EX-99.2 3 ea121603ex99-2_seabridge.htm TECHNICAL REPORT

Exhibit 99.2

Report to:
Seabridge Gold Inc. Northwestern British Columbia, Canada
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report
Effective Date: April 30, 2020
Prepared by:

William E. Threlkeld, P.Geo, P.G., Seabridge Gold Inc. Hassan Ghaffari, P.Eng., Tetra Tech, Inc. Jianhui (John) Huang, Ph.D., P.Eng., Tetra Tech, Inc. James H. Gray, P.Eng., Moose Mountain Technical Services Ross Hammett, Ph.D., P.Eng., Golder Associates Ltd. Brendon Masson, P.Eng., McElhanney Consulting Services Ltd. Derrek Kinakin, M.Sc., P.Geo., P.G., BGC Engineering Inc. Michael J. Lechner, P.Geo., RPG, CPG, Resource Modeling Inc. J. Graham Parkinson, P.Geo., Klohn Crippen Berger Ltd. Rolf Schmidt, P. Geo., ERM Consultants Canada Ltd. Neil Brazier, P.Eng., WN Brazier Associates Inc. Kirk Hanson, P.E., Wood Canada Ltd. Alan Keylock, P.Eng., Wood Canada Ltd. |

Important NOtice

This report was prepared as National Instrument 43-101 Technical Report for Seabridge Gold Inc. (Seabridge) by Tetra Tech, Inc., Moose Mountain Technical Services, Golder Associates Ltd., BGC Engineering Inc., Resource Modeling Inc., McElhanney Consulting Services Ltd., Klohn Crippen Berger Ltd, ERM Consultants Canada Ltd, WN Brazier Associates Inc., and Wood Canada Limited, collectively the Report Authors. The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in the Report Authors’ services, based on i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and qualifications set forth in this report. This report is intended for use by Seabridge subject to the respective terms and conditions of its contracts with the individual Report Authors. Those contracts permit Seabridge to file this report as a Technical Report with Canadian Securities Regulatory Authorities pursuant to provincial and territorial securities law. Except for the purposes legislated under Canadian provincial and territorial securities law, any other use of this report by any third party is at that party’s sole risk.

Seabridge Gold Inc. | ii | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Table of Contents

1.0 | Summary | 1-1
| 1.1 | Introduction | 1-1
| 1.2 | Key Study Outcomes | 1-2
|  | 1.2.1 | 2016 Preliminary Feasibility Study | 1-2
|  | 1.2.2 | 2020 Preliminary Economic Assessment | 1-4
| 1.3 | Property Description and Location | 1-5
| 1.4 | Accessibility, Climate, Local Resources, Physiography and Infrastructure | 1-8
| 1.5 | Geological Setting And Mineralization | 1-8
| 1.6 | History | 1-12
| 1.7 | Mineral Resources | 1-12
| 1.8 | 2016 PFS Data Verification | 1-15
| 1.9 | Mining Methods | 1-15
|  | 1.9.1 | Mineral Reserve Estimate | 1-15
|  | 1.9.2 | Mine Production Plan | 1-16
| 1.10 | Mineral Processing And Metallurgical Testing | 1-19
| 1.11 | 2016 PFS Recovery Methods | 1-19
| 1.12 | 2016 PFS Project Infrastructure | 1-21
|  | 1.12.1 | Geohazards | 1-21
|  | 1.12.2 | Tailings Management | 1-21
|  | 1.12.3 | Mine Site Water Management | 1-22
|  | 1.12.4 | Permanent Access Roads | 1-23
|  | 1.12.5 | Winter Access Road | 1-23
|  | 1.12.6 | Off-Site Infrastructure | 1-23
|  | 1.12.7 | Tunneling | 1-23
|  | 1.12.8 | Power Supply and Distribution | 1-24
| 1.13 | Environmental Studies, Permitting, and Social or Community Impact | 1-24
|  | 1.13.1 | Benefit Agreement | 1-25
|  | 1.13.2 | Closure and Reclamation | 1-25
| 1.14 | 2016 PFS Capital Cost Estimate | 1-25
| 1.15 | 2016 PFS Operating Cost Estimate | 1-27
| 1.16 | 2016 PFS Economic Evaluation | 1-28
|  | 1.16.1 | Sensitivity Analysis | 1-30
| 1.17 | 2020 Preliminary Economic Assessment | 1-32
|  | 1.17.1 | Introduction | 1-32
|  | 1.17.2 | Mining Methods | 1-33
|  | 1.17.3 | Open Pit Mining | 1-34
|  | 1.17.4 | Underground Mining | 1-35
|  | 1.17.5 | Mine Production Plan | 1-35

Seabridge Gold Inc. | iii | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

|  | 1.17.6 | Recovery Methods | 1-36
|  | 1.17.7 | Capital and Operating Costs | 1-36
|  | 1.17.8 | Economic Analysis | 1-37
|  | 1.17.9 | Conclusions | 1-38
| 1.18 | Recommendations | 1-39
|  | 1.18.1 | 2016 PFS Recommendations | 1-39
|  | 1.18.2 | 2020 PEA Recommendations | 1-39
2.0 | Introduction | 2-1
| 2.1 | Overview | 2-1
| 2.2 | Terms of Reference | 2-1
|  | 2.2.1 | 2016 Prefeasibility Study Mineral Resource Estimate Update | 2-1
|  | 2.2.2 | 2016 Prefeasibility Study Data Verification | 2-1
|  | 2.2.3 | 2020 Preliminary Economic Assessment Update | 2-2
| 2.3 | Sources of Information | 2-2
| 2.4 | Effective Dates | 2-2
| 2.5 | Qualified Persons | 2-2
| 2.6 | Site Visits | 2-2
3.0 | Reliance on Other Experts | 3-1
| 3.1 | 2016 Prefeasibility Study | 3-1
| 3.2 | 2020 Preliminary Economic Assessment | 3-2
4.0 | Property Description and Location | 4-1
5.0 | Accessibility, Climate, Infrastructure, Local Resources and Physiography | 5-1
6.0 | History | 6-1
| 6.1 | Exploration History | 6-1
| 6.2 | Historical Resource Estimates | 6-4
| 6.3 | History of Production | 6-4
7.0 | GEOLOGICAL SETTING AND MINERALIZATION | 7-1
| 7.1 | Geological Setting | 7-1
| 7.2 | Mineralization | 7-3
|  | 7.2.1 | Kerr Zone | 7-3
|  | 7.2.2 | Sulphurets Zone | 7-8
|  | 7.2.3 | Mitchell Zone | 7-13
|  | 7.2.4 | Iron Cap Zone | 7-18
8.0 | DEPOSIT TYPES | 8-1
9.0 | exploration | 9-1
| 9.1 | 2011 Geophysical Exploration Program | 9-1
|  | 9.1.1 | Results of 2011 Geophysical Program | 9-1
| 9.2 | 2013 Geophysical Exploration Program | 9-1
|  | 9.2.1 | Results of 2013 Geophysical Program | 9-2

Seabridge Gold Inc. | iv | 219221-01-RPT-001
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| 9.3 | 2014 Geophysical Exploration Program | 9-3
|  | 9.3.1 | Results of 2014 Geophysical Programs | 9-3
| 9.4 | 2015 Geophysical Exploration Program | 9-3
| 9.5 | 2019 Geophysical Program | 9-4
10.0 | Drilling | 10-1
| 10.1 | Introduction | 10-1
| 10.2 | Type and Extent of drilling | 10-4
| 10.3 | Drilling Procedures | 10-10
| 10.4 | QP Comments Regarding Drilling and Sampling Factors | 10-13
11.0 | Sample Preparation, Analysis and Security | 11-1
| 11.1 | Introduction | 11-1
| 11.2 | KSM Sample Preparation Methods and Procedures | 11-1
|  | 11.2.1 | Statement on Sample Preparation Personnel | 11-1
|  | 11.2.2 | Sample Preparation and Dispatch | 11-1
|  | 11.2.3 | Pre-2012 Analytical Procedures | 11-2
|  | 11.2.4 | Post-2012 Analytical Procedures | 11-3
| 11.3 | Summary of the Nature, Extent, and Results of Quality Control Procedures | 11-4
|  | 11.3.1 | Pre-2012 Quality Control Procedures | 11-9
|  | 11.3.2 | Post-2012 Quality Control Procedures | 11-10
| 11.4 | QP’s Opinion | 11-10
12.0 | Data Verification | 12-1
| 12.1 | Drill Hole Data Verification | 12-1
|  | 12.1.1 | Assay Verification | 12-1
|  | 12.1.2 | Drill Hole Logs | 12-2
|  | 12.1.3 | Drill Hole Collar Locations | 12-2
|  | 12.1.4 | Down-hole Surveys | 12-3
|  | 12.1.5 | Quality Assurance/Quality Control | 12-3
|  | 12.1.6 | Topographic Data | 12-3
|  | 12.1.7 | Bulk Density Data | 12-4
| 12.2 | Post Model Drilling Results | 12-4
| 12.3 | Verification that 2016 Prefeasibility Study Remains Current | 12-5
|  | 12.3.1 | Verification Checks on the PFS Mine Quantities | 12-6
|  | 12.3.2 | Verification Checks on Financial Analysis and Cost Estimates | 12-7
| 12.4 | QP’s Opinion | 12-8
|  | 12.4.1 | Drill Hole Data Verification | 12-8
|  | 12.4.2 | 2016 Prefeasibility Study Data Verification | 12-8
|  | 12.4.3 | 2020 PEA vs 2016 PFS | 12-8
13.0 | Mineral Processing and Metallurgical Testing | 13-1
| 13.1 | Introduction | 13-1
| 13.2 | Summary of Metallurgical Test Programs | 13-3
| 13.3 | Summary of Initial Test Work 1989–1991 | 13-5
| 13.4 | Summary of Test Work 2007–2016 | 13-5
|  | 13.4.1 | Test Programs | 13-5

Seabridge Gold Inc. | v | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

|  | 13.4.2 | Baseline Test Process Flowsheet and Conditions | 13-6
|  | 13.4.3 | Mitchell Zone Major Metallurgical Test Results | 13-6
|  | 13.4.4 | Sulphurets Zone Major Metallurgical Test Results | 13-19
|  | 13.4.5 | Upper Kerr Zone Metallurgical Test Results | 13-22
|  | 13.4.6 | Deep Kerr Zone Major Metallurgical Test Results | 13-24
|  | 13.4.7 | Iron Cap Zone Major Metallurgical Test Results | 13-28
| 13.5 | Recent Test Work 2017–2020 | 13-34
|  | 13.5.1 | Deep Kerr Zone Metallurgical Test Work (2016/2017) | 13-34
|  | 13.5.2 | Iron Cap Zone Metallurgical Test Work (2017–2020) | 13-43
|  | 13.5.3 | Flotation Concentrate Assay (2007–2020) | 13-53
|  | 13.5.4 | Ancillary Tests | 13-55
| 13.6 | Conclusions | 13-60
| 13.7 | Metallurgical Performance Projection | 13-61
|  | 13.7.1 | Metallurgical Performance Projection – Mitchell, Sulphurets, Upper Kerr, and Upper Iron Cap | 13-62
|  | 13.7.2 | Metallurgical Performance Projection – Deep Kerr and Lower Iron Cap | 13-65
14.0 | Mineral Resource Estimates | 14-1
| 14.1 | Kerr Deposit | 14-2
|  | 14.1.1 | Grade Distribution – Kerr Deposit | 14-2
|  | 14.1.2 | Assay Grade Capping – Kerr Deposit | 14-4
|  | 14.1.3 | Drill Hole Compositing – Kerr Deposit | 14-6
|  | 14.1.4 | Geologic Constraints - Kerr Deposit | 14-6
|  | 14.1.5 | Variography – Kerr Deposit | 14-6
|  | 14.1.6 | Grade Estimation Parameters – Kerr Deposit | 14-9
|  | 14.1.7 | Grade Model Verification - Kerr Deposit | 14-9
|  | 14.1.8 | Resource Classification – Kerr Deposit | 14-15
| 14.2 | Sulphurets Deposit | 14-15
|  | 14.2.1 | Grade Distribution – Sulphurets Deposit | 14-15
|  | 14.2.2 | Assay Grade Capping – Sulphurets Deposit | 14-17
|  | 14.2.3 | Drill Hole Compositing – Sulphurets Deposit | 14-19
|  | 14.2.4 | Geologic Constraints – Sulphurets Deposit | 14-19
|  | 14.2.5 | Variography – Sulphurets Deposit | 14-19
|  | 14.2.6 | Grade Estimation Parameters – Sulphurets Deposit | 14-22
|  | 14.2.7 | Model Validation – Sulphurets Deposit | 14-22
|  | 14.2.8 | Resource Classification – Sulphurets Deposit | 14-28
| 14.3 | Mitchell Deposit | 14-28
|  | 14.3.1 | Metal Distribution – Mitchell Deposit | 14-28
|  | 14.3.2 | Assay Grade Capping – Mitchell Deposit | 14-30
|  | 14.3.3 | Drill Hole Compositing – Mitchell Deposit | 14-32
|  | 14.3.4 | Variography – Mitchell Deposit | 14-32
|  | 14.3.5 | Grade Estimation Parameters – Mitchell Deposit | 14-35
|  | 14.3.6 | Grade Model Validation – Mitchell Deposit | 14-35
|  | 14.3.7 | Resource Classification – Mitchell Deposit | 14-41
| 14.4 | Iron Cap Deposit | 14-41
|  | 14.4.1 | Grade Distribution – Iron Cap Deposit | 14-41
|  | 14.4.2 | Assay Grade Capping – Iron Cap Deposit | 14-43

Seabridge Gold Inc. | vi | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

|  | 14.4.3 | Drill Hole Compositing – Iron Cap Deposit | 14-45
|  | 14.4.4 | Geologic Constraints – Iron Cap Deposit | 14-45
|  | 14.4.5 | Variography – Iron Cap Deposit | 14-45
|  | 14.4.6 | Grade Estimation Parameters – Iron Cap Deposit | 14-47
|  | 14.4.7 | Grade Model Verification – Iron Cap Deposit | 14-48
|  | 14.4.8 | Resource Classification – Iron Cap Deposit | 14-54
| 14.5 | Bulk Density | 14-54
| 14.6 | Resource Criteria | 14-54
| 14.7 | Summary of KSM Mineral Resources | 14-55
| 14.8 | General Discussion | 14-56
15.0 | Mineral Reserve Estimates | 15-1
| 15.1 | Introduction | 15-1
| 15.2 | Open Pit Reserve Parameters | 15-1
| 15.3 | Underground Mining Reserve Parameters | 15-2
| 15.4 | Mineral Reserves | 15-3
| 15.5 | Factors that could affect the mineral reserve estimate | 15-4
16.0 | Mining Methods | 16-1
| 16.1 | Introduction | 16-1
|  | 16.1.1 | Production Rate Consideration | 16-1
| 16.2 | Open Pit Mining Operations | 16-1
|  | 16.2.1 | Introduction | 16-1
|  | 16.2.2 | Mining Datum | 16-1
|  | 16.2.3 | Open Pit Production Rate Considerations | 16-2
|  | 16.2.4 | Open Pit Mine Planning 3D Block Model | 16-2
|  | 16.2.5 | Pit Slope Design Angles | 16-4
|  | 16.2.6 | Economic Pit Limits, Pit Designs | 16-9
|  | 16.2.7 | Detailed Pit Designs | 16-14
|  | 16.2.8 | Open Pit Mine Plan | 16-18
|  | 16.2.9 | Open Pit Production | 16-22
|  | 16.2.10 | Open Pit Mine Operations | 16-27
|  | 16.2.11 | Mine Closure and Reclamation | 16-30
|  | 16.2.12 | Open Pit Mine Equipment | 16-30
| 16.3 | Underground Mining Operations | 16-35
|  | 16.3.1 | Underground Mine Design Inputs | 16-36
|  | 16.3.2 | Mitchell Underground | 16-38
|  | 16.3.3 | Iron Cap Underground | 16-41
| 16.4 | Mine Production Schedule | 16-45
17.0 | Recovery Methods | 17-1
| 17.1 | Introduction | 17-1
| 17.2 | Major Process Design Criteria | 17-5
| 17.3 | Process Plant Description | 17-7
|  | 17.3.1 | Primary Crushing | 17-7
|  | 17.3.2 | Coarse Ore Transport From Mitchell Site to Treaty Site | 17-8
|  | 17.3.3 | Coarse Material Handling | 17-8

Seabridge Gold Inc. | vii | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

|  | 17.3.4 | Secondary Crushing | 17-9
|  | 17.3.5 | Tertiary Crushing Material Conveyance/Storage | 17-9
|  | 17.3.6 | Tertiary Crushing | 17-9
|  | 17.3.7 | Primary Grinding | 17-10
|  | 17.3.8 | Copper, Gold and Molybdenum Flotation | 17-10
|  | 17.3.9 | Concentrate Dewatering | 17-12
|  | 17.3.10 | Gold Recovery From Gold-bearing Pyrite Products | 17-12
|  | 17.3.11 | Treatment of Leach Residues | 17-14
|  | 17.3.12 | Tailing Management | 17-15
|  | 17.3.13 | Reagents Handling | 17-16
|  | 17.3.14 | Water Supply | 17-17
|  | 17.3.15 | Air Supply | 17-18
|  | 17.3.16 | Assay and Metallurgical Laboratory | 17-18
|  | 17.3.17 | Process Control and Instrumentation | 17-19
| 17.4 | Yearly Production Projection | 17-20
18.0 | Project Infrastructure | 18-1
| 18.1 | Site Layout | 18-1
| 18.2 | Tailing, Mine Rock, and Water Management | 18-4
|  | 18.2.1 | Introduction | 18-4
|  | 18.2.2 | Mine Site Characterization | 18-5
|  | 18.2.3 | TMF Site Characterization | 18-9
|  | 18.2.4 | Rock Storage Facilities | 18-12
|  | 18.2.5 | Mine Site Water Management | 18-12
|  | 18.2.6 | Water Treatment | 18-18
|  | 18.2.7 | Tailing Management Facility Design | 18-20
| 18.3 | Tunnels | 18-32
|  | 18.3.1 | Mitchell-Treaty Tunnels | 18-33
| 18.4 | Mine To Mill Ore Transport System | 18-36
|  | 18.4.1 | MTT Freight and Personnel Transport | 18-39
| 18.5 | Site Roads | 18-41
|  | 18.5.1 | Road Width | 18-42
| 18.6 | Ancillary Buildings | 18-43
|  | 18.6.1 | Treaty OPC | 18-44
|  | 18.6.2 | Mine Site | 18-45
| 18.7 | Sewage | 18-45
| 18.8 | Communications System | 18-46
| 18.9 | Fresh and Potable Water Supply | 18-46
| 18.10 | Power Supply and Primary Distribution | 18-46
|  | 18.10.1 | Northwest Transmission Line | 18-47
|  | 18.10.2 | Treaty Creek Switching Station | 18-48
|  | 18.10.3 | Transmission Line Extension To KSM | 18-48
|  | 18.10.4 | System Studies | 18-49
|  | 18.10.5 | Electric Utility Requirements, Tariffs, And Cost of Electric Power | 18-50
|  | 18.10.6 | Treaty Plant Main Substation No. 1 | 18-51
|  | 18.10.7 | 138 kV Cable | 18-52
|  | 18.10.8 | Mitchell Substation No. 2 | 18-52

Seabridge Gold Inc. | viii | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

|  | 18.10.9 | Site Power Distribution | 18-53
|  | 18.10.10 | Mine Power | 18-53
|  | 18.10.11 | Construction and Standby Power | 18-53
|  | 18.10.12 | Energy Recovery and Self Generation | 18-54
| 18.11 | Treaty OPC and Mine Site Secondary Electrical Power Distribution And Utilization | 18-55
|  | 18.11.1 | Mine And Plant Power Consumption | 18-55
|  | 18.11.2 | Power Distribution – Treaty Plant Main Substation No. 1 | 18-55
|  | 18.11.3 | Mitchell Substation No. 2 | 18-56
| 18.12 | Permanent and Construction Access Roads | 18-57
|  | 18.12.1 | Route Descriptions | 18-59
|  | 18.12.2 | Winter Access Road | 18-61
|  | 18.12.3 | Road Design Requirements | 18-63
|  | 18.12.4 | Design Progress | 18-63
| 18.13 | Logistics | 18-66
| 18.14 | Preliminary Construction Execution Plan | 18-67
|  | 18.14.1 | Introduction | 18-67
|  | 18.14.2 | Early Works Plan | 18-67
|  | 18.14.3 | construction Scope | 18-68
|  | 18.14.4 | Construction Schedule | 18-70
|  | 18.14.5 | Engineering and Procurement | 18-71
|  | 18.14.6 | Construction Management | 18-72
|  | 18.14.7 | Construction Supervision and Contractor Management | 18-72
|  | 18.14.8 | Contracting Packaging and Strategy Overview | 18-72
|  | 18.14.9 | Site Organization Structure | 18-73
|  | 18.14.10 | Environmental and Community Affairs | 18-73
|  | 18.14.11 | Pre-commissioning/Commissioning | 18-74
| 18.15 | Owner’s Implementation Plan | 18-74
19.0 | Market Studies and Contracts | 19-1
| 19.1 | Copper Concentrate | 19-1
|  | 19.1.1 | Marketability | 19-1
|  | 19.1.2 | Smelting Terms | 19-1
|  | 19.1.3 | Copper Concentrates Contracts and Terms | 19-2
| 19.2 | Molybdenite Concentrate | 19-5
|  | 19.2.1 | Smelting Charge | 19-5
| 19.3 | Gold and Silver Doré | 19-6
20.0 | Environmental Studies, Permitting, and Social or Community Impact | 20-1
| 20.1 | Licensing and Permitting | 20-1
|  | 20.1.1 | Provincial Process | 20-2
|  | 20.1.2 | Federal Process | 20-3
|  | 20.1.3 | Provincial Permits | 20-4
|  | 20.1.4 | Federal Permits | 20-5
|  | 20.1.5 | benefits agreement | 20-6
| 20.2 | Environmental Settings and Studies | 20-7
|  | 20.2.1 | Biophysical Setting | 20-7

Seabridge Gold Inc. | ix | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

|  | 20.2.2 | Economic, Social, and Cultural Setting | 20-9
| 20.3 | Water Management | 20-11
|  | 20.3.1 | Overview of Water Management | 20-11
|  | 20.3.2 | Summary of Water Management Plan | 20-13
| 20.4 | Waste Management | 20-14
|  | 20.4.1 | Tailing Management Facility Management and Monitoring Plan | 20-14
|  | 20.4.2 | Best Available Tailings Technology Assessment | 20-15
|  | 20.4.3 | Waste Rock Management | 20-15
|  | 20.4.4 | Domestic and Industrial non-hazardous and hazardous Waste Management | 20-16
| 20.5 | Air Quality Management including Greenhouse Gases | 20-17
| 20.6 | Environmental Management System | 20-17
| 20.7 | Closure and Reclamation | 20-18
|  | 20.7.1 | Closure and Reclamation Objectives | 20-18
|  | 20.7.2 | Soil Handling Plan | 20-18
|  | 20.7.3 | Closure and Reclamation Planning | 20-19
21.0 | Capital and Operating Cost Estimates | 21-1
| 21.1 | Initial Capital Costs | 21-1
|  | 21.1.1 | Exclusions | 21-3
|  | 21.1.2 | Direct Costs | 21-3
|  | 21.1.3 | Indirect Costs | 21-5
|  | 21.1.4 | Owner’s Costs | 21-6
|  | 21.1.5 | Contingency | 21-6
| 21.2 | Sustaining Capital Costs | 21-6
|  | 21.2.1 | Mine Site | 21-7
|  | 21.2.2 | Open Pit Mining | 21-7
|  | 21.2.3 | Underground Mining (Block Caves) | 21-7
|  | 21.2.4 | Mine Site Water Treatment | 21-8
|  | 21.2.5 | Process | 21-8
|  | 21.2.6 | Northwest Transmission Line Contribution | 21-9
|  | 21.2.7 | McTagg Diversion Tunnel Mini Hydro Generation Station | 21-9
|  | 21.2.8 | Tailing Management Facility | 21-9
|  | 21.2.9 | Other Sustaining Capital Costs | 21-10
| 21.3 | Operating Costs | 21-10
|  | 21.3.1 | Open Pit Mine Operating Costs | 21-12
|  | 21.3.2 | Underground Mining Operating Costs | 21-13
|  | 21.3.3 | Process Operating Costs | 21-14
|  | 21.3.4 | TMF Dam Management Operating Costs | 21-17
|  | 21.3.5 | Mine Site Water Management Costs | 21-17
|  | 21.3.6 | General and Administrative | 21-17
|  | 21.3.7 | Site Services | 21-18
22.0 | Economic Analysis | 22-1
| 22.1 | Introduction | 22-1
| 22.2 | Forward-looking Statements | 22-2
| 22.3 | Pre-Tax Model | 22-2

Seabridge Gold Inc. | x | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

|  | 22.3.1 | Financial Evaluations: NPV and IRR | 22-3
|  | 22.3.2 | Metal Price Scenarios | 22-4
| 22.4 | Sensitivity Analysis | 22-5
| 22.5 | Post-tax Financial Evaluations | 22-7
|  | 22.5.1 | Canadian Federal and BC Provincial Income Tax Regime | 22-7
|  | 22.5.2 | BC Mineral Tax Regime | 22-8
|  | 22.5.3 | Taxes and Post-tax Financial Results | 22-9
| 22.6 | Royalties | 22-11
| 22.7 | Smelter Terms | 22-11
| 22.8 | Miscellaneous Costs and Charges | 22-11
23.0 | Adjacent Properties | 23-1
24.0 | Other Relevant Data and Information | 24-1
| 24.16 | Mining Methods | 24-3
|  | 24.16.1 | Net Smelter Return Block Model | 24-3
|  | 24.16.2 | Open Pit Mining Method | 24-4
|  | 24.16.3 | Iron Cap Mining Methods | 24-23
|  | 24.16.4 | Deep Kerr Mining Methods | 24-36
|  | 24.16.5 | Mine Production Schedule (Open Pit and Underground Combined) | 24-51
| 24.17 | Recovery Methods | 24-54
|  | 24.17.1 | Process Plant | 24-56
|  | 24.17.2 | flowsheet description | 24-56
|  | 24.17.3 | Mill Design Criteria | 24-64
| 24.18 | Project Infrastructure | 24-67
|  | 24.18.1 | On-Site Infrastructure | 24-67
|  | 24.18.2 | Off-site Infrastructure | 24-73
|  | 24.18.2.1 | Electrical Power Supply | 24-73
|  | 24.18.2.2 | Access Roads | 24-74
|  | 24.18.2.3 | Logistics | 24-74
|  | 24.18.3 | Iron Cap Project Infrastructure | 24-74
|  | 24.18.4 | Deep Kerr Project Infrastructure | 24-76
| 24.19 | Market Studies and Contracts | 24-79
| 24.20 | Environmental Studies, Permitting and Social or Community Impact | 24-82
| 24.21 | Capital and Operating Cost Estimates | 24-83
|  | 24.21.1 | Capital Cost Estimate | 24-84
|  | 24.21.2 | Operating Cost Estimate | 24-86
| 24.22 | Economic Analysis | 24-91
|  | 24.22.1 | Methodology Used | 24-92
|  | 24.22.2 | Financial Model Parameters | 24-92
|  | 24.22.3 | Sensitivity Analysis | 24-96
25.0 | Interpretations and Conclusions | 25-1
| 25.1 | Introduction | 25-1
| 25.2 | 2016 Prefeasibility Study Conclusions | 25-2
|  | 25.2.1 | 2016 PFS Data Verification | 25-2
|  | 25.2.2 | Mineral Reserves | 25-2

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|  | 25.2.3 | Mining Methods | 25-2
|  | 25.2.3.1 | Open Pit Mining | 25-2
|  | 25.2.3.2 | Underground Mining | 25-3
|  | 25.2.4 | Recovery Methods | 25-3
|  | 25.2.5 | Project Infrastructure | 25-4
|  | 25.2.5.1 | MTT Transportation System | 25-4
|  | 25.2.5.2 | Tunneling | 25-4
|  | 25.2.5.3 | Infrastructure Dams | 25-5
|  | 25.2.6 | Economic Analysis | 25-6
| 25.3 | 2020 Preliminary Economic Assessment Conclusions | 25-7
|  | 25.3.1 | Mining Methods | 25-7
|  | 25.3.1.1 | Open Pit Mining | 25-7
|  | 25.3.1.2 | Underground Mining | 25-8
|  | 25.3.2 | Recovery Methods | 25-8
|  | 25.3.3 | Project Infrastructure | 25-9
|  | 25.3.4 | Economic Analysis | 25-10
| 25.4 | 2016 Prefeasibility Study Risks | 25-11
|  | 25.4.1 | Open Pit Mining | 25-12
|  | 25.4.2 | Underground Mining | 25-13
|  | 25.4.3 | Tunnels | 25-16
|  | 25.4.4 | Construction Critical Path | 25-17
|  | 25.4.5 | Metallurgical Performance and Process | 25-17
| 25.5 | 2020 Preliminary Economic Assessment Risks | 25-18
|  | 25.5.1 | Open Pit Mining | 25-18
|  | 25.5.2 | Underground Mining | 25-18
|  | 25.5.3 | Tunnels | 25-20
|  | 25.5.4 | Metallurgical Performance and Process | 25-20
|  | 25.5.5 | Marketing Risks | 25-20
26.0 | Recommendations | 26-1
| 26.1 | Introduction | 26-1
| 26.2 | 2016 Prefeasibility Study Recommendations | 26-1
|  | 26.2.1 | Open Pit Mining and Reserves | 26-1
|  | 26.2.2 | Pit Slopes | 26-2
|  | 26.2.3 | Underground Mining and Reserves | 26-3
|  | 26.2.4 | Rock Storage Facilities | 26-5
|  | 26.2.5 | Metallurgical Testing and Process Engineering | 26-5
|  | 26.2.6 | Water Management | 26-6
|  | 26.2.7 | TMF Area | 26-8
|  | 26.2.8 | Tunnels | 26-9
|  | 26.2.9 | MTT Transport System | 26-11
| 26.3 | 2020 Preliminary Economic Assessment Recommendations | 26-11
|  | 26.3.1 | Geology | 26-12
|  | 26.3.2 | Open Pit Mining | 26-13
|  | 26.3.3 | Underground Mining & Geotechnical | 26-14
|  | 26.3.4 | Iron Cap and Deep Kerr Metallurgical Testwork | 26-16
|  | 26.3.5 | Rock Storage Facilities | 26-17
|  | 26.3.6 | Water Management | 26-17
|  | 26.3.7 | TMF Area | 26-18
|  | 26.3.8 | Tunnels | 26-18
|  | 26.3.9 | Power Supply and Energy Recovery Studies | 26-19
27.0 | References | 27-1
28.0 | Certificates of Qualified Persons | 28-1

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List of Tables

Table 1.1 | 2020 PEA Projected Economic Results | 1-4
Table 1.2 | KSM Mineral Resources | 1-14
Table 1.3 | Summary of the 2016 PFS Verification Checks | 1-15
Table 1.4 | KSM Proven and Probable Reserves as of July 31, 2016 | 1-16
Table 1.5 | 2016 PFS Initial Capital Cost Summary | 1-26
Table 1.6 | 2016 PFS Average Operating Cost Summary | 1-27
Table 1.7 | 2016 PFS Metal Production | 1-29
Table 1.8 | 2016 PFS Summary of the Pre- and Post-tax Economic Evaluations | 1-29
Table 1.9 | Pit Delineated PEA Mine Plan Quantities | 1-34
Table 1.10 | Production Tonnes and Grade in 2020 PEA Underground Mine Plan | 1-35
Table 1.11 | 2020 PEA Capital Cost Estimate Summary in US$M | 1-37
Table 1.12 | 2020 PEA LOM Average Unit Operating Costs | 1-37
Table 1.13 | 2020 PEA Economic Results | 1-38
Table 4.1 | KSM Mineral Claims and Leases | 4-5
Table 4.2 | KSM Placer Claims | 4-6
Table 4.3 | Seabee/Tina KSM Claims | 4-7
Table 6.1 | Exploration Summary of the Kerr Zone | 6-2
Table 6.2 | Exploration Summary of the Sulphurets, Mitchell, Iron Cap Zones, and other Exploration Targets | 6-3
Table 10.1 | KSM Historic Drilling Through 2019 | 10-2
Table 10.2 | Historic Drilling by Company Through 2018 | 10-2
Table 10.3 | KSM Drill Hole Summary by Area and Company Through 2018 | 10-3
Table 11.1 | ICP Detection Limits – Pre-2012 Data | 11-3
Table 11.2 | ICP Detection Limits – Post 2012 | 11-4
Table 11.3 | Summary of KSM Control Samples Submitted Thru Time | 11-5
Table 11.4 | ¼ Core Duplicate Sample Statistics (2007-2018) | 11-5
Table 12.1 | Reserve Assessment Variance Check from 2016 PFS to 2020 Information | 12-6
Table 12.2 | Summary of the 2016 PFS Verification Checks | 12-7
Table 13.1 | Typical Mineralogical Characteristics and Average Copper-to-Gold Grade Ratios in Planned Mill Feed | 13-1
Table 13.2 | Average Ball Mill Grindability and Abrasion Index | 13-2
Table 13.3 | Locked Cycle Flotation Test Result Summary | 13-2
Table 13.4 | Metallurgical Test Work Programs | 13-3
Table 13.5 | Test Samples – Mitchell (2007–2012) | 13-7
Table 13.6 | JK SimMet Simulation Results (60,000 t/d SABC Circuit, 2008) | 13-8
Table 13.7 | Locked Cycle Test Results – Mitchell | 13-13
Table 13.8 | Locked Cycle Test Results – Blended Samples (Mitchell and Other Deposits) | 13-14
Table 13.9 | Cu-Mo Separation LCT Results, 2010 | 13-15
Table 13.10 | Cyanidation Test Results on LCT Products – Mitchell | 13-16
Table 13.11 | Metal Contents of Leach Test Head Samples – Mitchell, 2017 (ALS KM5367) | 13-17
Table 13.12 | Metal Contents of Leach Test Head Samples – Mitchell, 2018 (ALS KM5367) | 13-17
Table 13.13 | Optimization Cyanidation Test Results – Mitchell 2017 (ALS) | 13-18
Table 13.14 | Test Samples – Sulphurets (2009-2012) | 13-20
Table 13.15 | LCT Results – Sulphurets | 13-21
Table 13.16 | Cyanidation Test Results – Flotation LCT Products, Sulphurets, 2009–2011 | 13-22
Table 13.17 | Metal Contents of Composites – Upper Kerr, 2010 (G&T) | 13-23

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Table 13.18 | LCT Results – Upper Kerr (G&T) | 13-24
Table 13.19 | Cyanidation Test Results on LCT Products – Upper Kerr (G&T) | 13-24
Table 13.20 | Individual Test Samples – Deep Kerr (2012–2016) | 13-25
Table 13.21 | Master Test Samples – Deep Kerr (2012–2016) | 13-25
Table 13.22 | Flotation LCT Results – Deep Kerr | 13-27
Table 13.23 | Preliminary Cyanidation Test Results – Deep Kerr | 13-28
Table 13.24 | Individual Test Samples – Iron Cap (2010–2015) | 13-29
Table 13.25 | Master Test Samples – Iron Cap (2010–2015) | 13-29
Table 13.26 | Locked Cycle Test Results – Iron Cap | 13-33
Table 13.27 | Cyanidation Test Results on LCT Products – Iron Cap | 13-34
Table 13.28 | Metal Contents of Composites – Deep Kerr, 2016/2017 (ALS KM5063) | 13-35
Table 13.29 | Metal Contents of Composites – Deep Kerr, 2017 (ALS KM5266) | 13-36
Table 13.30 | Bond Ball Mill Work Index Test Results – Deep Kerr, 2017 (ALS) | 13-36
Table 13.31 | Mineral Composition Data – Deep Kerr 2016/2017 (ALS) | 13-37
Table 13.32 | Flotation Locked Cycle Test Results – Deep Kerr 2016/2017 | 13-41
Table 13.33 | Preliminary Cyanidation Test Results – Deep Kerr | 13-42
Table 13.34 | Metal Contents of Composites – Iron Cap, 2017/2018 (ALS KM5248/5501) | 13-44
Table 13.35 | Bond Ball Mill Work Index Test Results – Deep Kerr, 2018 (ALS) | 13-45
Table 13.36 | Mineral Composition Data – Iron Cap 2017-2020 (ALS) | 13-46
Table 13.37 | Flotation Locked Cycle Test Results – Iron Cap 2017 -2020 (KM5248/KM5501/KM5806/KM6004) | 13-50
Table 13.38 | Preliminary Cyanidation Test Results – Iron Cap (ALS 2017) | 13-52
Table 13.39 | Flotation Concentrate Assay from Different Deposit Samples | 13-54
Table 13.40 | Flotation Concentrate Assay from Blended Head Samples | 13-55
Table 13.41 | Chemical Analysis of Cyanide Recovery Test Solution and Cyanide Destruction Pulp | 13-56
Table 13.42 | Cyanide Recovery Test Results – AVR | 13-56
Table 13.43 | Cyanide Destruction Test Results – 2009/2010 (SGS) | 13-57
Table 13.44 | Recommended Conventional Thickener Operating Parameters – 2009 (Pocock) | 13-59
Table 13.45 | Recommended High Rate Thickener Operating Parameters – 2009 (Pocock) | 13-59
Table 13.46 | Filtration Test Results – 2009 (Pocock) | 13-60
Table 13.47 | Cu-Au Flotation Concentrate Grade Versus Cu Head Grade | 13-62
Table 13.48 | Cu-Au Flotation Concentrate – Metal Recovery Projections | 13-63
Table 13.49 | Au-Ag Doré – Cyanide Leach Metal Recovery Projections | 13-64
Table 13.50 | Mo Flotation Concentrate Metal Recovery and Grade (Molybdenum Concentrate Grade = 50%) | 13-65
Table 13.51 | Cu-Au Flotation Concentrate –Metal Recovery Projections | 13-65
Table 13.52 | Au-Ag Doré – Cyanide Leach Metal Recovery Projections | 13-66
Table 13.53 | Mo Flotation Concentrate Metal Recovery and Grade (Molybdenum Concentrate Grade = 50%) | 13-66
Table 14.1 | Summary of KSM Block Model Parameters by Deposit | 14-1
Table 14.2 | Distribution of Gold by AUZON – Kerr Deposit | 14-3
Table 14.3 | Distribution of Copper by CUZON – Kerr Deposit | 14-4
Table 14.4 | Distribution of Gold by AUZON – Sulphurets Deposit | 14-16
Table 14.5 | Distribution of Copper by CUZON – Sulphurets Deposit | 14-17
Table 14.6 | Distribution of Gold by AUZON – Mitchell Deposit | 14-29
Table 14.7 | Distribution of Copper by CUZON – Mitchell Deposit | 14-30
Table 14.8 | Distribution of Gold by AUZON – Iron Cap Deposit | 14-42
Table 14.9 | Distribution of Copper by CUZON – Iron Cap Deposit | 14-43
Table 14.10 | Key Mineral Resource Parameters | 14-54

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Table 14.11 | KSM Mineral Resources | 14-56
Table 15.1 | Pit Mining Loss and Dilution | 15-1
Table 15.2 | Grade of Dilution Material by Pit Area | 15-1
Table 15.3 | Site Operating Cost – Drawpoint Shut-off | 15-2
Table 15.4 | Underground Mining Dilution | 15-3
Table 15.5 | Proven and Probable Reserves | 15-3
Table 16.1 | Major Smelter Terms Used in the NSR Calculation | 16-3
Table 16.2 | Metal Prices for Reserve NSR Calculation | 16-3
Table 16.3 | Estimated NSP by Mining Area | 16-3
Table 16.4 | LG Pit Limit Primary Assumptions | 16-9
Table 16.5 | Production Schedule Assumptions | 16-20
Table 16.6 | Blasting Assumptions | 16-29
Table 16.7 | Major Equipment Requirements | 16-31
Table 16.8 | Open Pit Production Drilling Assumptions | 16-31
Table 16.9 | Mine Support Equipment Fleet | 16-32
Table 16.10 | Open Pit Ancillary Equipment Fleet | 16-33
Table 16.11 | Mitchell Block Cave Mineral Reserves ($15/t NSR Shut-off) | 16-38
Table 16.12 | Iron Cap Block Cave Reserves ($16/t NSR Shut-off) | 16-42
Table 16.13 | Summarized Production Schedule – Open Pit and Underground | 16-46
Table 17.1 | Major Design Criteria | 17-5
Table 17.2 | Projected Metallurgical Performance | 17-21
Table 17.3 | Projected Copper Concentrate Quality | 17-22
Table 18.1 | Climate Data for the Mine Site (Sulphurets Creek Climate Station) | 18-8
Table 18.2 | Climate Data for the TMF (Teigen Creek Climate Station)1 | 18-11
Table 18.3 | Temporary Water Treatment Plant Locations | 18-19
Table 18.4 | Annual Reagent Consumption for the HDS WTP | 18-20
Table 18.5 | Tailing Dam Summary | 18-24
Table 18.6 | KSM Pre-Production Tunnels Summary | 18-32
Table 18.7 | KSM Operational Phase Tunnels Summary | 18-33
Table 18.8 | Mini Hydro and Energy Recovery Power Generation | 18-54
Table 18.9 | Owner’s Key Activities by Year | 18-77
Table 19.1 | Benchmark Smelting Terms | 19-2
Table 21.1 | Initial Capital Cost Summary | 21-2
Table 21.2 | Mine Site Capital Costs | 21-4
Table 21.3 | Process-Treaty OPC Capital Cost Estimate | 21-4
Table 21.4 | Indirect Capital Costs | 21-5
Table 21.5 | Sustaining Capital Costs | 21-6
Table 21.6 | Mine Site Sustaining Capital Costs | 21-7
Table 21.7 | Operating Cost Summary | 21-11
Table 21.8 | Summary of Process Operating Costs by Deposit | 21-15
Table 21.9 | Operating Costs per Area of Operation by Deposit | 21-16
Table 22.1 | Metal Production from the KSM Mine | 22-2
Table 22.2 | Summary of the Pre-tax Economic Evaluations | 22-4
Table 22.3 | Component of the Various Taxes for all Scenarios | 22-9
Table 22.4 | Summary of Post-tax Financial Results | 22-9
Table 22.5 | 2016 PFS Annual Cash Flow for Pre-production Period, Years 1 to 7 and LOM | 22-10
Table 23.1 | Pretium Snowfield Mineral Resources Using a 0.30 g/t Cut-off | 23-1
Table 24.1 | Metal Prices for 2020 PEA NSR Calculation | 24-3
Table 24.2 | Major Smelter Terms Used in the 2020 PEA NSR Calculation | 24-3
Table 24.3 | LG Pit Limit Primary Assumptions | 24-5
Table 24.4 | Pit Delineated Resource | 24-11

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Table 24.5 | Open Pit Production Schedule Assumptions | 24-14
Table 24.6 | Blasting Assumptions | 24-18
Table 24.7 | Peak Major Equipment Requirements | 24-19
Table 24.8 | Peak Mine Support Equipment Fleet | 24-20
Table 24.9 | Open Pit Ancillary Equipment Fleet | 24-20
Table 24.10 | Iron Cap Footprint Finder Input Parameters (CAD) | 24-25
Table 24.11 | Main Drift Types and Dimensions | 24-30
Table 24.12 | Vertical Development Dimensions | 24-31
Table 24.13 | Iron Cap Production Tonnes And Grade in PEA Mine Plan | 24-33
Table 24.14 | Mobile Equipment Requirements for Iron Cap (KSM operation Year 14) | 24-35
Table 24.15 | Deep Kerr Footprint Finder Input Parameters (CAD) | 24-37
Table 24.16 | Main Drift Types and Dimensions | 24-45
Table 24.17 | Vertical Development Dimensions | 24-46
Table 24.18 | Deep Kerr Production Tonnes and Grade in PEA Mine Plan | 24-49
Table 24.19 | Mobile Equipment Requirements for Deep Kerr (Year 17) | 24-50
Table 24.20 | Summarized Production Schedule – Open Pit and Underground | 24-52
Table 24.21 | Major Process Design Criteria | 24-64
Table 24.22 | Comminution – Major Equipment Sizing | 24-65
Table 24.23 | Copper-Gold Flotation – Major Equipment List | 24-66
Table 24.24 | Pyrite Flotation & Cyanidation – Major Equipment List | 24-67
Table 24.25 | Capital Cost Estimate Responsibilities by Firm | 24-84
Table 24.26 | 2020 PEA Capital Cost Estimate Summary | 24-85
Table 24.27 | 2019 Q4 PEA Commodity Pricing | 24-86
Table 24.28 | Operating Cost Estimate Responsibilities by Firm | 24-86
Table 24.29 | 2020 PEA Average Operating Costs | 24-87
Table 24.30 | Open Pit Mine Operating Costs | 24-88
Table 24.31 | Underground Mine Operating Costs | 24-88
Table 24.32 | Process Operating Costs | 24-89
Table 24.33 | Water Management Costs | 24-90
Table 24.34 | PST Rates | 24-91
Table 24.35 | Metal Price Assumptions in 2020 PEA | 24-93
Table 24.36 | 2020 PEA Financial Analysis Summary | 24-95
Table 24.37 | PEA Cashflow on an Annualized Basis | 24-98
Table 24.38 | Alternate Cases | 24-100
Table 25.1 | Summary of Major Pre- and Post-tax Results by Metal Price Scenario | 25-6

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List of Figures

Figure 1.1 | Panoramic View of KSM Deposits (Looking East) | 1-1
Figure 1.2 | General Location Map | 1-7
Figure 1.3 | Geology of the KSM District | 1-9
Figure 1.4 | Mitchell Zone Cross Section | 1-11
Figure 1.5 | 2016 PFS Mill Feed Production Schedule | 1-17
Figure 1.6 | 2016 PFS Open Pit LOM | 1-18
Figure 1.7 | 2016 PFS Simplified Process Flow Sheet | 1-20
Figure 1.8 | 2016 PFS Sensitivity Analysis of Pre-tax NPV at a 5% Discount Rate | 1-31
Figure 1.9 | 2016 PFS Sensitivity Analysis of Pre-tax IRR | 1-31
Figure 1.10 | 2016 PFS Sensitivity Analysis of Pre-tax Payback Period | 1-32
Figure 4.1 | KSM Mineral Claim Map | 4-3
Figure 4.2 | KSM Placer Claim Map | 4-4
Figure 7.1 | Geology of the KSM District | 7-2
Figure 7.2 | Geological Map of the Kerr Deposit | 7-5
Figure 7.3 | Vertical E-W Section through the Kerr Deposit, at 6,258,650N | 7-6
Figure 7.4 | Vertical E-W Section through the Kerr Deposit, at 6,259,650N | 7-7
Figure 7.5 | Map of the Sulphurets Deposit | 7-10
Figure 7.6 | Vertical Cross-Section through the Sulphurets Deposit, Looking ENE | 7-11
Figure 7.7 | Vertical Cross-Section through the Sulphurets Deposit, Looking NNE | 7-12
Figure 7.8 | Geology Map of the Mitchell Deposit | 7-15
Figure 7.9 | Vertical Section through the Mitchell Deposit | 7-16
Figure 7.10 | Plan View of the Mitchell Deposit at 500 masl | 7-17
Figure 7.11 | Iron Cap Geology Map | 7-20
Figure 7.12 | Vertical Cross-Section through the Iron Cap Deposit, Looking NNE | 7-21
Figure 7.13 | Plan View Through the Iron Cap Deposit at 1200 m Elevation | 7-22
Figure 10.1 | KSM Drill Hole Locations | 10-5
Figure 10.2 | Drill Hole Locations – Kerr Deposit | 10-6
Figure 10.3 | Drill Hole Locations – Sulphurets Deposit | 10-7
Figure 10.4 | Drill Hole Locations – Mitchell Deposit | 10-8
Figure 10.5 | Drill Hole Locations – Iron Cap Deposit | 10-9
Figure 11.1 | Duplicate Sample Box Plots | 11-6
Figure 11.2 | Gold Duplicate Sample Graphs | 11-7
Figure 11.3 | Copper Duplicate Sample Graphs | 11-8
Figure 13.1 | Copper and Gold Open Cycle Flotation Variability Test Results (KM2153) | 13-9
Figure 13.2 | Copper Open Cycle Flotation Performance vs Copper Head Grade at a Concentrate Grade of 25% Copper (KM2153) | 13-10
Figure 13.3 | Flotation Performance – Open Circuit Flotation Tests, Mitchell (KM2153) | 13-11
Figure 13.4 | Copper Recovery vs. Copper Feed – Open Cleaner Circuit Tests (KM2153) | 13-11
Figure 13.5 | Leaching Test Results – Mitchell 2017 and 2018 (ALS KM5455) | 13-19
Figure 13.6 | Variability Test Results – Copper – Iron Cap, 2015 (ALS) – KM4672 | 13-31
Figure 13.7 | Variability Test Results – Gold – Iron Cap, 2015 (ALS) – KM4672 | 13-31
Figure 13.8 | Copper Recovery vs. Rougher Mass Recovery and Grind Size – Iron Cap, 2015 (ALS) – KM4672 | 13-32
Figure 13.9 | Gold Recovery vs. Rougher Mass Recovery and Grind Size – Iron Cap, 2015 (ALS) – KM4672 | 13-32
Figure 13.10 | KM5063 Copper Sulphides Liberation – Mitchell and Deep Kerr 2017 (ALS) | 13-37
Figure 13.11 | KM5266 Copper Sulphides Liberation – Deep Kerr 2017 (ALS) | 13-38

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Figure 13.12 | KM5063 and KM5266 Open Circuit Flotation Flowsheet –Deep Kerr 2016/2017 (ALS) | 13-39
Figure 13.13 | KM5063 and KM5266 LCT Flotation Flowsheet – Deep Kerr 2016/2017 (ALS) | 13-40
Figure 13.14 | Effect of Slurry Solid Density and Primary Grind Size _ Rougher Flotation _ Copper – KM6004 | 13-47
Figure 13.15 | Effect of Slurry Solid Density and Primary Grind Size _ Rougher Flotation _ Gold – KM6004 | 13-48
Figure 13.16 | LCT Flotation Flowsheet – Iron Cap (2017 – 2020 ALS) | 13-48
Figure 13.17 | Gold Recovery Flowsheet | 13-51
Figure 14.1 | Gold-Copper Probability Plots – Kerr Deposit | 14-5
Figure 14.2 | Down-hole Grade Correlograms – Kerr Deposit | 14-7
Figure 14.3 | Directional Grade Correlograms – Kerr Deposit | 14-8
Figure 14.4 | Kerr Cross Section 6,259,650 N. – Gold | 14-10
Figure 14.5 | Kerr Cross Section 6,259,650 N. – Copper | 14-11
Figure 14.6 | Kerr 850 m Level Plan – Gold | 14-12
Figure 14.7 | Kerr 850 m Level Plan - Copper | 14-13
Figure 14.8 | Kerr Gold-Copper Swath Plots by Elevation | 14-14
Figure 14.9 | Gold-Copper Probability Plots – Sulphurets Deposit | 14-18
Figure 14.10 | Down-hole Grade Correlograms – Sulphurets Deposit | 14-20
Figure 14.11 | Directional Grade Correlograms – Sulphurets Deposit | 14-21
Figure 14.12 | Sulphurets Cross Section 20 - Gold | 14-23
Figure 14.13 | Sulphurets Cross Section 20 – Copper | 14-24
Figure 14.14 | Sulphurets 1135m Level Plan - Gold | 14-25
Figure 14.15 | Sulpurets 1135 m Level Plan – Copper | 14-26
Figure 14.16 | Sulphurets Gold-Copper Swath Plots by Elevation | 14-27
Figure 14.17 | Gold-Copper Cumulative Probability Plots – Mitchell Deposit | 14-31
Figure 14.18 | Down-hole Grade Correlograms – Mitchell Deposit | 14-33
Figure 14.19 | Directional Grade Correlograms – Mitchell Deposit | 14-34
Figure 14.20 | Mitchell Cross Section 11 - Gold | 14-36
Figure 14.21 | Mitchell Cross Section 11 - Copper | 14-37
Figure 14.22 | Mitchell760 m Level Plan - Gold | 14-38
Figure 14.23 | Mitchell 760 m Level Plan - Copper | 14-39
Figure 14.24 | Mitchell Swath Plots by Elevation | 14-40
Figure 14.25 | Gold-Copper Cumulative Probability Plots – Iron Cap Deposit | 14-44
Figure 14.26 | Down-hole Grade Correlograms – Iron Cap Deposit | 14-46
Figure 14.27 | Directional Grade Correlograms – Iron Cap Deposit | 14-47
Figure 14.28 | Iron Cap Cross Section 12 – Gold | 14-49
Figure 14.29 | Iron Cap Cross Section 12 – Copper | 14-50
Figure 14.30 | Iron Cap 1200 m Level Plan – Gold | 14-51
Figure 14.31 | Iron Cap 1200 m Level Plan – Copper | 14-52
Figure 14.32 | Iron Cap Gold-Copper Swath Plots by Elevation | 14-53
Figure 16.1 | Mitchell Sensitivity of Ore Tonnes to Pit Size | 16-10
Figure 16.2 | Sulphurets Sensitivity of Ore Tonnes to Pit Size | 16-11
Figure 16.3 | Kerr Sensitivity of Ore Tonnes to Pit Slope and Pit Size | 16-11
Figure 16.4 | Plan View of the KSM LG Pit Limits | 16-12
Figure 16.5 | Mitchell Open/Underground Pit and Economic Pit Limit – North-South Section at East 422950, Viewed from the East | 16-13
Figure 16.6 | Sulphurets Economic Pit Limit – North-South Section at East 421725, Viewed from the East | 16-13
Figure 16.7 | Kerr Economic Pit Limit – East-West Section at North 6258800, Viewed from the South | 16-14

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Figure 16.8 | Plan View of Mitchell Pit Phases | 16-16
Figure 16.9 | Plan View of Sulphurets Pit Phases | 16-17
Figure 16.10 | Plan View of Kerr Pit Phases | 16-18
Figure 16.11 | End of Pre-production (Year -1) | 16-24
Figure 16.12 | End of Year 5 | 16-25
Figure 16.13 | Open Pit Life of Mine | 16-26
Figure 16.14 | Plan View of the Mitchell and Iron Cap Block Cave Mines | 16-35
Figure 16.15 | Section View of the Mitchell and Iron Cap Block Cave Mines (Looking North) | 16-36
Figure 16.16 | Section View of the Mitchell Block Cave Mine Design (Looking South) | 16-39
Figure 16.17 | Mitchell Block Cave Mine Development and Production Schedules | 16-41
Figure 16.18 | Section Looking East of the Iron Cap block Cave Mine Design | 16-43
Figure 16.19 | Iron Cap Block Cave Mine Development and Production Schedules | 16-45
Figure 16.20 | KSM Mill Feed Production Schedule | 16-47
Figure 17.1 | Simplified Process Flowsheet | 17-2
Figure 17.2 | Treaty Process Plant Layout | 17-6
Figure 18.1 | Mine Site Layout after Initial Construction | 18-2
Figure 18.2 | Ultimate TMF Layout | 18-3
Figure 18.3 | Mine Site Mapped Geology | 18-6
Figure 18.4 | Mine Site Ultimate Water Management Facilities | 18-13
Figure 18.5 | Mine Site Monthly Water Treatment Rate | 18-15
Figure 18.6 | Water Storage Dam Sections | 18-17
Figure 18.7 | TMF Staging Plan | 18-22
Figure 18.8 | North Tailing Dam | 18-25
Figure 18.9 | Saddle and Splitter Tailing Dams | 18-26
Figure 18.10 | Southeast Tailing Dam | 18-27
Figure 18.11 | Schematic TMF Water Cycle | 18-29
Figure 18.12 | Ultimate TMF with Catchments and Diversion Channels | 18-31
Figure 18.13 | MTT Dual Track Plan View (Distances in Metres) | 18-37
Figure 18.14 | MTT Train Transport Drift Section (Dimensions in Millimeters) | 18-38
Figure 18.15 | Treaty Personnel, Freight, and Fuel Staging and Marshalling | 18-40
Figure 18.16 | Mitchell Personnel, Freight, and Fuel Staging and Marshalling | 18-41
Figure 18.17 | Mine Site Roads | 18-42
Figure 18.18 | NTL Route Map | 18-47
Figure 18.19 | Proposed Access Roads Network | 18-58
Figure 18.20 | Proposed Winter Glacier Access Route | 18-62
Figure 18.21 | Construction Schedule Summary (Level 1) | 18-71
Figure 18.22 | EPCM Organizational Chart | 18-73
Figure 20.1 | KSM Mine Site Water Management Schematic | 20-12
Figure 21.1 | Operating Cost Distribution | 21-11
Figure 21.2 | LOM Average Unit Operating Cost for Open Pit Mining (US$/t Material Mined) | 21-13
Figure 22.1 | Pre-tax Undiscounted Annual and Cumulative Cash Flow | 22-4
Figure 22.2 | Sensitivity Analysis of Pre-tax NPV at a 5% Discount Rate | 22-6
Figure 22.3 | Sensitivity Analysis of Pre-tax IRR | 22-6
Figure 22.4 | Sensitivity Analysis of Pre-tax Payback Period | 22-7
Figure 24.1 | Mine Site General Arrangement (PEA) | 24-2
Figure 24.2 | Sensitivity of PEA Mill Feed Tonnes to Pit Case | 24-5
Figure 24.3 | Sensitivity of Unit Net Cash Flow (NCF) to Pit Case | 24-7
Figure 24.4 | Mitchell Slope Sensitivity Section View Looking West | 24-8
Figure 24.5 | PEA Mitchell Pit Design – Plan View | 24-9
Figure 24.6 | Sulphurets Pit – Plan View | 24-10
Figure 24.7 | Kerr Pit – Plan View | 24-10

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Figure 24.8 | Rock Storage Facility Design | 24-12
Figure 24.9 | Mitchell Pit, End of Pre-Production | 24-13
Figure 24.10 | Open Pit Mill Feed Grades | 24-15
Figure 24.11 | Open Pit Production Schedule | 24-16
Figure 24.12 | Iron Cap Footprint for Lift at 870 m Elevation | 24-26
Figure 24.13 | Plan View of Iron Cap Mine with Contour | 24-28
Figure 24.14 | Typical Level Arrangement for Iron Cap | 24-29
Figure 24.15 | Drawpoint Configuration for Iron Cap | 24-30
Figure 24.16 | Ventilation Schematic for Iron Cap Mine (Section Looking West) | 24-32
Figure 24.17 | Iron Cap Mine Production Plan | 24-33
Figure 24.18 | Proposed Drawpoint Development Sequence for the Iron Cap Mine | 24-34
Figure 24.19 | Plan View of Deep Kerr 625 m Lift 1 Footprint. | 24-38
Figure 24.20 | Plan View of Deep Kerr 130 m Lift 2 Footprint. | 24-39
Figure 24.21 | Deep Kerr Mine Plan View | 24-41
Figure 24.22 | Deep Kerr Lift 1 (625L Mine Area) General Arrangement | 24-42
Figure 24.23 | Deep Kerr Lift 2 (130L Mine Area) General Arrangement | 24-43
Figure 24.24 | Typical Level Arrangement for Deep Kerr | 24-43
Figure 24.25 | Drawpoint Configuration for Deep Kerr | 24-45
Figure 24.26 | Ventilation Schematic | 24-46
Figure 24.27 | Mine Dewatering General Arrangement | 24-47
Figure 24.28 | Deep Kerr Mine Production Plan | 24-48
Figure 24.29 | KSM Mill Feed Production Schedule | 24-53
Figure 24.30 | Overall Process Block Flow Diagram | 24-61
Figure 24.31 | Main Process Plant Layout | 24-62
Figure 24.32 | Leaching Plant Layout | 24-63
Figure 24.33 | Water Treatment Trains Required and Constructed | 24-70
Figure 24.34 | Post-Tax Undiscounted Cash Flow Sensitivity Analysis | 24-96
Figure 24.35 | Post-Tax NPV 5% Sensitivity Analysis | 24-96
Figure 24.36 | Post-Tax IRR Sensitivity Analysis | 24-97

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Glossary

Units of Measure

ampere | A
annum (year) | a
bank cubic metres | bcm
billion tonnes | Bt
billion | B
centimetre | cm
Coefficients of Variation | CVs
cubic centimetre | cm3
cubic kilometre | km3
cubic metre | m3
day | d
days per week | d/wk
days per annum (year) | d/a
dead weight tonnes | DWT
degree | °
degrees Celsius | °C
diameter | ø
dollar (American) | US$
dollar (Canadian) | Cdn$
dry metric tonne | dmt
foot | ft
gallon | gal
gallons per minute (US) | gpm
gigawatt hours | Gwh
gigawatt | GW
gram | g
grams per cubic centimetre | g/cc
grams per litre | g/L
grams per tonne | g/t
gravitational constant | g
greater than | >
hectare | ha
hertz | Hz
horsepower | hp
hour | h
hours per day | h/d

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hours per week | h/wk
hours per annum (year) | h/a
inch | in
kilo (thousand) | k
kilogram | kg
kilograms per cubic metre | kg/m3
kilograms per day | kg/d
kilograms per hour | kg/h
kilograms per square metre | kg/m2
kilometre | km
kilometres per hour | km/h
kilonewton | kN
kilopascal | kPa
kilotonne | kt
kilovolt | kV
kilowatt hour | kWh
kilowatt hours per tonne | kWh/t
kilowatt hours per annum (year) | kWh/a
kilowatt | kW
less than | <
litre | L
litres per hour | L/h
litres per second | L/s
Megaannum (million years) | Ma
megavolt-ampere | MVA
megawatt | MW
metre | m
metres above sea level | masl
metres per second | m/s
metres per annum (year) | m/a
microns | µm
milligram | mg
milligrams per litre | mg/L
millilitre | mL
millimetre | mm
million | M
million tonnes | Mt
minute (plane angle) | '
minute (time) | min
month | mo
Ohm-metre | Wm
ounce | oz

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Ounce per tonne | oz/t
parts per billion | ppb
parts per million | ppm
Pascal | Pa
percent | %
pound(s) | lb
pound(s) per square inch | psi
revolutions per minute | rpm
second (plane angle) | "
second (time) | s
specific gravity | SG
square centimetre | cm2
square kilometre | km2
square metre | m2
square millimetre | mm2
three-dimensional | 3D
tonne (1,000 kg) (metric ton) | t
tonnes per day | t/d
tonnes per hour | t/h
tonnes per annum (year) | t/a
troy ounce | troy oz
volt | V
week | wk
weight/weight | w/w
wet metric ton | wmt
Abbreviations and Acronyms |
AACE® International | AACE®
abrasion index | Ai
acid rock drainage | ARD
Acid-base accounting | ABA
acidification, volatilization of hydrogen cyanide gas, and re-neutralization | AVR
Acoustic Televiewer | ATV
Application Information Requirements | AIR
ALS Canada Ltd. | ALS
alternating current | AC
aluminum oxide | Al2O3
American Institute of Professional Geologists | AIPG
American National Standards Institute | ANSI
ammonium nitrate-fuel oil | ANFO
antimony | Sb

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Aquatic Effects Monitoring Program | AEMP
arsenic | As
atomic absorption spectrometry | AAS
atomic absorption | AA
atomic emission spectroscopy | AES
audio magneto telluric | AMT
ball mill work index | BMWi
Barrick Gold Corp. | Barrick
Basal Shear Fault | BSF
Basal Shear Zone | BSZ
BC Environmental Assessment Act | BCEAA
BC Mineral Inventory | Minfile
best applicable practices | BAP
best available technology | BAT
best available tailings technology | BATT
BGC Engineering Inc. | BGC
Biogeoclimatic Ecosystem Classification | BEC
BioteQ Environmental Technologies Inc. | BioteQ
bismuth | Bi
British Columbia | BC
British Columbia Environmental Assessment Office | BCEAO
British Columbia Utilities Commission | BCUC
Bulk Mineral Analysis with Liberation | BMAL
bulk mineral analysis | BMA
cadmium | Cd
calcium oxide | CaO
California Air Resources Board | CARB
Canadian Dam Association | CDA
Canadian Development Expense | CDE
Canadian Environmental Assessment | CEA
Canadian Environmental Assessment Act | CEAA
Canadian Exploration Expense | CEE
Canadian Institute of Mining, Metallurgy and Petroleum | CIM
Canadian National Railroad | CNR
Capital Cost Allowance | CCA
carbon-in-leach | CIL
carboxymethyl cellulose | CMC
Caro’s acid | H2SO5
CDN Resource Laboratories Ltd. | CDN Resource
chalcopyrite | Cp
chlorine | Cl
closed-circuit television | CCTV

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coarse ore stockpile | COS
cobalt | Co
common voltage reference | CVR
comparative work index | CWi
Construction Diversion Tunnel | CDT
conventional counter-current decantation | CCD
copper | Cu
copper sulphate | CuSO4
cost breakdown structure | CBS
Cost, Insurance and Freight – Free Out | CIF-FO
Coulter Creek Access Road | CCAR
counter current decantation | CCD
cross-linked polyethylene | XLPE
Cumulative Tax Credit Account | CTCA
Deep Kerr | DK
Delegation of Authority Guideline | DOAG
Demand Side Management | DSM
direct current | DC
direct cyanide leaching | DCN
discounted cash flow | DCF
Distributed Control System | DCS
east | E
EBC Inc. | EBC
economic, social, and cultural impact assessment | ESCIA
effective grinding length | EGL
Electricity Supply Agreement | ESA
electromagnetic | EM
emergency medical technician | EMT
engineering, procurement, construction management | EPCM
environmental assessment | EA
Environmental Design Flood | EDF
Environmental Effects Monitoring | EEM
environmental impact statement | EIS
Environmental Management System | EMS
ERM Consultants Canada Inc. | ERM
Esso Minerals Canada Ltd | Esso Minerals
Factor of Safety | FOS
Feasibility Study | FS
Fisheries and Oceans Canada | DFO
fluorine | F
Footprint Finder | FF
Free Carrier | FCA

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front-end loader | FEL
G&T Metallurgical Services Ltd. | G&T
galena | Gn
gas insulated | GIS
general and administrative | G&A
general mine expense | GME
Global Climatic Models | GCMs
global positioning system | GPS
gold | Au
gold equivalent | AuEQ
Golder Associates Ltd. | Golder
Goods and Services Tax | GST
Granduc Mines Ltd. | Granduc
Granmac Services Ltd. | Granmac
greenhouse gas | GHG
gross vehicle weight | GVW
Ground Penetrating Radar | GPR
harmful alteration, disruption or destruction | HADD
Harmonized Sales Tax | HST
Hazelton Volcanics | HV
Hazen Research Inc. | Hazen
health, safety and security | HS&S
heating, ventilation, and air conditioning | HVAC
height of draw | HOD
hematite | He
high-density polyethylene | HDPE
high-density sludge | HDS
high-pressure grinding roll | HPGR
hydrochloric acid | HCl
hydrogen peroxide | H2O2
Impact Benefits Agreement | IBA
Independent Geotechnical Review Board | IGRB
Independent Power Producer | IPP
induced polarization | IP
inductively coupled plasma | ICP
Inflow Design Flood | IDF
Input Tax Credit | ITC
internal rate of return | IRR
International Electrotechnical Commission | IEC
International Organization for Standardization | ISO
inter-ramp angle | IRA
inter-ramp height | IRH

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inverse distance weighting | IDW
Iron Cap Fault | ICF
iron | Fe
Joint Health and Safety Committee | JHSC
joint venture | JV
Kerr-Sulphurets-Mitchell | KSM
KSM Mining ULC | KSM Mining
Klohn Crippen Berger Ltd. | KCB
Köeppern Machinery Australia Pty Ltd. | Köeppern
Lambert Civil Consulting Ltd. | LCC
Land Resource Management Plan | LRMP
lead | Pb
Lerchs-Grossmann | LG
life-of-mine | LOM
Light Detection and Ranging | LIDAR
Lilburn & Associates LLC | Lilburn
linear low-density polyethylene | LLDPE
liquefied natural gas | LNG
liquefied petroleum gas | LPG
load factor | LF
load-haul-dump | LHD
local study area | LSA
locked cycle tests | LCT
magnesium oxide | MgO
magnetite | Ma
magneto telluric | MT
maintenance and repair contracts | MARC
manganese oxide | MnO
material take-off | MTO
McElhanney Consulting Services Inc. | McElhanney
McTagg Diversion Tunnels | MTDT
mercury | Hg
metabisulphite | MBS
Metal and Diamond Mining Effluent Regulations | MDMER
metal leaching | ML
Metal Mining Effluent Regulations | MMER
Methyl isobutyl carbinol | MIBC
Metso Minerals Industries Inc. | Metso
mineral titles online. | MTO
Mining Rock Mass Rating | MRMR
Ministry of Energy and Mines | MEM
Ministry of Energy, Mines and Natural Gas | MEMNG
Ministry of Energy, Mines and Petroleum Resources | MEMPR

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Ministry of Environment | MOE
Ministry of Forests | MOF
Ministry of Forests, Lands and Natural Resource Operations | MFLNRO
Ministry of Forests, Lands and Natural Resource Operations Road Division | MFLNRORD
Ministry of Transportation and Infrastructure | MOTI
Mitchell Diversion Tunnel | MDT
Mitchell Thrust Fault | MTF
Mitchell Valley Drainage Tunnel | MVDT
Mitchell-Treaty Twinned Tunnels | MTT
molybdenum | Mo
Moose Mountain Technical Services | MMTS
motor control centre | MCC
Multiple Accounts Analysis | MAA
Multiple Pulse in Air | MPiA
Municipal Wastewater Regulation | MWR
Nass Timber Supply Area | Nass TSA
Nass Wildlife Area | NWA
National Ambient Air Quality Objectives | NAAQOs
National Instrument 43-101 | NI 43-101
nearest neighbor | NN
Neil S. Seldon & Associates Ltd. | NSA
net cash flow | NCF
net present value | NPV
net smelter price | NSP
net smelter return | NSR
Newhawk Gold Mines Ltd. | Newhawk
Newmont Exploration of Canada Ltd. | Newmont
nickel | Ni
Nisga’a Final Agreement | NFA
Nisga’a Lisims Government | NLG
no net loss | NNL
non-potentially acid generating | NPAG
Nordic Minesteel Technologies Inc. | NMT
North American Datum | NAD
North Pit Wall Dewatering Adit | NPWDA
North Treaty Access Road | NTAR
north | N
Northwest Fault | NWF
Northwest Transmission Line | NTL
Operator Interface Stations | OIS
Ore Control System | OCS
Ore Preparation Complex | OPC

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peak ground acceleration | PGA
personal protective equipment | PPE
Phelps Dodge Corp. | Phelps Dodge
phosphorus | P
pit slope angle | PSA
Placer Dome | Placer
Pocock Industrial Inc. | Pocock
point of delivery | POD
potassium amyl xanthate | PAX
potentially acid generating | PAG
Prefeasibility Study | PFS
Preliminary Economic Assessment | PEA
Pretium Resources Inc. | Pretium
PricewaterhouseCoopers | PwC
prilled ammonium | AN Prill
probable maximum flood | PMF
Probable Maximum Precipitation | PMP
Process Control System | PCS
Process Tailing and Management Area | PTMA
programmable logic controller | PLC
Provincial Sales Tax | PST
pyrite | Py
Qualified Person | QP
quality assurance | QA
quality control | QC
quantile-quantile | QQ
Quantitative Evaluation of Minerals by Scanning | QEMSCAN®
quartz vein fragments | QABX
Raewyn Copper | RC
Raewyn fault | RF
Rate Design Application | RDA
real-time kinematic | RTK
regional study area | RSA
regulation | Reg.
request for information | RFI
Resource Management Zone | RMZ
Resource Modelling Inc. | RMI
reverse osmosis | RO
Revised Statutes of British Columbia | RSBC
rhenium | Re
rock quality designation | RQD
Rock Storage Facility | RSF
rotations per minute | RPM

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run-of-mine | ROM
Seabridge Gold Inc. | Seabridge
selenium | Se
self-contained breathing apparatus | SCBA
semi-autogenous grinding mill comminution | SMC
semi-autogenous grinding | SAG
SGS Minerals Services | SGS
silica | SiO2
silver | Ag
site water management | SWM
S.J.V. Consultants Ltd. | SJV
Skeena fold and thrust belt | SFTB
Snowfields Slide Dewatering Adit | SSDA
Society for Mining, Metallurgy, and Exploration | SME
sodium cyanide | NaCN
sodium hydrosulphide | NaHS
sodium hydroxide | NaOH
sodium silicate | Na2SiO3
sodium sulphide | Na2S
solids liquid separation | SLS
south | S
Special Use Permit | SUP
Species at Risk Act | SARA
Standard Penetration Test | SPT
standard reference material | SRM
sulphide | S-2
sulphidization, acidification, recycling, and thickening of precipitate | SART
sulphur | S
Sulphurets Fault Zone | SFZ
Sulphurets Thrust Fault | STF
Sulphurets-Mitchell Conveyor Tunnel | SMCT
sulphuric acid | H2SO4
Surface Science Western | SSW
Sustainable Resource Management Plan | SRMP
System Impact Study | SIS
Tailing Management Facility | TMF
Tariff Supplement | TS
temporary water treatment plants | TWTP
tennantite | Tn
Tetra Tech, Inc. | Tetra Tech
tetrahedrite | Tt
The Claim Group Inc. | TCG
the Environmental Assessment Application/Environmental Impact Statement | the Application/EIS

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time domain electromagnetic | TDEM
total sulphur | ST
total suspended sediments | TSS
treatment charge/refining charge | TC/RC
Treaty Creek Access Road | TCAR
tunnel support classes | TSC
ultra-high frequency | UHF
undercut | UC
Uniform Hazard Response Spectra | UHRS
Universal Transverse Mercator | UTM
University of British Columbia | UBC
Unlimited Liability Corporation | ULC
valued component | VC
Voice over Internet Protocol | VoIP
Water Storage Dam | WSD
Water Storage Facility | WSF
Water Treatment Plant | WTP
weak acid dissociable | WAD
west | W
WN Brazier Associates Inc. | Brazier
Wood Canada Ltd. | Wood
work breakdown structure | WBS
work index | Wi
x-ray fluorescence | XRF
Z-Axis Tipper Electromagnetic | ZTEM
zinc | Zn

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| 1.0 | Summary

| 1.1 | Introduction

Seabridge Gold Inc.’s (Seabridge) Kerr-Sulphurets-Mitchell (KSM) Property (the Property) involves the development of major gold-copper deposits located in northwest British Columbia (BC) off Highway 37, approximately 65 km by air north-northwest of the ice free Port of Stewart, BC. The Property is situated within the coastal mountains of BC, approximately 30 km topographically upgradient of the Alaska-BC border. KSM is one of the few undeveloped projects in the world that has received its environmental approvals, these having been granted by both the Government of Canada and the Government of BC. KSM includes four major mineralized zones, identified as the Mitchell, Kerr, Sulphurets, and Iron Cap deposits. The deposits contain significant gold, copper, silver, and molybdenum mineralization. Figure 1.1 is a panoramic view looking east towards the aforementioned deposits.

Figure 1.1 Panoramic View of KSM Deposits (Looking East)

In conjunction with the environmental approvals, Seabridge also received early-stage construction permits for KSM from the Province of BC in September 2014. The permits issued include:

| ● | authority to construct and use roadways along Coulter Creek and Treaty Creek

| ● | rights-of-way for the proposed Mitchell-Treaty Twinned Tunnel (MTT) alignment connecting KSM facilities

| ● | permits for constructing and operating various camps required to support construction activities

| ● | permits authorizing early-stage construction activities at the mine site and Tailings Management Facility (TMF).

Seabridge also received permits from the BC Government in October 2016, which allows the construction of an exploration adit to explore mineralization associated with the Deep Kerr deposit.

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In this report, this “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report” (the Report), the KSM mine development has been evaluated with two alternative studies that each evaluate different options for mine development. The 2016 Prefeasibility Study (2016 PFS) evaluated mining mostly by open pit method at a specified processing rate, while the 2020 Preliminary Economic Assessment (2020 PEA) is an alternative study that leverages the technical information of the previous studies and evaluates mine development dominated by underground mining methods at a processing rate significantly higher than that used in the 2016 PFS.

The results of the economic analyses for the 2016 PFS and 2020 PEA represent forward-looking information that is subject to a number of known and unknown risks, uncertainties and other factors that may cause actual results to differ materially from those presented in this Report. Forward-looking statements in this report include, but are not limited to, statements with respect to future metal prices, the estimation of Mineral Resources and Mineral Reserves, the estimated mine production and metals recovered, the estimated capital and operating costs, and the estimated cash flows generated from the planned mine production for the different development options. The material factors or assumptions used to develop the forward-looking information are identified in the relevant sections of this Report.

| 1.2 | Key Study Outcomes

The 2016 PFS and the 2020 PEA are stand alone studies. The results of the 2016 PFS remain valid and represent a viable option for developing KSM, with the 2020 PEA assessing an alternative development option at a scoping level. The 2020 PEA does not impact the 2016 PFS. The 2020 PEA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the results of the 2020 PEA will be realized. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

The results of the 2020 PEA were disclosed in Seabridge’s press release entitled “Updated PEA Study Enhances Seabridge Gold’s KSM Project”, dated April 27th, 2020. The Report will be filed in support of the disclosure of the 2020 PEA results.

| 1.2.1 | 2016 Preliminary Feasibility Study

This 2016 PFS was prepared under the direction of Tetra Tech, Inc. (Tetra Tech), for Seabridge in 2016. The Qualified Persons (QPs) of this Report have verified that the 2016 PFS remains current and is suitable to be included unchanged in this Report. The 2016 PFS consists of work produced by Tetra Tech and the following independent consultants:

| ● | Moose Mountain Technical Services (MMTS)

| ● | Golder Associates Ltd. (Golder)

| ● | McElhanney Consulting Services Ltd. (McElhanney)

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| ● | BGC Engineering Inc. (BGC)

| ● | Resource Modeling Inc. (RMI)

| ● | Klohn Crippen Berger Ltd. (KCB)

| ● | ERM Consultants Canada Ltd. (ERM)

| ● | WN Brazier Associates Inc. (Brazier).

The 2016 PFS envisaged a combined open pit/underground block caving mining operation that is scheduled to operate for 53 years. During the initial 33 years of mine life, the majority of ore would be derived from open pit mines, with the tail end of this period supplemented by the initial development of underground block cave mines. Ore delivery to the mill during Year 2 to Year 35 is designed to be maintained at an average of 130,000 t/d. After depletion of the open pits, the mill processing rate would be reduced to just over 95,000 t/d for 10 additional years before ramping down to just over 60,000 t/d for the remaining few years of stockpile reclaim at the end of the mine life. Over the entire 53-year mine life, ore would be fed to a flotation and gold extraction mill. The flotation plant would produce a gold/copper/silver concentrate for transport by truck to the nearby sea port for shipment to Pacific Rim smelters. Extensive metallurgical testing confirms that KSM can produce a clean concentrate with an average copper grade of 25% with a high gold and silver content, making it readily saleable. A separate molybdenum concentrate and gold-silver doré would also be produced at the KSM processing facility.

All dollar figures presented in this report are stated in US dollars, unless otherwise specified. The 2016 PFS concluded:

| ● | the estimated Proven and Probable Mineral Reserves as of July 31, 2016 are 38.8 Moz of gold and 10.2 Blb of copper (2.2 Bt at an average grade of 0.55 g/t gold and 0.21% copper per tonne).

| ● | projected initial capital costs are US$5.0 billion and sustaining capital costs are US$5.5 billion estimated to a +25%/-10% level of accuracy and incorporate substantial enhancements to meet environmental improvements that were committed to in the environmental assessment (EA) review process

| ● | the 2016 Base Case after-tax NPV at a 5% discount rate is US$1.5 billion with an IRR of 8% and payback period is approximately 6.8 years of a 53-year operating mine life, using three-year average price assumptions of US$1,230.00/oz gold, US$2.75/lb copper, US$17.75/oz silver, US$8.49/lb molybdenum and a foreign exchange rate of US$0.80 per Cdn$1.00

| ● | overall, the 2016 PFS confirmed that KSM is an economic project with an unusually long life in a low-risk jurisdiction.

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| 1.2.2 | 2020 Preliminary Economic Assessment

Wood Canada Limited (Wood) and the following independent consultants have prepared an updated Preliminary Economic Assessment study (the 2020 PEA):

| ● | MMTS

| ● | Golder

| ● | KCB

| ● | BGC

| ● | Brazier

| ● | ERM.

The 2020 PEA has been prepared as an alternative option to the 2016 PFS for the KSM development. The 2020 PEA was undertaken to evaluate a different approach to mine development by emphasizing low cost block cave mining early in the production plan and reducing the size of the open pits, which reduces the surface disturbances in the re-designed mine development. The 2020 PEA assesses the potential impacts of incorporating higher grade Inferred Mineral Resources delineated at Iron Cap and Deep Kerr into the mine design, and increasing the annual average mill throughput from 130,000 t/d envisioned in the 2016 PFS to 170,000 t/d in the 2020 PEA.

The key 2020 PEA outcomes for the projected economic results are presented in Table 1.1. The 2020 PEA show improvements over the 2016 PFS in unit operating costs, net cash flow, NPV, IRR and project payback by producing a higher percentage of mill feed through underground mining, and processing the mill feed at a higher plant throughput rate. Annual gold and copper production in the 2020 PEA increase over those shown in the 2016 PEA through the benefit of applying an improved variable cutoff strategy to early mining, improved mine sequencing and achieving this with reduced initial and sustaining capital.

Table 1.1 2020 PEA Projected Economic Results

Financial Results | 2020 Comparative Case | 2020 Base Case 3yr Avg. | 2020 Alternate Upside Case
Metal Prices: |  |  |
Gold (US$/ounce) | 1,230 | 1,340 | 1,700
Copper (US$/pound) | 2.75 | 2.80 | 2.50
Silver (US$/ounce) | 17.75 | 16.25 | 16.00
Exchange Rate: (US$/Cdn$) | 0.80 | 0.76 | 0.72
Cost Summary: |  |  |
Op cost (US$)/ Au oz recovered byproduct | (410) | (472) | (327)
Total cost (US$)/ Au oz recovered byproduct | 92 | 4 | 124
Initial Capital (US$ Billion) | 5.4 | 5.2 | 4.9
Sustaining Capital (US$ Billion) | 7.2 | 6.9 | 6.5
Unit Operating Cost On-site (US$/t milled) | 10.50 | 9.98 | 9.45
|  | table continues…

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Financial Results | 2020 Comparative Case | 2020 Base Case 3yr Avg. | 2020 Alternate Upside Case
Pre-Tax Results: |  |  |
Net Cash Flow (US$ Billion) | 31.3 | 36.8 | 43.4
NPV @ 5% Discount Rate (US$ Billion) | 8.2 | 10.3 | 13.2
Internal Rate of Return (%) | 14.9% | 17.5% | 21.9%
Payback Period (Years) | 4.3 | 3.6 | 2.9
Post-Tax Results: |  |  |
Net Cash Flow (US$ Billion) | 19.6 | 23.0 | 27.2
NPV @ 5% Discount Rate (US$ Billion) | 4. 7 | 6.0 | 7.9
Internal Rate of Return (%) | 11.9% | 14.0% | 17.4%
Payback Period (Years) | 4.9 | 4.0 | 3.2

| Notes: | 1. Base Case metal and foreign exchange prices based on April 9, 2020 3-year trailing average. The Comparative Case and Alternate Upside Case show sensitivities to a combination of changes in metal prices and exchange rate.
|  | 2. Operating and total cost per ounce of gold are after copper and silver credits. Total cost per ounce includes all start-up capital, sustaining capital, and reclamation/closure costs. The post-tax results include the BC Mineral Tax and corporate provincial and federal taxes.

The material factors or assumptions used to develop the forward-looking information in the 2020 PEA, as well as the material risk factors that could cause actual results to differ materially from the forward-looking information are more fully described in Section 24 and 25 of the Report. A portion of the Mineral Resources in the 2020 PEA mine plans, production schedules, and cash flows include Inferred Mineral Resources, that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the 2020 PEA will be realized. Due to the conceptual nature of the 2020 PEA, none of the Mineral Resources in the 2020 PEA have been converted to Mineral Reserves and therefore do not have demonstrated economic viability.

| 1.3 | Property Description and Location

The Property is located in the coastal mountains of northwest BC at a latitude and longitude of approximately 56.50° north (N) and 130.30° west (W), respectively. The Property is situated approximately 950 km northwest of Vancouver, BC; 65 km by air north-northwest of Stewart, BC; and 21 km south-southeast of the former Eskay Creek Mine. The proposed pit areas lie within the headwaters of Sulphurets Creek, which is a tributary of the Unuk River, which flows into the Pacific Ocean through Alaska. The proposed TMF will be located within the tributaries of Teigen and Treaty creeks. The Teigen and Treaty creeks are tributaries of the Bell-Irving River, which is itself a major tributary of the Nass River. The Nass River also flows to the Pacific Ocean through the northwestern portion of British Columbia, entirely within Canadian jurisdiction. Figure 1.2 is a general location map of the Property.

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The Property comprises four discrete claim blocks. The claim blocks are referred to as:

| ● | the KSM claims

| ● | the Seabee claims

| ● | the Tina claims

| ● | the Treaty Creek Switching Station claims

The four KSM claim blocks include 79 mineral claims (cell and legacy) and 2 mining leases with a combined area of 40,784.97 ha. There are also 17 KSM placer claims held by KSM Mining ULC covering part of the KSM claims. The placer claims secure rights in a historically designated placer district. The Claim Group Inc. (TCG) acts as agent on behalf of Seabridge with respect to maintaining all pertinent records associated with the Property tenures. All claims and leases are in good standing under the Mining Tenure Act of BC and are recorded as owned 100% by KSM Mining ULC, a wholly owned subsidiary of Seabridge.

Annual holding costs for all leases and claims vary by year depending on whether the fees are paid in cash or the value of work completed on developing the claims is used in lieu of a cash payment. Over the next five years, the annual cash holding costs to keep the claims and leases valid range between Cdn$450,000 to Cdn$970,000. Those estimated costs can be reduced significantly if work expenditures are applied in lieu of cash fees. No additional permits are required to address the recommendations in this report; part of the expenditures for that work can be applied in lieu of cash fees. Seabridge believes they have addressed all issues to secure access, mineral title, and ability to perform work on the property and are not aware of any risks, other than those identified in this Report, that could materially affect proposed work plans.

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Figure 1.2 General Location Map

Source: ERM (2012)

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| 1.4 | Accessibility, Climate, Local Resources, Physiography and Infrastructure

The Property lies within the rugged coastal mountains of northwestern BC, with elevations ranging from 520 m in Sulphurets Creek Valley, to over 2,300 m at the highest peaks.

There are multiple deep-water loading facilities for shipping bulk mineral concentrates located in the ice free Port of Stewart, BC. Those port facilities are currently used by the Red Chris Mine. The nearest railway is the Canadian National Railroad (CNR) Yellowhead route, which is located approximately 220 km southeast of the Property. This line runs east-west, and can deliver concentrate to deep water ports near Prince Rupert and Vancouver, BC.

The Property and its access routes are on Crown land; therefore, surface and access rights are granted under, and subject to compliance with, the Mineral Tenure Act or the Land Act or, at the discretion of the Crown, under the Mining Right of Way Act.

The closest power transmission lines, the Northwest Transmission Line (NTL), run along the Highway 37 corridor up to the Red Chris Mine. The Red Chris Mine is approximately 120 km north of KSM, whereas the NTL is less than 15 km east of KSM or approximately 30 km away from the Treaty OPC by way of the proposed Treaty Creek Access Road (TCAR).

| 1.5 | Geological Setting And Mineralization

The Property lies within “Stikinia”, a long-lived volcanic island-arc terrane that was accreted onto the Paleozoic basement of the North American continental margin in the Middle Jurassic.

Early Jurassic sub-volcanic intrusive complexes are common in the Stikinia terrane, and several host well-known precious- and base-metal-rich hydrothermal systems. These include copper-gold porphyry zones such as Galore Creek, Red Chris, Kemess, Mt. Milligan, and KSM. The geology of the KSM District is shown in Figure 1.3.

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Figure 1.3 Geology of the KSM District

Source: Seabridge (2019)

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The Kerr deposit is centered on an Early Jurassic, north-south trending, steep westerly dipping tabular intrusive complex. Drilling demonstrates that the recognized mineralized system has an overall strike extent of 2,400 m, a width of approximately 800 m, and a vertical extent of at least 2,200 m. Deep drilling since 2012 has identified two sub-parallel, north-south trending, steep west-dipping mineralized zones approximately 1,700 m long that appear to coalesce near the topographic surface. The west limb is up to 500 m thick, while the east limb is up to 300 m thick. After significant deep drilling was completed at the Kerr deposit, an updated geological interpretation and subsequent updated Mineral Resource model were completed. That new model forms the basis for the 2019 Mineral Resources and Mineral Reserves. Approximately 223 diamond core holes totaling about 85,000 m of drilling data were used to construct the Kerr block model used for this Report.

Sulphurets is a structurally complex deposit intersected by numerous east-vergent faults associated with the mid-Cretaceous Skeena fold and thrust belt. The deposit is composed of stacked thrust fault panels made up of Upper Triassic Stuhini Group and Lower Jurassic Hazelton Group volcano-sedimentary strata that are intruded by a number of dykes and stocks. The majority of mineralization occurs in the “Lower Panel”, that was historically divided into several discrete mineralized zones, including, from southwest to northeast: the Canyon zone, Breccia Gold zone, and Raewyn Copper-Gold zone. The main body of the Sulphurets deposit has a lensoidal geometry, dipping approximately 30 degrees to the northwest with a horizontal extent of 2,200m, down dip extent of 550 m, and a true thickness of up to 330 m. Approximately 139 core holes totaling about 45,000 m were used to construct the Sulphurets block model used for this Report.

The Mitchell Zone (Figure 1.4) is underlain by foliated, schistose, intrusive, volcanic, and clastic rocks that are exposed in an erosional window below the shallow north dipping Mitchell Thrust Fault (MTF). Mineralization at the Mitchell deposit is genetically and spatially related to the Early Jurassic Mitchell intrusive complex, which is composed of Sulphurets (Texas Creek) suite diorite, monzodiorite, and granodiorite stocks and dykes. The intrusive complex cuts sedimentary and volcanic rocks of the Upper Triassic Stuhini Group and sandstones, conglomerates, and andesitic rocks of the Lower Jurassic Jack Formation (basal Hazelton Group). The Mitchell complex has been subdivided into three major intrusive phases. The successive intrusive phases were accompanied by the development of different hydrothermal assemblages, veining and mineralization. The Mitchell deposit features many characteristics typical of gold-enriched calc-alkaline porphyry copper deposits. Metals, chiefly gold and copper are generally at low concentrations, finely disseminated, stockwork or sheeted veinlet controlled and pervasively dispersed over hundreds of metres. Grades diminish slowly over large distances, distinct from the Sulphurets and Kerr zones. The basis for Mitchell Mineral Resources is the block model that was used in the 2016 PFS (Tetra Tech, 2016). That model was constructed using drilling data collected through 2011. Since that model was completed, 23 holes were drilled within the Mitchell resource area, although only 8 holes intersected estimated Mineral Resources based on the end-of-year 2011 block model. Grade comparisons between the 8 new holes and the end-of-year 2011 Mitchell block model showed no material differences between the new holes and the estimated block grades.

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Figure 1.4 Mitchell Zone Cross Section

Notes: The vertical section through the Mitchell deposit is shown as A-A’ (Seabridge, 2019). The traces of the Mitchell thrust fault (MTF) and Mitchell basal shear zone (BSZ) are shown on each panel. a) Simplified geology of the Mitchell deposit, showing undifferentiated wallrock (gray), the Phase 1 diorite porphyry (pink), the sheeted vein body (red), Phase 2 granodiorite porphyry dikes (darker pink), and post-mineral dolerite dikes (dark green). The traces of drill holes occurring within ± 50 m of the section are shown as black lines. b) Simplified diagram of hydrothermal alteration zoning within the Mitchell deposit, illustrating the extent of potassic alteration, moderate and strong sericitic alteration (see text), as well as the occurrence of hydrothermal anhydrite; c) contours of logged volume percent of total quartz veins within Mitchell drill core; d) Au grades from the Mitchell block model, showing clear concentric zoning; e) block model Cu grades, showing a concentric zoning pattern similar to that of Au; and f) block model Mo grades, showing a zoning pattern with highest Mo grades rimming the central zone of high Au and Cu.

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The Iron Cap deposit is the northernmost porphyry gold-copper-molybdenum deposit in the KSM district, and occurs structurally above the Mitchell deposit, in a panel of rocks located between the Mitchell Thrust Fault and Sulphurets Trust Fault. The deposit is hosted by an Early Jurassic intrusive complex that is roughly contemporaneous with intrusions located at the Mitchell and Kerr zones. The intrusive complex is composed of multiple intrusion and breccia phases, the earliest of which is a pre-mineral diorite. A second dioritic phase located in the northwestern portion of the deposit is thought to be syn-mineral and is spatially associated with some of the highest gold and copper grades observed at Iron Cap. Mineralized hydrothermal breccias are significantly more abundant and voluminous at Iron Cap than at Mitchell, Sulphurets, and Kerr. The Iron Cap mineralized zone forms a tabular body striking roughly north-south, dipping approximately -60 degrees to the west. The deposit has dimensions of approximately 1,500 m along strike, 1,500 m down dip, and up to 800 m in thickness. Mineralization remains open down dip. The Iron Cap grade model was updated following the completion of the 2018 drilling campaign. A total of 99 diamond core holes totaling about 62,000 m were used in the update.

| 1.6 | History

The modern exploration history of the property began in the 1960s, with brief programs conducted by Newmont Mining Corp. (Newmont), Granduc, Phelps Dodge, and the Meridian Syndicate. All of these programs were focused towards gold exploration. Various explorers were attracted to this area due to the numerous large, prominent pyritic gossans that are exposed in alpine areas. There is evidence that prospectors were active in the area prior to 1935. Several short hole, reconnaissance level drilling programs were undertaken between 1969 and 1991. The Sulphurets Zone was first drilled by Esso Minerals in 1969, Kerr was first drilled by Brinco Ltd. in 1985, Mitchell Creek by Newhawk in 1991, and Iron Cap by Esso Minerals in 1980.

In 1989, Placer Dome acquired a 100% interest in the Kerr deposit from Western Canadian Mines; in the following year, they acquired the adjacent Sulphurets Property from Newhawk. The Sulphurets Property also hosts the Mitchell Creek deposit and other mineral occurrences. In 2000, Seabridge acquired a 100% interest from Placer Dome in both the Kerr and Sulphurets properties, subject to capped royalties.

There is no recorded mineral production, nor evidence of it, from the Property. Immediately west of the Property, small-scale placer gold mining has occurred in the Sulphurets and Mitchell creeks.

There is no recorded mineral production, nor evidence of it, from the Property. Immediately west of the Property, small-scale placer gold mining has occurred downstream in Sulphurets Creek.

| 1.7 | Mineral Resources

The 2019 Mineral Resource estimate is the current and only Mineral Resource estimate used in this Report.

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Mineral Resources were estimated for the Property by Resource Modeling Inc. (RMI). The four mineralized zones, Kerr, Sulphurets, Mitchell and Iron Cap, were modeled within a single block model using 25m x 25m x 15m blocks. As more understanding was gained after each annual drilling campaign, individual block models were created for each area. Grade interpolation parameters have also evolved over time, reflecting changes required for modeling deeper mineralization intersected below the Kerr and Iron Cap deposits. A variety of basic descriptive statistics and spatial analyses were completed for each area upon the completion of annual drilling campaigns. These investigations include the generation of grade distribution tables, grade histograms, cumulative probability plots, grade box plots, grade contact plots, down-hole variograms, and directional variograms. In addition, new drill hole results were typically compared against the previous grade model to assess model performance.

RMI established the Mineral Resources for the various KSM mineralized zones using constraining conceptual open pit and block cave shapes which were used to establish reasonable prospects for eventual economic extraction as outlined in the CIM Definition Standards for Mineral Resources and Mineral Reserves (CIM, 2014). The following gold, copper, silver, and molybdenum metal prices were used for determining block NSR values, US$1,300/oz, US $3.00/lb, US $20.00/oz, and US $9.70/lb, respectively. Open pit and underground mining costs of Cdn$1.80/tonne and Cdn$6.00 to Cdn$7.00/tonne were used to establish conceptual open pit and underground resource shapes, along with a processing and G&A cost of Cdn$9.00/tonne.

The conceptual open pit and underground mining shapes were generated for each resource area based on calculated block model NSR values. The NSR values were generated for each deposit. MMTS generated conceptual pits for the Kerr, Sulphurets, and Mitchell deposits using MineSight® software and Lerchs-Grossmann algorithms. Golder developed conceptual block cave footprints using the block NSR values and Geovia’s PCBC™ Footprint Finder software. The footprint polygons were extruded vertically based on guidance from Golder.

The draw point extraction elevations were extruded vertically to create 3D solids that were used for resource tabulation. Conceptual caves were clipped against surface topography (Iron Cap) or conceptual resource pits (Kerr and Mitchell). Mineral Resources are determined, at Cdn$9 and Cdn$16 NSR cutoffs for open-pit constrained and underground mining constrained resources, respectively.

Table 1.2 summarizes the estimated Measured, Indicated, and Inferred Mineral Resources for each zone.

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Table 1.2 KSM Mineral Resources

Notes: Mineral Resources have an effective date of December 31, 2019. The 2019 Sulphurets drill holes were not used in the construction of the resource model but were used to validate the interpretations of the model. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. It is reasonably expected that the majority of Inferred Mineral Resources could be upgraded to Indicated Mineral Resources with continued exploration. The Mineral Resources tabulated in Table 1.2 are inclusive of those Mineral Resources that have been converted to Mineral Reserves. Numbers may not add due to precision and roundoff of tonnes and grade. Details regarding key assumptions, parameters, and methods used in estimating the Mineral Resources are included in Section 14 of this Report.

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| 1.8 | 2016 PFS Data Verification

The 2016 Prefeasibility Study has been reviewed to determine whether it remains current and is suitable to be carried forward unchanged and included in the 2020 Technical Report on the KSM Property as part of the data verification process. The QPs reviewed information on the KSM Property generated since the effective date of the 2016 PFS to confirm there has been no material change to the 2016 PFS and that the Mineral Reserves and economic outcomes from the 2016 PFS remain current.

The data verification checks used updated resource models with updated process recoveries and escalated cost assumptions and estimated that variances in total mill feed would be limited to +1.3% on mill feed tonnes, no change on gold grade, and +2.4% on copper grade. These overall differences are not material to the 2016 PFS Mineral Reserves. The financial results using the 2016 PFS mine plan with updates to capital and operating costs, 3-year average metal prices and tax code revisions are also not materially different from the 2016 PFS financial results. Specifically, IRR, NPV and payback variances would be different by less than 0.2%, US$333M and 0.2 years, respectively, on a pre-tax basis. Results are summarized in Table 1.3.

Table 1.3 Summary of the 2016 PFS Verification Checks

Item | Unit | 2016 PFS (Base Case) | 2020 Verification Check | Variance
NPV5 | US$M | 3,263 | 3,596 | 333
IRR | % | 10.4 | 10.6 | 0.2
Payback | years | 6 | 5.8 | -0.2

| 1.9 | Mining Methods

The 2016 PFS uses conventional large-scale open pit and block cave underground mining methods. Pit phases at the Mitchell, Kerr, and Sulphurets deposits have been engineered based on the results of economic pit limit analysis. Starter pits have been selected in higher-grade areas. Underground mining has been proposed for the Iron Cap deposit and below the Mitchell open pit to limit the volume of waste generated from the potential open pits.

| 1.9.1 | Mineral Reserve Estimate

In the 2016 PFS, the waste to ore open pit cut-offs and underground shut-offs, including process recovery, were determined using metal prices of US$1,200.00/oz of gold, US$2.70/lb of copper, US$17.50/oz of silver, and US$9.70/lb of molybdenum and a foreign exchange rate of US$0.83 per Cdn$1.00 for NSR calculations.

The open pit Mineral Reserves are repeated from the 2016 PFS pit designs and the 2016 PFS Mineral Resource models. These estimations include mining loss and dilution that varies by pit ranging from 2.2% to 5.3% for loss, and 0.8% to 3.9% for dilution. A variable cut-off grade strategy has been applied with a minimum NSR of Cdn$9.00/t.

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The mining NSR shut-off is Cdn$15.00/t for the Mitchell underground mine and Cdn$16.00/t for the Iron Cap underground mine. The Mitchell Mineral Reserves include 59 Mt of non-mineralized dilution at zero grade (13%) and 7 Mt of mineralized dilution (2%). The Iron Cap Mineral Reserves include 20 Mt of dilution at zero grade (9%) and 25 Mt of mineralized dilution (11%).

Proven and Probable Mineral Reserves for the KSM mineral deposits with an effective date of July 31, 2016 are stated in Table 1.4.

Table 1.4 KSM Proven and Probable Reserves as of July 31, 2016

|  |  |  | Average Grades | Contained Metal
Zone | Mining Method | Reserve Category | Tonnes (Mt) | Au (g/t) | Cu (%) | Ag (g/t) | Mo (ppm) | Au (Moz | Cu (Mlb) | Ag (Moz) | Mo (Mlb)
Mitchell | Open Pit | Proven | 460 | 0.68 | 0.17 | 3.1 | 59 | 10.1 | 1,767 | 45 | 60
|  | Probable | 481 | 0.63 | 0.16 | 2.9 | 66 | 9.7 | 1,677 | 44 | 70
| Block Cave | Probable | 453 | 0.53 | 0.17 | 3.5 | 34 | 7.7 | 1,648 | 51 | 34
Iron Cap | Block Cave | Probable | 224 | 0.49 | 0.20 | 3.6 | 13 | 3.5 | 983 | 26 | 6
Sulphurets | Open Pit | Probable | 304 | 0.59 | 0.22 | 0.8 | 52 | 5.8 | 1,495 | 8 | 35
Kerr | Open Pit | Probable | 276 | 0.22 | 0.43 | 1.0 | 3 | 2.0 | 2,586 | 9 | 2
Totals |  | Proven | 460 | 0.68 | 0.17 | 3.1 | 59 | 10.1 | 1,767 | 45 | 60
|  | Probable | 1,738 | 0.51 | 0.22 | 2.5 | 38 | 28.7 | 8,388 | 138 | 147
|  | Total | 2,198 | 0.55 | 0.21 | 2.6 | 43 | 38.8 | 10,155 | 183 | 207

Note: All Mineral Reserves stated in Table 1.4 account for mining loss and dilution.

| 1.9.2 | Mine Production Plan

During the initial 33 years of mine life, the majority of ore is derived from open pits, with the tail end of this period supplemented by the initial development of underground block cave mines. After Year 1 ramp up, ore delivery to the mill from Year 2 to Year 35 is designed to be maintained at an average of 130,000 t/d. After depletion of the open pits, the mill processing rate will be reduced to about 96,000 t/d for 10 additional years, before ramping down to just over 61,000 t/d. The change in throughput matches the production levels from the block cave with appropriate ramp ups and ramp downs applied. The remaining few years use stockpile reclaim to supplement the declining production from the block caves at the end of the mine life.

LOM production is summarized in Figure 1.5.

The topographic relief in the areas of the open pits, block cave mines, and the Rock Storage Facilities (RSFs) requires specific geotechnical consideration. Conservative designs, alternative/mitigating scenarios, and extra data and analyses have been included in the mine designs.

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Potential geohazards have been identified in the area of the proposed open pits, block cave mine, RSFs, roads, and other infrastructure; designs include the mitigation of geohazards such as avalanche control, provision of avalanche run-out routes, barriers, and avalanche area and slope hazard avoidance as appropriate.

The mining progression is designed to build RSFs in lifts (bottom-up construction) to consolidate the foundations and reduce downslope risks. Final RSF configurations are designed with terraces at “as dumped” angle of repose, with flat benches between terraces. The overall slope angle is between 26° and 30° to provide the ability for re-sloping to accommodate the end land use and reclamation plan.

Figure 1.5 2016 PFS Mill Feed Production Schedule

Source: MMTS

Ore is mined from Mitchell open pit from Years 1 to 24. Mitchell transitions to block cave mining as the Mitchell pit is mined out. Ore is mined from Sulphurets open pit from Years 1 to 17. Kerr open pit supplements block cave mining from Year 25 to Year 34, and during these years, ore will be transported by an overland conveyor and rope conveyor system starting at the Kerr pit. Mitchell block cave is estimated to have a production ramp-up period of six years, steady state production at 20 Mt/a for 17 years, and then ramp-down production for another 7 years. Iron Cap is estimated to have a production ramp-up period of four years, steady state production at 15 Mt/a for 10 years, and then ramp-down production for another 9 years. The underground pre-production period would be six years, with first underground ore production from Mitchell and Iron Cap in Years 23 and 32, respectively. Figure 1.6 shows the 2016 PFS Open Pit LOM general arrangement.

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Figure 1.6 2016 PFS Open Pit LOM

Source: MMTS

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| 1.10 | Mineral Processing And Metallurgical Testing

Several wide-ranging metallurgical test programs have been carried out since 2007 to assess the metallurgical responses of the mineral samples from the KSM deposits, especially the samples from the Mitchell deposit.

The test results indicate that the mineral samples from the four separate mineralized deposits are amenable to the flotation-cyanidation combined process. The process consists of:

| ● | copper-gold-molybdenum bulk rougher flotation followed by gold-bearing pyrite flotation

| ● | regrinding the bulk rougher concentrate followed by three stages of cleaner flotation to produce a copper-gold-molybdenum bulk cleaner flotation concentrate

| ● | molybdenum separation of the bulk cleaner flotation concentrate to produce a molybdenum concentrate and a copper/gold concentrate containing associated silver

| ● | cyanide leaching of the gold-bearing pyrite flotation concentrate and the scavenger cleaner tailing- to further recover gold and silver values as doré bullion.

The testing programs from 2017 to 2020, as described in Section 13, show the following:

| ● | a finer primary grind size can improve the copper and gold metallurgical performance, especially for copper from the Iron Cap and Deep Kerr samples.

| ● | rougher flotation at a low slurry solid density can improve copper and gold metallurgical performance, especially for the mineralization with more clay-type minerals

| ● | the test results suggest that the gold-bearing sulphide products (first cleaner scavenger tailing and pyrite concentrate) from Deep Kerr and Iron Cap zones did not seem to respond well to the gold recovery by the established cyanide leaching treatment.

| 1.11 | 2016 PFS Recovery Methods

The mill feed from the Mitchell, Sulphurets, Kerr, and Iron Cap deposits will be processed at an average rate of 130,000 t/d. The mill feed will come from open pit mines (upper Mitchell Zone, Sulphurets and upper Kerr Zone) and underground block caving operations (lower Mitchell Zone and Iron Cap deposits). The Mitchell deposit will be the dominant source of mill feed for the process plant and will be processed through the entire mine life, excluding Years 24 and 25. The ore from the Sulphurets deposit will be fed to the plant together with the ore from the Mitchell pit from Years 1 to 17, excluding Years 4, 5, 12, and 13 and with ore from the other deposits during the last four years. Ore from the Kerr deposit, together with ore from the other deposits, will be introduced to the plant during Years 24 to 34, and Year 53, while the Iron Cap ore will be fed to the process plant during Year 32 to the end of mine life.

The 23 km MTT tunnel system has been designed to connect the Mitchell OPC and the Treaty OPC. The crushed ore will be transported through the tunnels by train. This tunnel will also be used for electrical power transmission sourced from the Northwest Transmission Line and for the transport of personnel and supplies for mine operating and water management activities.

The proposed process flow sheet is shown in Figure 1.7.

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Figure 1.7 2016 PFS Simplified Process Flow Sheet

| Source: | Tetra Tech

| Note: | ROM – run-of-mine

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| 1.12 | 2016 PFS Project Infrastructure

| 1.12.1 | Geohazards

Geohazard and risk assessments were completed for the proposed facilities within the KSM footprint. As expected for a mountainous, high-relief project site, snow avalanche and landslide hazards exist, with the potential to affect mine construction, operations, and closure.

Geohazard scenarios were identified for the mine facilities considered. Using unmitigated geohazard levels as a baseline, these scenarios were assessed in terms of risk to human safety, economic loss, environmental loss, and reputation loss. Geohazard risk levels were assigned to each scenario with ratings ranging from very low to very high.

Mitigation strategies have been identified to reduce the high and very high risk scenarios to a target residual risk not exceeding moderate. Further risk reduction will be achieved where practical and cost-efficient.

| 1.12.2 | Tailings Management

The TMF for the 2016 PFS will be constructed in three cells: the North and South cells for flotation tailings, and a lined Central Cell for CIL residue tailings. The cells are confined between four dams (North, Splitter, Saddle, and Southeast dams) located within the Teigen-Treaty creek cross-valley. Design criteria for the dams are based on the Canadian Dam Association (CDA) guidelines. The area is moderately seismic and the dams are designed to resist earthquake loads. A site-specific seismic hazard assessment indicates peak ground acceleration (PGA) at 10,000-year return period of 0.14 g. The TMF cells are designed to store the 30-day probable maximum flood (PMF) with snowmelt, although an operational phase discharge pipeline and closure spillways are also provided to route the critical duration PMF.

De-pyritized flotation tailings will be stored in the North and South cells. The pyrite bearing CIL residue tailings will be stored in the lined Central Cell. In total, the TMF will have a capacity of 2.3 Bt.

The North and Central cells will be constructed and operated first; they will store tailings produced in the first 25 years. The North Cell will then be reclaimed while the Central and South cells are in operation.

The North, Splitter, and Saddle earth-fill starter dams will be constructed over a two-year period, in advance of the start of milling, to form the North and Central cells and will provide start-up tailings storage for two years. Cyclone sand dams will be progressively raised above the starter dams over the operating LOM. The North Starter Dam will be constructed with a low-permeability glacial till core and raised with compacted cyclone sand shells, using the centerline geometry method. The Splitter and Saddle starter dams will form the CIL pond. These dams will also subsequently be raised with cyclone sand shells, with the CIL pond and the Splitter and Saddle dams incorporating high-density polyethylene (HDPE) and linear low-density polyethylene (LLDPE) liners in the core and basin floor in order to surround the CIL tailings within a completely lined impoundment.

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An Independent Geotechnical Review Board (IGRB) was established in January 2015 to independently review and to provide expert oversight, opinion, and advice to Seabridge on the design, construction, operational management, and ultimate closure of the TMF and Water Storage Dam (WSD). The IGRB will review the TMF and WSD on an ongoing basis to ensure that these structures meet internationally accepted standards and practices which effectively minimize risks to employees, lands, and communities.

Seabridge commissioned KCB to undertake the Best Available Tailings Technology (BATT) review in August 2015 in response to the Independent Expert Engineering Investigation and Review Panel report on the breach of the Mount Polley tailings storage facility. The Review Panel concluded that future projects require not only an improved adoption of best applicable practices (BAP), but also a migration to best available technology (BAT). The KCB report (KCB 2016a) also meets the new BC Mining Code requirement that new mines must provide an alternate assessment of BAT in their provincial permit applications.

The BATT study confirmed that the existing TMF design, consisting of centerline dams constructed with double cyclone sand and a till core in association with wet tailings deposition, is the best available technology for tailings deposition, and the most environmentally responsible design to minimize long-term risks associated with the proposed TMF. This conclusion confirms the findings from KSM's IGRB that the TMF's design is robust and appropriate for KSM's site-specific characteristics.

The BATT study also confirmed that the TMF design that was included in the overall project design, that received provincial and federal EA approval, is the best possible design for eliminating risks associated with operation and closure. The study specifically determined that filtered tailings options are impractical for KSM and would result in greater environmental impacts and risks, contrary to the assertions of many environmental groups who have advocated that only filtered tailings disposal technologies should be implemented.

| 1.12.3 | Mine Site Water Management

The overall site water management strategy for the 2016 PFS, including the discharge from the WSF via the High-density Sludge (HDS) water treatment plant (WTP) was the strategy that was reviewed and approved during the EA review process.

Diversion Tunnels and Surface Diversion

Two main diversion tunnel routes will be required to route glacial melt water and non-contact valley runoff from the Mitchell and McTagg valleys around the mine site.

Lined surface diversion channels will be constructed progressively during operations, along the contact of the RSF and the hillside, to divert surface flows.

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Water Storage Facility

All contact water from the mine site areas (open pits, RSFs, roads, infrastructure) will be directed to the WSF, located in the lower Mitchell Creek area. The WSF will be formed with a 165 m high rock fill asphalt core dam built to full height by Year -1 and is sized to store annual freshet flows and volumes resulting from a 200-year wet year. The core zones of the WSF dam will be founded on competent sedimentary rock foundations. Seepage will be controlled by the asphalt core in the dam and the dam foundation will be grouted. A seepage collection pond will return seepage water to the WSF.

Water Treatment

Mine area contact water will be treated with a HDS lime water treatment, of which the discharge from the plant was approved as a component of the environmental assessment review process. A Selenium WTP will be constructed and operational by Year 5 to treat up to 500 L/s of seepage principally from the RSF and select point sources within Mitchell Valley with selenium loading waters, compared to lower concentrations within the WSF.

The HDS WTP and the WSF will be operational before mill start-up to allow pre-production activity in the Mitchell Valley and Mitchell pit area.

| 1.12.4 | Permanent Access Roads

There will be two primary permanent access roads to the mine site and PTMA, the Coulter Creek Access Road (CCAR) and the TCAR, respectively. The CCAR will be primarily a single lane road starting from the former Eskay Creek mine and ending near the HDS WTP at the western extent of mine site disturbance. The TCAR will be a two lane road coming off of Highway 37 that provides access to both the PTMA and the Saddle construction area.

Additional roads will also be required at mine start-up to facilitate maintenance access and construction of the proposed uphill cut-off drainage ditch.

| 1.12.5 | Winter Access Road

A winter access road will be constructed during the first construction year to provide the initial access to the KSM mine site for logistics support.

| 1.12.6 | Off-Site Infrastructure

Copper concentrates produced at the Treaty OPC will be filtered at the processing plant and transported by contract trucking firms via Highway 37 and 37A to one of the port vendors in Stewart, BC.

| 1.12.7 | Tunneling

A total of nine major tunnels will be excavated during the pre-production period and during operations. These tunnels will be classified as either infrastructure tunnels or water tunnels. This does not include development work for the block caves.

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Engineering for the major components of the MTT has been developed by two experienced tunnel contractors and have been adapted to form the advance rate and cost estimates for the infrastructure and water tunnels.

The tunnels will be driven in accordance with the BC Mines Act and Regulations using mechanized drill and blast techniques and will follow the conditions contained within the License of Occupation.

The detailed contractor tunneling estimate for the MTT has been adapted to estimate the excavation costs and advance rates for the other associated excavations in the MTT, as mentioned above. They have also been used for the other infrastructure and water tunnels.

| 1.12.8 | Power Supply and Distribution

Power supply for KSM will be from BC Hydro’s existing NTL transmission line. Interconnection to the site will be via a new 30 km long, 287 kV, spur transmission line to be constructed from the Treaty OPC to a new BC Hydro switching station on the NTL, to be located beside Highway 37 at a point near the intersection of the Treaty Creek Access Road and the highway. Refer to Section 18.10 herein.

The power supply will be supplemented by several energy recovery plants. Refer to Section 18.10.12.

| 1.13 | Environmental Studies, Permitting, and Social or Community Impact

The KSM mine development is subject to the BC Environmental Assessment Act, the Canadian Environmental Assessment Act, and Chapter 10 of the Nisga’a Final Agreement (NFA).

As of January 2020, KSM has successfully gone through the provincial and federal environmental assessment review processes, and the appropriate certificates/approvals have been obtained. Additionally, permits for early-stage construction activities, continuation of exploration, and certain permit and project approval renewals have also been obtained. Seabridge continues to advance permitting to allow for the construction of the Property, as well as to expand exploration activities. Details of the provincial, federal, and NFA processes and current statuses, as well as the current permitting status of KSM, are included in Section 20.0.

KSM underwent a harmonized EA process with the provincial and federal governments, in accordance with the principles of the Canada-BC Agreement on Environmental Assessment Cooperation (Cooperation Agreement 2004). The process included a working group comprising federal and provincial officials, the Nisga’a Lisims Government (NLG), Aboriginal groups, and local government agencies. Representatives of the US federal and Alaska state agencies were extensively involved in the EA review processes, as a matter of courtesy at the insistence of Seabridge, given that the mineral deposits are located on a tributary of the Unuk River, a transboundary river, 30 km upstream of the US/Canada border. Authorizations are not required from any US federal or state regulatory agency for KSM to proceed into construction and operation.

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| 1.13.1 | Benefit Agreements

Tahltan Nation

On June 10, 2019 the Tahltan Nation and Seabridge Gold Inc. announced that the parties reached agreement on the terms of a Co-operation and Benefits Agreement in connection with KSM. Tahltan Nation voted 77.8% in favour of the KSM Project Impact Benefits Agreement (IBA). The IBA provides a thorough and co-operative framework for the parties to continue building the social licence of KSM through commitments to economic benefits and environmental management of the land.

Nisga’a Nation

On June 17, 2014, Seabridge Gold entered into a comprehensive Benefits Agreement with the Nisga’a Nation in respect of Seabridge Gold’s KSM Property located in northwest British Columbia. The Benefits Agreement established a long-term co-operative relationship between Seabridge and the Nisga’a Nation under which the Nisga’a Nation will support development of KSM, participate in economic benefits from KSM and provide ongoing advice. The Agreement includes commitments by Seabridge regarding jobs and contracting opportunities at KSM, education and training, financial payments and a framework for working together on ongoing development matters. This comprehensive agreement also addresses concerns expressed by the Nisga’a Nation around the potential environmental and social impacts of KSM.

| 1.13.2 | Closure and Reclamation

In the 2016 PFS, the KSM Mine will be closed in accordance with the closure plan outlined in Section 20.7, and in further detail in the Application/EIS (Rescan 2013).

The closure and reclamation plan has three objectives that provide assurance to the province that the site will be left in a condition that will limit any future liability to the people of BC:

| ● | to provide stable landforms

| ● | to re-establish productive land use

| ● | to protect terrestrial and aquatic resources.

| 1.14 | 2016 PFS Capital Cost Estimate

An initial capital of US$5.005 billion is estimated for the 2016 PFS. All currencies in are expressed in US dollars, unless otherwise stated. Costs have been converted using a fixed currency exchange rate of US$0.80 to Cdn$1.00. The expected accuracy range of the capital cost estimate is +25%/-10%.

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This capital cost estimate includes only initial capital, which is defined as all capital expenditures that are required to produce concentrate and doré. A summary of the capital costs is shown in Table 1.5.

Table 1.5 2016 PFS Initial Capital Cost Summary

Major Area No. | Major Area Description | Cost (US$ M)
1 – Direct Costs
1.1 | Mine Site | 1,218
1.2 | Process | 1,336
1.3 | TMF | 441
1.4 | Environmental | 15
1.5 | On-site Infrastructure | 23
1.6 | Off-site Infrastructure | 120
1.7 | Permanent Electrical Power Supply and Energy Recovery | 159
Total Direct Costs | 3,312
2 – Indirect Costs
2.91 | Construction Indirect Costs | 449
2.92 | Spares | 34
2.93 | Initial Fills | 20
2.94 | Freight and Logistics | 99
2.95 | Commissioning and Start-up | 6
2.96 | EPCM | 231
2.97 | Vendor’s Assistance | 23
Total Indirect Costs | 862
3 – Owner’s Costs
3.98 | Owner’s Costs | 160
4 – Contingency
4.99 | Contingency | 671
2016 PFS Capital Cost Total | 5,005

Notes: Costs have been rounded to the nearest million dollars.

Capital costs exclude reclamation and closure costs that are accounted for separately in the economic analysis. Sustaining capital costs were also estimated leveraging the same basis of information applied to the initial capital estimate with respect to vendor quotations, labour, and material costs. The sustaining capital costs total US$5.503 billion and consist of:

| ● | open pit mine development, principally mobile fleet replacement

| ● | underground mine development at Mitchell and Iron Cap block cave mines

| ● | process improvements principally at Mitchell and Treaty OPC, and SMCT

| ● | TMF expansions, mainly comprising dam raises and CIL basin expansions

| ● | permanent electrical power supply and energy recovery systems

| ● | indirect costs, including construction indirects, spares, freight and logistics, EPCM, vendor assistance, and contingency.

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| 1.15 | 2016 PFS Operating Cost Estimate

The average operating cost for the 2016 PFS is estimated at US$12.03/t milled at the nominal process rate of 130,000 t/d or US$12.33/t for the LOM average as shown in Table 1.6.

The cost estimates in this section are based upon budget prices in Q1/Q2 2016 or based on the data from the database of the consulting firms involved in the cost estimates. When required, costs in this report have been converted using a three-year average currency exchange rate of US$0.80 to Cdn$1.00. The expected accuracy range of the operating cost estimate is +25%/-10%.

The estimates do not include energy recovery credit (approximately US$0.12/t milled LOM) from mini hydropower stations and the cost (approximately US$0.15/t milled LOM) related to Provincial Sales Tax (PST).

Table 1.6 2016 PFS Average Operating Cost Summary

| At the Nominal Feed Rate of 130,000 t/d* | LOM Average (US$/t milled)
(US$ M/a) | (US$/t milled)
Mine
Mining Costs – Mill Feed | 190.2 | 4.59 | 4.59
Open Pit – Mill Feed | - | 4.40 | 4.40
Block Caving – Mill Feed | - | 4.99 | 4.99
Mill
Process | 251.1 | 5.29 | 5.34
G&A and Site Service
G&A | 43.3 | 0.91 | 1.03
Site Service | 18.9 | 0.40 | 0.44
TMF and SWM
TMF Dam Management | 6.1 | 0.13 | 0.13
Selenium Water Treatment | 9.4 | 0.20 | 0.21
HDS Water Treatment | 22.0 | 0.46 | 0.53
Mine Site Water Pumping | 2.5 | 0.05 | 0.06
Total Operating Cost | 543.5 | 12.03 | 12.33

Notes: *The nominal feed rate estimate excludes mine operating costs and is based on a mill feed rate of 130,000 t/d; the costs do not reflect higher unit costs late in the mine life when the mill feed rates are lower. Costs have been rounded to the nearest hundred thousands of dollars.

The operating costs are defined as the direct operating costs including mining, processing, tailings storage, water treatment, and G&A. The hydropower credit from recovered hydro-energy during mining operations is not accounted for directly against operating cost estimate, but is included in the economic analysis. Sustaining capital costs including all capital expenditures after the process plant has first been put into production are excluded from the operating cost estimate.

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| 1.16 | 2016 PFS Economic Evaluation

Tetra Tech prepared an economic evaluation for the 2016 PFS based on a pre-tax financial model. For the 53-year mine life and 2,198 Mt Mineral Reserve, the following pre-tax financial parameters were calculated using the 2016 Base Case metal prices:

| ● | 10.4% IRR

| ● | 6.0-year payback on US$5.005 billion capital

| ● | US$3.263 billion NPV at a 5% discount rate.

The tax component of the financial model used for the post-tax economic evaluation was prepared and reviewed by other consultants (see Section 22.0 for further details).

Based on the tax analysis and review, the following post-tax financial results were calculated:

| ● | 8.0% IRR

| ● | 6.8-year payback on US$5.005 billion capital

| ● | US$1.539 billion NPV at a 5% discount rate.

The 2016 Base Case results incorporate historical three-year trailing averages for metal prices as of July 31, 2016 as follows:

| ● | gold – US$1,230/oz

| ● | copper – US$2.75/lb

| ● | silver – US$17.75/oz

| ● | molybdenum – US$8.49/lb

| ● | exchange rate –US$0.80 to Cdn$1.00.

Metal revenues projected in the KSM cash flow models were based on the average metal values indicated in Table 1.7.

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Table 1.7 2016 PFS Metal Production

| Years 1 to 7 | LOM
Total Tonnes to Mill (000s) | 323,000 | 2,199,000
Annual Tonnes to Mill (000s) | 46,100 | 41,500
Average Grades
Gold (g/t) | 0.82 | 0.55
Copper (%) | 0.24 | 0.21
Silver (g/t) | 2.8 | 2.6
Molybdenum (ppm) | 48 | 43
Total Production
Gold ('000 oz) | 6,530 | 28,600
Copper ('000 lb) | 1,435,000 | 8,270,000
Silver ('000 oz) | 18,200 | 114,700
Molybdenum ('000 lb) | 11,200 | 62,100
Average Annual Production
Gold ('000 oz) | 933 | 540
Copper ('000 lb) | 205,000 | 156,000
Silver ('000 oz) | 2,600 | 2,160
Molybdenum ('000 lb) | 1,590 | 1,170

In addition to the 2016 Base Case, three metal price/exchange rate scenarios were also developed: the first uses the metal prices and exchange rate used in mine optimization and design (2016 Design Case); the second uses the spot metal prices and closing exchange rate on July 1, 2016 (2016 Spot Case); the third uses higher metal prices to indicate upside potential (2016 Alternate Case). The input parameters and pre- and post-tax results of all scenarios can be found in Table 1.8.

Table 1.8 2016 PFS Summary of the Pre- and Post-tax Economic Evaluations

| Unit | 2016 Base Case | 2016 Design Case | 2016 Spot Case | 2016 Alternate Case
Metal Price
Gold | US$/oz | 1,230.00 | 1,200.00 | 1,350.00 | 1,500.00
Copper | US$/lb | 2.75 | 2.70 | 2.20 | 3.00
Silver | US$/oz | 17.75 | 17.50 | 20.00 | 25.00
Molybdenum | US$/lb | 8.49 | 9.70 | 7.00 | 10.00
Exchange Rate | US$:Cdn$ | 0.80 | 0.83 | 0.77 | 0.80
|  |  | table continues…

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| Unit | 2016 Base Case | 2016 Design Case | 2016 Spot Case | 2016 Alternate Case
Pre-tax Economic Results
Cumulative Cash Flow | US$ million | 15,933 | 13,727 | 16,101 | 26,319
NPV (at 3%) | US$ million | 6,217 | 5,128 | 6,461 | 11,138
NPV (at 5%) | US$ million | 3,263 | 2,510 | 3,507 | 6,541
NPV (at 8%) | US$ million | 960 | 475 | 1,175 | 2,928
IRR | % | 10.4 | 9.2 | 11.1 | 14.6
Payback | years | 6.0 | 6.5 | 5.6 | 4.1
Pre-tax Economic Results
Cash Cost/oz Au* | US$/oz | 277 | 311 | 404 | 183
Total Cost/oz Au* | US$/oz | 673 | 720 | 787 | 580
Post-tax Economic Results
Cumulative Cash Flow | US$ million | 9,983 | 8,537 | 10,109 | 16,721
NPV (at 3%) | US$ million | 3,513 | 2,789 | 3,691 | 6,696
NPV (at 5%) | US$ million | 1,539 | 1,028 | 1,718 | 3,663
NPV (at 8%) | US$ million | -2 | -343 | 161 | 1,282
IRR | % | 8.0 | 7.0 | 8.5 | 11.4
Payback | years | 6.8 | 7.4 | 6.4 | 4.9

Note: *Net of by-product

| 1.16.1 | Sensitivity Analysis

Sensitivity analyses were carried out on the pre-tax 2016 Base Case for the following parameters:

| ● | gold, copper, silver, and molybdenum metal prices

| ● | exchange rate

| ● | capital expenditure

| ● | operating costs.

The 2016 PFS NPV is most sensitive to gold price and exchange rate followed by operating costs, copper price and capital costs. The NPV, IRR, and payback sensitivities can be seen in Figure 1.8, Figure 1.9, and Figure 1.10.

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Figure 1.8 2016 PFS Sensitivity Analysis of Pre-tax NPV at a 5% Discount Rate

Source: Tetra Tech, 2016

Figure 1.9 2016 PFS Sensitivity Analysis of Pre-tax IRR

Source: Tetra Tech, 2016

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Figure 1.10 2016 PFS Sensitivity Analysis of Pre-tax Payback Period

Source: Tetra Tech, 2016

| 1.17 | 2020 Preliminary Economic Assessment

| 1.17.1 | Introduction

The 2020 PEA has been undertaken to evaluate an alternative approach to developing the KSM Property by incorporating high value block cave mines at Iron Cap and Deep Kerr deposits, with reduced open pit production and associated surface disturbance, and an increased processing rate.

Iron Cap and Deep Kerr Mineral Resources have expanded since the 2016 PEA.

Mining of Mineral Resources from open pit and underground are limited in the 2020 PEA plan to a capacity of 2.4 Bt, which is less than the potential economic limits of mining for the 2020 PEA and which maintains the existing footprint of the approved tailings facility. In addition, mine planning optimization within the larger (effective date December 31, 2019) Mineral Resources has resulted in increased mill feed grades and an improved production schedule compared to the 2016 PEA. The mineral deposits remain open at the end of the 2020 PEA mine plan presented in this report.

The 2020 PEA is a scoping level of study based on the Mineral Resources and include Inferred Mineral Resources in the 2020 PEA design and projected economics.

The 2020 PEA envisages a combined open pit/underground block cave mining operation that is planned to operate for 44 years.

The proposed process plant for the 2020 PEA design has an average process rate of 170,000 t/d. The 2020 PEA is based on a throughput 31% greater than stated in the approved EA as an initial case for evaluating improved business cases that maintain reasonable and achievable mining and processing rates. Future mine to mill and process throughput studies will optimize the mining and processing rates for further business case improvement.

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Mitchell open pit is the sole source of mill feed during initial capital payback. The mill feed then transitions to Iron Cap and Deep Kerr underground mines, supplemented by Kerr and Sulphurets open pit during the transition. There is no Mitchell underground production in the 2020 PEA.

The flotation plant produces a gold/copper/silver concentrate for transport by truck to a nearby seaport at Stewart for shipment to overseas smelters. Metallurgical testing indicates that KSM can produce a clean concentrate at an average copper grade of 25% with a high gold and silver content. Separate gold-silver doré is also produced at the KSM processing facility from gold rich pyrite products rejected from the gold/copper flotation. No molybdenum concentrate is produced in the 2020 PEA.

The 2020 PEA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the results of the 2020 PEA will be realized. Mineral Resources in the 2020 PEA mine plan are not Mineral Reserves and do not have demonstrated economic viability.

The results of the 2016 PFS remain valid and represent a viable option for developing KSM, with the 2020 PEA assessing an alternative development option at a scoping level of design.

| 1.17.2 | Mining Methods

The 2020 PEA utilizes Measured, Indicated, and Inferred Mineral Resources in mine planning. Mill feed in the 2020 PEA is drawn from open pit mining of the Mitchell, Sulphurets and Kerr deposits and underground block cave mining of the Iron Cap and Deep Kerr deposits. Approximately 34% of the mill feed would come from open pit operations and 66% from underground block cave operations. Waste to mill feed cut-offs were determined using a NSR for each block in the model for the open pit mines, and a NSR shut-off for the block cave underground mines. NSR is calculated using prices and process recoveries for each metal accounting for all off-site losses, transportation, smelting and refining charges. Metal prices of US$1,200/oz of gold, US$2.70/lb of copper, US$17.50/oz of silver and a fixed currency exchange rate of US$0.83 to Cdn$1.00 are used in the NSR calculations.

LG pit shell optimizations were used to define open pit mine plans in the 2020 PEA. The pit limits of the 2020 PEA are contained inside the pit limits of the 2016 PFS. The mine design for the 2020 PEA focuses on reducing waste and selecting blocks with higher values. As a result, the 2020 PEA mine plan contains 1.7 Bt less waste in the open pit mine plan compared to the 2016 PFS.

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An elevated shut-off is used in the 2020 PEA underground mine designs compared to what was used in the 2016 PFS relative to the mining operating cost. In the 2020 PEA, the mining NSR shut-off is Cdn$20/t for the Iron Cap underground mine and Cdn$18/t for Deep Kerr.

The 2020 PEA mine plan was carried out with the aim of reducing the amount of waste rock produced in the open pits with the mill feed drawing more on the underground resources. The mine production plan starts in lower-cost open pit areas using conventional large scale equipment before transitioning into block cave underground bulk mining later in the mine life. Starter pits have been selected in higher-grade areas and cut-off grade strategy optimizes revenues to minimize the payback duration. Smaller pits were designed for the Mitchell, Sulphurets and Kerr deposits as well as mining the Iron Cap and Deep Kerr deposits by underground mining methods. This approach shrinks the KSM mine site footprint.

| 1.17.3 | Open Pit Mining

The 2020 PEA open pit mining is designed as a conventional truck-shovel operation. The pit design for the Mitchell deposit includes three nested phases to balance stripping requirements while satisfying the process plant requirements. In the case of the Sulphurets and Kerr deposits, the materials are mined in one phase. The 2020 PEA pit delineated resources summarized in Table 1.9 are a subset of the Mineral Resources and quantities that are included in Section 14.0. The pit delineated resources use a variable NSR cut-off grade varying between Cdn$15/t and Cdn$25/t for Mitchell, and a constant NSR cut-off grade of Cdn$17/t for Kerr and Sulphurets.

Table 1.9 Pit Delineated 2020 PEA Mine Plan Quantities

Pit | Mill Feed (Mt) | Au (g/t) | Cu (%) | Ag (g/t) | Waste (Mt) | Strip Ratio
Mitchell Pit Phase M1 | - | - | - | - | 33 | n/a
Mitchell Pit Phase M2 | 467 | 0.81 | 0.21 | 3.4 | 565 | 1.2
Mitchell Pit Phase M3 | 145 | 0.63 | 0.16 | 3.4 | 238 | 1.6
Mitchell Pit Total | 612 | 0.77 | 0.20 | 3.4 | 836 | 1.4
Sulphurets Pit | 105 | 0.73 | 0.31 | 1.0 | 233 | 2.2
Kerr Pit | 89 | 0.31 | 0.59 | 1.5 | 252 | 2.8
Total | 807 | 0.71 | 0.25 | 2.9 | 1,321 | 1.6

Note: Numbers may not add due to rounding.

The 2020 PEA production schedule results in an open pit production of 25 years (split into two main time periods, Years 1 to 16 and Years 26 to 34) with stockpile reclaim supplementing the mill feed from Year 5 to 20. The mine will require four years of pre-production before the start of the process plant operations. Eight 363 t haul trucks will be required in Year -2, increasing to 34 by Year 1, and peaking at 62 in production Years 6 to 10.

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| 1.17.4 | Underground Mining

The 2020 PEA underground block caving mine designs for Iron Cap and Deep Kerr are based on modeling using GEOVIA’s PCBC™ Footprint Finder software. Iron Cap has been designed for using battery-electric loaders and electric trains for material handling and employs both tele-operation and automation for these units. This enables a reduction to the number of primary ventilation intake and exhaust drifts due to less ventilation demand and cost savings from lower diesel fuel, labour and equipment maintenance costs.

Iron Cap is estimated to have a production ramp-up period of six years, steady state production at 33 Mt/a for 17 years, and then ramp-down production for another six years. Deep Kerr is estimated to have a production ramp-up period of five years, steady state production at 29 Mt/a for 20 total years with a seven year production dip during years where the operation transitions from the first to second lift. The underground pre-production period is four years for Iron Cap and five years for Deep Kerr.

The first underground mill feed production from Iron Cap and Deep Kerr comes in Years 6 and 12, respectively. Table 1.10 shows the production tonnes and grade in the 2020 PEA mine plan. Only Indicated and Inferred Mineral Resources are included in the production plan.

Table 1.10 Production Tonnes and Grade in 2020 PEA Underground Mine Plan

| Iron Cap | Deep Kerr
Mill Feed (Mt) | 742.7 | 825.1
Au (g/t) | 0.58 | 0.29
Cu (%) | 0.35 | 0.46
Ag (g/t) | 3.0 | 1.6

| 1.17.5 | Mine Production Plan

The 2020 PEA mine production plan starts in lower-cost open pit areas using conventional large scale equipment before transitioning into block cave underground bulk mining later in the mine life. Starter pits have been selected in higher-grade areas and cut-off grade strategy optimizes revenues to minimize the payback duration.

After initial ramp-up the throughput averages of 170,000 t/d for the first 33 years reflecting mill fed from multiple mine sources. Then the production rate is reduced to 120,000 t/d in Year 34 and then further reduced to around 80,000 t/d for the following 10 years as mill feed is provided from a single mine source. The 2020 PEA mine life is estimated at 44 years. Mitchell open pit production takes place from Year 1 to Year 15 except for Year 8, 11 and 12 and then resumes in Year 29 and lasts until Year 34. Kerr open pit production starts in Year 6 and Sulphurets in Year 9 and last until Years 9 and 16 of production, respectively.

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At Mitchell, a near-surface higher-grade gold zone outcrops allowing for gold production in the first six years that is substantially above the mine life average grade. The mine plan is specifically designed for mining highest gold grade first to facilitate an early capital investment payback.

Development of the Iron Cap cave starts in Year 2 and first mill feed is produced in Year 6 of production. Deep Kerr development starts sequentially to that of Iron Cap and delivers first mill feed in Year 12. Underground production ends first at Iron Cap in Year 34 and finally at Kerr in Year 44 of production.

| 1.17.6 | Recovery Methods

The proposed process plant for the 2020 PEA will have an average process rate of 170,000 t/d. The plant will receive mill feed from the Mitchell, Kerr and Sulphurets open pits and the Iron Cap and Deep Kerr underground mines. The Mitchell open pit and Iron Cap and Deep Kerr underground mines will be the main source of mill feed, contributing about 92% of the total plant feed over the LOM, as follows:

| ● | 25.8% Open pit – Mitchell

| ● | 31.3% Underground mine - Iron Cap

| ● | 34.7% Underground mine - Deep Kerr.

The 2020 PEA process circuit incorporates three-stage crushing (tertiary crushing by HPGR), milling, conventional flotation and cyanidation processes for the recovery of copper, gold and silver, with equipment sized for the design throughput. Recent metallurgical testing on the copper rich Iron Cap and Deep Kerr mill feed suggests incremental CIL gold production is marginal or sub-economic. Thus, the process plant is split into two lines which are able to separately process Mitchell/Sulphurets and Iron Cap/Kerr feed. Characterization and metallurgical testwork on Mitchell, Sulphurets, Iron Cap, and Deep Kerr samples is presented in Section 13.0.

| 1.17.7 | Capital and Operating Costs

Initial capital costs of the 2020 PEA (including contingency of US$859 million and pre-production mining costs) are estimated at US$5.2 billion, approximately 3.0% higher than the initial capital estimate in the 2016 PFS. This increased initial capital is related to the higher plant throughput that required increased mine production in pre-production for the WSD construction and higher pre-stripping costs in Mitchell pit, larger size of equipment at the mill and changes in the tailings management facility due to a higher mill rate. Also, contingency is higher to reflect the lower level of cost accuracy of the 2020 PEA compared to the 2016 PFS. Initial capital and sustaining capital for the 2020 PEA are estimated to an AACE® Class 5 Estimate classification.

Sustaining capital over the 44-year mine life is estimated at US$6.9 billion and is dominated by capitalizing the underground mine expansions at Iron Cap and Deep Kerr block caves. In addition to sustaining capital, a further US$938 million has been charged against the 2020 PEA for closure and post-closure costs including US$526 million set aside in a sinking fund from operating cash flow to pay for estimated post-closure water treatment obligations. Initial capital and sustaining capital estimates for the 2020 PEA are summarized in Table 1.11.

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Table 1.11 2020 PEA Capital Cost Estimate Summary

| Initial | Sustaining | Total
Area | US$ M | US$ M | US$ M
Direct Costs |  |  |
Mine | 1,193 | 4,406 | 5,599
Process | 1,359 | 190 | 1,549
TMF | 488 | 509 | 998
Environmental | 14 | - | 14
On-site Infrastructure | 22 | - | 22
Off-site Infrastructure | 106 | 10 | 116
Power Supply/Energy Recovery | 159 | 214 | 373
Total Direct Capital | 3,342 | 5,330 | 8,672
Indirect cost | 802 | 648 | 1,450
Owners cost | 152 | - | 152
Contingency | 859 | 907 | 1,767
Total Estimated Total | 5,156 | 6,886 | 12,041

Note: Numbers may not add due to rounding.

Average mine, process, and G&A operating costs over the 2020 PEA calculated mine life (including waste mining and on-site power credits, excluding off-site shipping and smelting costs) are estimated at US$9.98/t milled (before base metal credits). Estimated unit operating costs are 19% lower than the 2016 PFS primarily due to reduction in process and G&A cost associated with higher throughput and elimination of cyanide leaching for the Iron Cap and Deep Kerr mineralization. A breakdown of estimated unit operating costs is shown in Table 1.12.

Table 1.12 2020 PEA LOM Average Unit Operating Costs

Area | Cost (US$/t milled)
Mining | 3.91
Milling, TMF and train transport | 4.32
G&A | 0.89
Others | 0.85
Total Operating Costs | 9.98

Notes: 1. Mining operating cost excludes pre-production cost of both open pit and underground mining.

2. Others includes site services, water management, annual energy recovery credit and PST.

3. Numbers may not add due to rounding.

| 1.17.8 | Economic Analysis

The 2020 PEA has been evaluated using a discounted cash flow (DCF) analysis. Cash inflows consist of annual revenue projections for the mine. Cash outflows such as capital (including pre-production costs), operating costs, taxes, and royalties are subtracted from the inflows to arrive at the annual cash flow projections. Cash flows are taken to occur at the end of each period.

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Under the assumptions presented in this report, the 2020 PEA demonstrates positive economics. In the 2020 PEA Base Case, the after-tax NPV at a 5% discount rate is US$ 6.01 billion with an after-tax IRR of 14.0%. Payback of the initial capital investment is estimated to occur in 4.0 years after the start of production or 9% of mine life. The 2020 PEA Economic Results are presented in Table 1.13.

Table 1.13 2020 PEA Economic Results

Financial Results | 2020 PEA Comparative Case | 2020 PEA Base Case 3yr Avg. | 2020 PEA Alternate Upside Case
Metal Prices: |  |  |
Gold (US$/ounce) | 1,230 | 1,340 | 1,700
Copper (US$/pound) | 2.75 | 2.80 | 2.50
Silver (US$/ounce) | 17.75 | 16.25 | 16.00
Exchange Rate: (US$/Cdn$) | 0.80 | 0.76 | 0.72
Pre-Tax Results:
Net Cash Flow (US$ Billion) | 31.34 | 36.79 | 43.40
NPV @ 5% Discount Rate (US$ Billion) | 8.20 | 10.26 | 13.22
Internal Rate of Return (%) | 14.9% | 17.5% | 21.9%
Payback Period (Years) | 4.3 | 3.6 | 2.9
Post-Tax Results:
Net Cash Flow (US$ Billion) | 19.58 | 23.03 | 27.22
NPV @ 5% Discount Rate (US$ Billion) | 4.67 | 6.01 | 7.92
Internal Rate of Return (%) | 11.9% | 14.0% | 17.4%
Payback Period (Years) | 4.9 | 4.0 | 3.2

The 2020 PEA financials are most sensitive to changes in metal prices and grade and less sensitive to changes in capital costs and operating costs.

| 1.17.9 | Conclusions

The 2020 PEA offers a viable option for KSM development and reduces a number of the project risks. By including large block cave mines at Iron Cap and Deep Kerr, supplemented by a production plan starting with open pits, annual average maximum throughput of 130,000 t/d envisioned in the 2016 PFS could be increased to 170,000 t/d in the 2020 PEA to yield a viable KSM mine to mill plan. Increased throughput boosts the metal production, reduces the payback period and improves estimated NPV and IRR. The 2020 PEA mine plan in total would reduce the amount of waste rock by ~57% (approximately 1.7 Bt) compared to the 2016 PFS, shrinking the KSM mine site footprint and its environmental impact.

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| 1.18 | Recommendations

Based on the work carried out in the 2016 PFS and the 2020 PEA, and the resultant economic evaluations, the next phase of KSM development should comprise more advanced studies in order to further assess the technical and economic viability of KSM. Specific recommendations made by Qualified Persons (QPs) are detailed in Section 26.0 and are briefly summarized in the following subsections.

| 1.18.1 | 2016 PFS Recommendations

The key recommendations for advanced studies emanating from the 2016 PFS focus on improving both open pit and underground mine design through additional drilling and testing; water related topics to further refine the inputs and results of site wide water balance analyses from the construction period through closure; and tunnels to develop more design-specific information to assist in reducing construction and operational risks and associated capital and operating costs. Other recommendations address data collection needs for RSFs, ore transportation, TMF, metallurgical testing, and process engineering. The estimated total cost to implement the 2016 PFS recommendations in Section 26 ranges between US$35 to US$45 million.

| 1.18.2 | 2020 PEA Recommendations

RMI recommends an infill drilling program on Iron Cap and Deep Kerr for upgrading the Inferred Mineral Resources that are included in the 2020 PEA mine plan, RMI also recommends developing an exploration adit to Iron Cap to provide underground drill access and opportunities in bulk sampling, geotechnical data collection and geologic mapping.

MMTS recommends design work to a FS level focusing on:

| ● | pre-stripping and for the first five years of operation

| ● | pit phase design with revised geotechnical parameters

| ● | fleet optimization and power technologies with OEM budgetary quotes

| ● | evaluate backfill into the mined out Mitchell pit

| ● | evaluate basal drain design to reduce required quantities

| ● | refine early year production sequencing and cutoff grade strategies to enhance project economics.

Golder recommends the following works for underground block cave mining of the Iron Cap and Deep Kerr deposits for the next level of study:

| ● | undertake a geotechnical drilling and logging program for data collection and rock characterization

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| ● | undertake a detailed study of the optimum pit profile at Deep Kerr to transition from open pit to underground and assess the optimum drawpoint layout and spacing

| ● | undertake additional laboratory rock strength tests. Evaluate the impact of fragmentation, oversize, and hang ups on productivity, develop a hydraulic fracturing plan and conduct stress analyses

| ● | assess alternative development techniques, conduct mine productivity simulations and develop designs of underground infrastructure

| ● | evaluate the use of electric vehicles for Deep Kerr and optimize the train haulage level and train operations at Iron Cap.

Wood recommends that Seabridge perform further metallurgical test work, especially for the Iron Cap and Deep Kerr mineralization investigating the metallurgical response of various zones to both flotation and cyanidation.

KCB recommends that a geotechnical site investigation program be carried out for further delineation and characterization of foundation soils in Mitchell Valley and additional hydrogeological and geotechnical site investigations in the WSD and TMF footprint areas.

The estimated total cost to implement the 2020 PEA recommendations in Section 26.0 ranges between US$174 to US$205 million.

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| 2.0 | Introduction

| 2.1 | Overview

The KSM Property is currently 100%-owned by Seabridge. Description and location of the KSM Property is presented in Section 4.0 of this report.

| 2.2 | Terms of Reference

This report was prepared for Seabridge to summarize the updated Mineral Resource estimate on the Property and results of the data verification conducted on the 2016 PFS, as well as a 2020 PEA update of an alternative development option, performed at a conceptual level, of the previous studies.

This 2016 KSM PFS data verification and 2020 KSM PEA Update have been prepared for Seabridge based on work performed by the following independent consultants:

| ● | Tetra Tech

| ● | Wood

| ● | MMTS

| ● | Golder

| ● | McElhanney

| ● | BGC

| ● | RMI

| ● | KCB

| ● | Brazier

| ● | ERM.

| 2.2.1 | Mineral Resource Estimate Update

The Mineral Resource estimate has been updated in 2019 and is presented in Section 14.0 of this report.

| 2.2.2 | 2016 Prefeasibility Study Data Verification

The 2019 Mineral Resource estimate is the current and only Mineral Resource estimate used in this Report. To assess whether the updated 2019 information to date would materially change the 2016 PFS results, several verification checks have been made to confirm the validity of the 2016 PFS. Details and results of the checks are presented in Section 12.0 of this report.

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| 2.2.3 | 2020 Preliminary Economic Assessment Update

The results of the 2020 PEA Update were disclosed in Seabridge’s press release dated April 27, 2020. This Report will be filed in support of the disclosure of the PEA results.

| 2.3 | Sources of Information

The key information sources for the Report were:

| ● | previously filed technical report titled 2016 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update and Preliminary Economic Assessment, effective date Oct 06, 2016

| ● | documents referenced in Section 3.0 (Reliance on Other Experts)

| ● | documents referenced in Section 27.0 (References) of this Report

| ● | additional information provided by Seabridge personnel where required.

| 2.4 | Effective Dates

The Report has a number of effective dates as follows:

| ● | KSM Mineral Resource estimate: December 31, 2019

| ● | KSM Mineral Reserve estimate: July 31, 2016

| ● | Latest information on mineral tenure, surface rights and Property ownership: February 25, 2020

The overall effective date of the Report is April 30, 2020.

| 2.5 | Qualified Persons

The name of the QPs of this report and their QP certificates are included at the end of the report.

| 2.6 | Site Visits

The following QPs conducted a site visit of the Property:

| ● | Derek Kinakin (M.Sc., P.Geo., P.G.) of BGC visited the Property from August 13 to 17, 2018 and from August 19 to 21, 2013. The site visit involved a general review of the site conditions for the Mitchell and Sulphurets deposits, and detailed inspection and qualitative geotechnical assessment of the core from many of the holes that had been drilled up to that time.

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| ● | Hassan Ghaffari (P.Eng.) of Tetra Tech visited the Property on September 20, 2014 and conducted a general site overview in the proposed PTMA area.

| ● | James H. Gray (P.Eng.) of MMTS visited the Property from June 21 to 22, 2017, and also during 2008, 2009 and 2010. The 2008, 2009 and 2017 visits were general site overview of the potential open pit and rock storage sites and water management structures. Site access alternatives and the route of the MTT were also viewed. The 2010 visit was specifically to inspect the open pit mining areas before the winter snowpack receded.

| ● | Michael J. Lechner (P. Geo., RPG, CPG) of RMI last visited the Property from October 5 to October 7, 2019. During the site visit, the geologic models were reviewed with Seabridge’s geologic staff, drill core examined and mineralized outcrops at the Mitchell deposit were inspected.

| ● | Jianhui (John) Huang (Ph.D., P.Eng.) of Tetra Tech visited the Property from June 21 to 22, 2017, from April 27 to 28, 2017 and on September 16, 2008 and conducted an overview of the proposed general project and process plant sites and inspected the drill cores.

| ● | Neil Brazier (P.Eng.) of Brazier visited the Property from September 12 to 16, 2011, July 20 to 25, 2012 and from September 1 to 4, 2013 and inspected transmission line routes, the plant-site area and locations of energy recovery projects.

| ● | J. Graham Parkinson (P.Geo.) of KCB visited the Property from September 10 to 14, 2018 as well as during 2007, 2008, 2009 and 2014. The purposes of these site visits included geotechnical site investigations and infrastructure planning.

| ● | Rolf Schmitt (P.Geo.) of ERM visited the property on July 12, 2019, via helicopter access accompanied by Jessy Chaplin, Seabridge Director, Permitting. The visit entailed a fly-over of the PTMA area, including Treaty Creek access road and powerline route from Bell-Irving River. The site visit also included fly-over of Iron Cap Exploration adit area and ground visit in upper Mitchell Creek valley at the glacier terminus, Camp 9, and the KSM exploration camp at Sulphurets Lake to examine geochemical kinetic field tests and examples of drill core from ICEA adit area.

| ● | Ross D. Hammett (Ph.D., P.Eng.) of Golder visited the Property from September 12 to 15, 2017 and from August 8 to 10, 2010 and from October 18 and 19, 2011. The 2017 site visit involved a general inspection of the site conditions, with a specific focus on the stability of the mountain slopes where portals are proposed to be constructed, and all aspects of surface water management related to the proposed mining. It also involved visits to the drill rigs where core was being logged, and detailed inspection and qualitative geotechnical assessment of the core from many of the holes that had been drilled since 2012. The 2010 and 2011 site visits involved a general review of the site conditions for the Mitchell, Iron Cap, and Sulphurets deposits, and detailed inspection and qualitative geotechnical assessment of the core from many of the holes that had been drilled up to that time.

| ● | Brendon Masson (P.Eng.) of McElhanney visited the Property on September 13, 2012 and conducted an aerial reconnaissance over the existing and proposed site access road alignments.

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| 3.0 | Reliance on Other Experts

| 3.1 | 2016 Prefeasibility Study

Mr. Hassan Ghaffari, P.Eng. relied on:

| ● | Seabridge management concerning private royalties applicable to the PFS and applied in Section 22.0.

| ● | Lilburn & Associates LLC (Lilburn) concerning tax matters relevant to this PFS and detailed in Section 22.0. The reliance is based on a letter from Lilburn to Seabridge titled “Assistance with the tax portion of the economic analysis prepared by Tetra Tech Canada Inc. (“Tetra Tech”), dated September 13, 2016, in connection with the 2016 KSM (Kerr-Sulphurets-Mitchell) PFS, NI 43-101 Technical Report.

| ● | PricewaterhouseCoopers (PwC) concerning tax matters relevant to the PFS and detailed in Section 22.0. This reliance is based on a letter from PwC to Seabridge titled “Assistance with the review of income and mining tax portions of the economic analysis prepared by Seabridge Gold Inc. (“SGI”), dated September 14, 2016, in connection with the 2016 KSM (Kerr-Sulphurets-Mitchell) PFS, NI 43-101 Technical Report.

Dr. John Huang, Ph.D., P.Eng. relied on:

| ● | Mr. Neil Seldon of Neil S. Seldon & Associates Ltd. (NSA) for matters relating to the smelting terms, refining terms, saleability, and sales terms for copper concentrate and molybdenite concentrate. These terms are summarized in Section 19.0 and applied in Section 22.0.

Mr. William E. Threlkeld, P.Geo., PG relied on:

| ● | Mr. John Brassard, President of The Claim Group Inc., for matters relating to mineral and placer claims status and ownership. The reliance is based on a letter from TGC to Seabridge titled “Seabridge Gold Inc., Title Review-KSM Property, Province of British Columbia”, dated February 25, 2020. Mr. William E. Threlkeld, who is responsible for the information in Section 4.0, has relied entirely on the information provided by Mr. Brassard regarding the claims which comprise the KSM property, their ownership, and their status in Section 4.0.

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| 3.2 | 2020 Preliminary Economic Assessment

Ownership, Mineral Tenure, and Surface Rights

The Wood QPs have not independently reviewed ownership of the PEA area, including mineral and surface rights, and the underlying property agreements, including royalty agreements. The Wood QPs have fully relied upon, and disclaim responsibility for information provided by Seabridge through the following QP:

| ● | William E. Threlkeld, Senior Vice President of Seabridge, for matters relating to the mineral and surface rights, and the underlying property agreements, including royalty agreements.

This information is used in Section 24.22 of the Report.

Mr. William E. Threlkeld, P.Geo., PG relied on:

| ● | Mr. John Brassard, President of The Claim Group Inc., for matters relating to mineral and placer claims status and ownership. The reliance is based on a letter from TGC to Seabridge titled “Seabridge Gold Inc., Title Review-KSM Property, Province of British Columbia”, dated February 25, 2020.

Marketing Studies and Contracts

Dr. John Huang, Ph.D., P.Eng. relied on:

| ● | Mr. Neil Seldon of NSA for matters relating to the smelting terms, refining terms, saleability, and sales terms for copper concentrate and molybdenite concentrate. These terms are summarized in Section 24.19 and applied in Section 24.22.

Reclamation Bonds and Special Use Permit Costs

The Wood QPs have not reviewed the reclamation bonds and special use permits and have fully relied upon and disclaim responsibility for information provided by Seabridge through the following document.

| ● | Electronic mail from Mike Skurski, Vice President of Seabridge, titled “Updated reclamation costs” dated March 27, 2020.

This information is used in Section 24.22 of the Report.

Taxes

The Wood QPs have not independently reviewed the taxation information. The Wood QPs have fully relied upon, and disclaim responsibility for, taxation information derived from experts retained by Seabridge contained in the following document:

A letter authored by PwC with the title:

“NI 43-101 Technical Report Prepared for Seabridge Gold Inc. – Taxation Narrative”, dated April 24, 2020.

PwC is a limited liability partnership, which is a member firm of PriceWaterhouseCoopers International Limited, each member firm of which is a separate legal entity.

This information is used in Section 24.22 of the Report.

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| 4.0 | Property Description and Location

The KSM Property is in northwest BC at an approximate latitude of 56.50 N and a longitude of 130.30 W. The Mineral Resources that are the subject of this report are located relative to the North American Datum (NAD)83 Universal Transverse Mercator (UTM) coordinate system. The Property is approximately 950 km northwest of Vancouver, 65 km north-northwest of Stewart, and 21 km south-southeast of the Eskay Creek Mine (production ceased in 2009). Figure 1.2 is a general location map.

The KSM Property comprises four discrete claim blocks and a group of placer claims. Placer claims are on only part of the westernmost claim block of the KSM Property. Claim blocks of the KSM Property are referred to as:

| ● | the KSM claims

| ● | the Seabee claims

| ● | the Tina claims

| ● | the Treaty Creek Switching Station claims

The four KSM claim blocks include 79 mineral claims (cell and legacy) and 2 mining leases with a combined area of 40,784.97 ha. There are also 17 KSM placer claims held by KSM Mining ULC covering part of the KSM claims. The placer claims secure rights in a historically designated placer district. The Claim Group Inc. (TCG) acts as Agent on behalf of Seabridge with respect to maintaining all pertinent records associated with the KSM Property tenures. All claims and leases are in good standing under the Mining Tenure Act of BC and are recorded as owned 100% by KSM Mining ULC, a wholly owned subsidiary of Seabridge.

The Seabee and Tina claim blocks are located about 19 km northeast of the Kerr-Sulphurets-Mitchell-Iron Cap mineralized zones. These claim blocks are currently being considered for proposed infrastructure siting. The Treaty Creek Switching Station claims, adjacent to the NTL and east of the Seabee claims, are also being considered for power infrastructure siting.

The initial group of KSM mineral claims were purchased by Seabridge from Placer Dome in 2001. The mineral claims were converted from legacy claims to BC’s new Mineral Titles Online (MTO) system in 2005. In the MTO system, claims are located digitally using a fixed grid on lines of latitude and longitude with cells measuring 15 seconds north-south and 22.5 seconds east-west (approximately 460 m by 380 m). The legacy claims were located by previous owners by placing tagged posts along the boundaries; however, the survey method employed in locating the legacy claims is not known. The MTO system, where no markings are required on the ground, eliminates the potential for gaps and/or overlapping claims inherent in the old system.

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There is no record or evidence of any significant historical mining on the Property other than minimal placer mining. There are no known environmental liabilities to which the Property is subject apart from current exploration activity that is fully permitted. The BC Mineral Inventory (Minfile) contains 25 mineral occurrences associated with previous work in this area (mostly copper and gold).

The original KSM claim group consisted of two contiguous claim blocks known as the Kerr and Sulphurets (or Sulphside) properties. The claims are 100% owned by KSM Mining ULC, a wholly-owned subsidiary of Seabridge. Placer Dome retained a 1% net smelter returns royalty that is capped at Cdn$4.5 million, which is now owned by Newmont. Two of the pre-converted claims (Xray 2 and 6) are subject to a 2% net smelter returns royalty obligation capped at US$650,000, and Seabridge has been paying, annually, minimum advance royalty payments that may be credited against future royalties. The lands covered by these claims are now contained within mining lease No. 1031440.

Since it acquired the original KSM claim group, Seabridge has expanded its property holdings around the original KSM claims through staking and purchase of several claim groups. These groups include the Seabee group, acquired by staking, the Tina and BJ groups purchased in 2009, the New BJ group purchased in 2011, and the Treaty Creek Switching Station claims purchased in 2018. The Seabee, Tina, and Treaty Creek Switching Station claims are together referred to as the Seabee Property, and the original KSM group, BJ, and New BJ groups are referred to as the KSM Property (Figure 4.1). In 2014, most of the original KSM claims and the BJ claims acquired in 2009 were converted into two mining leases. The 17 KSM placer claims are shown in Figure 4.2.

Annual holding costs for all leases and claims vary by year depending on whether the fees are paid in cash or the value of work completed on developing the claims is used in lieu of a cash payment. Over the next five years, the annual cash holding costs to keep the claims and leases valid range between Cdn$450,000 to Cdn$970,000. Those estimated costs can be reduced significantly if work expenditures are applied in lieu of cash fees. No additional permits are required to address the recommendations in this report; part of the expenditures for that work can be applied in lieu of cash fees. Seabridge believes they have addressed all issues to secure access, mineral title, and ability to perform work on the property and are not aware of any risks, other than those identified in this Report, that could materially affect proposed work plans.

Mr. Threlkeld has relied on information with respect to all mining claim matters as provided by TCG in a letter titled “Title Review – KSM Property”, by John Brassard, dated February 25, 2020

Seabridge’s mineral claim blocks including the KSM, Seabee, Tina, and Treaty Creek Switching Station groups are shown in Table 4.1 and Table 4.3. The location of the four mineralized zones (Kerr, Sulphurets, Mitchell, and Iron Cap) is depicted in the southwestern portion of Figure 4.1. Table 4.2 shows Seabridge’s placer claims.

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Figure 4.1 KSM Mineral Claim Map

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Figure 4.2 KSM Placer Claim Map

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Table 4.1 KSM Mineral Claims and Leases

Claim No. | Claim Name | Area (ha) | Good To Date | Map Number
Mineral Claims (20)
394780 | BJ 5 | 100.0000 | 25-Nov-2024 | 104B059
394781 | BJ 6 | 100.0000 | 25-Nov-2024 | 104B059
394786 | BJ 11 | 500.0000 | 25-Nov-2024 | 104B059
394787 | BJ 12 | 500.0000 | 25-Nov-2024 | 104B059
394788 | BJ 13 | 100.0000 | 25-Nov-2024 | 104B059
394789 | BJ 13A | 25.0000 | 25-Nov-2024 | 104B059
394790 | BJ 14 | 100.0000 | 25-Nov-2024 | 104B059
394791 | BJ 15 | 250.0000 | 25-Nov-2024 | 104B059
394794 | BJ 18 | 300.0000 | 25-Nov-2024 | 104B059
394808 | BJ 31A | 375.0000 | 25-Nov-2024 | 104B049
394809 | BJ 32 | 150.0000 | 25-Nov-2024 | 104B049
394810 | BJ 33 | 450.0000 | 25-Nov-2024 | 104B049
394811 | BJ 34 | 150.0000 | 25-Nov-2024 | 104B049
394812 | BJ 35 | 450.0000 | 25-Nov-2024 | 104B049
683463 | - | 1,246.5185 | 25-Nov-2024 | 104B
683483 | - | 837.5991 | 25-Nov-2024 | 104B
705591 | BJ GAP1 | 231.6166 | 05-Feb-2024 | 104B
705592 | BJ GAP2 | 160.4624 | 05-Feb-2024 | 104B
1036269 | KSM 1 | 53.4100 | 21-May-2020 | 104B
1036270 | KSM 2 | 17.8100 | 21-May-2020 | 104B
Totals | - | 6,097.4166 | - | -
Mineral Leases (2)
1031440 | - | 6,085.0000 | 06-Oct-2020 | -
1031441 | - | 5,162.0000 | 06-Oct-2020 | -
Totals | - | 11,247.0000 | - | -

Seabridge Gold Inc. | 4-5 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Table 4.2 KSM Placer Claims

Claim No. | Claim Name | Area (ha) | Good to Date | Map Number
Placer Claims (17)
986922 | PL21 | 35.7243 | 16-May-2020 | 104B
986924 | PL22 | 35.7244 | 16-May-2020 | 104B
986925 | PL23 | 107.1670 | 16-May-2020 | 104B
541785 | Mitchell Placer 3 | 178.6000 | 10-Mar-2022 | 104B
542065 | Mitchell Sulphurets Junct | 71.4400 | 10-Mar-2022 | 104B
543053 | Add-On Placer 2 | 17.8600 | 10-Mar-2022 | 104B
543054 | Add-On Placer 3 | 17.8500 | 10-Mar-2022 | 104B
558630 | South Sulphurets Placer | 71.4600 | 10-Mar-2022 | 104B
575058 | Mitchell Placer 1 & 2 | 499.7600 | 10-Mar-2022 | 104B
577710 | 2008 Add-On Placer | 17.8600 | 10-Mar-2022 | 104B
577712 | 2008 Add-On Placer 2 | 17.8600 | 10-Mar-2022 | 104B
577715 | 2008 Add-On Placer 3 | 17.8600 | 10-Mar-2022 | 104B
577716 | 2008 Add-On Placer 4 | 17.8600 | 10-Mar-2022 | 104B
578100 | Sulphuret Creek Placer 2 | 17.8600 | 10-Mar-2022 | 104B
578154 | Lower Sulphuret Creek | 285.8200 | 10-Mar-2022 | 104B
578160 | Lower Sulphuret Creek | 17.8600 | 10-Mar-2022 | 104B
583353 | Unnamed & Add-On Placer 1 | 124.9900 | 10-Mar-2022 | 104B
Totals | - | 1,553.5557 | - | -

Seabridge Gold Inc. | 4-6 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Table 4.3 Seabee/Tina KSM Claims

Claim No. | Claim Name | Area (ha) | Good To Date | Map Number
SEABEE Property Mineral Claims (46)
566467 | BRIDGE1 | 445.8258 | 08-Apr-2022 | 104A
566468 | BRIDGE2 | 445.5733 | 08-Apr-2022 | 104A
566469 | BRIDGE3 | 427.7919 | 08-Apr-2022 | 104A
566470 | BRIDGE4 | 427.9770 | 08-Apr-2022 | 104A
566471 | BRIDGE5 | 445.7336 | 08-Apr-2022 | 104A
566472 | BRIDGE6 | 445.5766 | 08-Apr-2022 | 104A
566473 | BRIDGE7 | 427.9217 | 08-Apr-2022 | 104A
566474 | BRIDGE8 | 427.7599 | 08-Apr-2022 | 104A
566475 | BRIDGE9 | 427.6131 | 08-Apr-2022 | 104A
566476 | BRIDGE10 | 445.5312 | 08-Apr-2022 | 104A
566477 | BRIDGE11 | 302.8823 | 08-Apr-2022 | 104A
566478 | BRIDGE12 | 427.4311 | 08-Apr-2022 | 104A
566479 | BRIDGE13 | 445.1533 | 08-Apr-2022 | 104A
566481 | BRIDGE14 | 445.0611 | 08-Apr-2022 | 104A
566482 | BRIDGE15 | 444.8427 | 08-Apr-2022 | 104A
566484 | BRIDGE16 | 444.5621 | 08-Apr-2022 | 104A
566485 | BRIDGE17 | 426.7283 | 08-Apr-2022 | 104A
566487 | BRIDGE18 | 444.7114 | 08-Apr-2022 | 104A
566488 | BRIDGE19 | 444.8346 | 08-Apr-2022 | 104A
566489 | BRIDGE20 | 444.9690 | 08-Apr-2022 | 104A
566490 | BRIDGE21 | 427.2642 | 08-Apr-2022 | 104A
566491 | BRIDGE22 | 445.1671 | 08-Apr-2022 | 104A
566492 | BRIDGE23 | 427.3078 | 08-Apr-2022 | 104A
566493 | BRIDGE24 | 427.9239 | 08-Apr-2022 | 104A
566494 | BRIDGE25 | 427.9246 | 08-Apr-2022 | 104A
566495 | BRIDGE26 | 444.8785 | 08-Apr-2022 | 104A
566496 | BRIDGE27 | 391.3145 | 08-Apr-2022 | 104B
566497 | BRIDGE28 | 444.4573 | 08-Apr-2022 | 104A
566567 | BRIDGE29 | 427.4572 | 08-Apr-2022 | 104A
571582 | SEABEE1 | 408.8286 | 08-Apr-2022 | 104A
571583 | SEABEE2 | 373.1368 | 08-Apr-2022 | 104A
571584 | SEABEE3 | 444.0680 | 08-Apr-2022 | 104A
571585 | SEABEE4 | 426.0832 | 08-Apr-2022 | 104A
571586 | SEABEE5 | 372.6392 | 08-Apr-2022 | 104A
571587 | SEABEE6 | 159.6419 | 08-Apr-2022 | 104A
573813 | SEABEE7 | 213.2634 | 08-Apr-2022 | 104A
575633 | SEA 1 | 445.1987 | 08-Apr-2022 | 104A
575635 | SEA 2 | 445.3012 | 08-Apr-2022 | 104A
575636 | SEA 3 | 445.4096 | 08-Apr-2022 | 104A
575638 | SEA 4 | 445.4484 | 08-Apr-2022 | 104A
|  |  | table continues…

Seabridge Gold Inc. | 4-7 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Claim No. | Claim Name | Area (ha) | Good To Date | Map Number

575639 | SEA 5 | 445.3365 | 08-Apr-2022 | 104A
575642 | SEA 6 | 445.0850 | 08-Apr-2022 | 104A
575643 | SEA 7 | 213.4398 | 08-Apr-2022 | 104A
575645 | SEA 8 | 427.0822 | 08-Apr-2022 | 104A
575646 | SEA 9 | 35.5980 | 08-Apr-2022 | 104A
603133 | SEABEE 8 | 426.5614 | 08-Apr-2022 | 104B
Totals | - | 18,674.2970 | - | -
TINA Property Mineral Claims (11)
401548 | TINA 1 | 500.0000 | 28-Apr-2023 | 104B070
401549 | TINA 2 | 500.0000 | 28-Apr-2023 | 104B070
401550 | TINA 3 | 500.0000 | 28-Apr-2023 | 104B070
401551 | TINA 4 | 500.0000 | 28-Apr-2023 | 104B070
401552 | TINA 5 | 500.0000 | 28-Apr-2023 | 104B070
401553 | TINA 6 | 250.0000 | 28-Apr-2023 | 104B070
603134 | SEABEE 9 | 53.3796 | 28-Feb-2021 | 104B
1044111 | - | 71.1200 | 14-Nov-2020 | 104B
1044114 | - | 106.7800 | 14-May-2020 | 104B
1044281 | - | 53.3700 | 23-Nov-2020 | 104B
1047928 | - | 17.7900 | 17-Nov-2020 | 104B
Totals | - | 3,052.4396 | - | -
Treaty Creek Switching Station Claims (2)
1064020 | - | 53.4100 | 20-Nov-2020 | 104A
1064024 | - | 106.8400 | 20-Nov-2020 | 104A
Totals | - | 160.2500 | - | -

Seabridge Gold Inc. | 4-8 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 5.0 | Accessibility, Climate, Infrastructure, Local Resources and Physiography

The following section was taken from RMI’s April 6, 2007 NI 43-101 report entitled “Mitchell Creek Technical Report, Northern British Columbia” (Lechner, 2007), remains largely unchanged, and has only been updated for consistency in abbreviations and grammar.

The Property lies in the rugged Coastal Mountains of northwestern BC, with elevations ranging from 520 m in Sulphurets Creek valley to over 2,300 m at the highest peaks. Valley glaciers fill the upper portions of the larger valleys from just below the tree line and upwards. The glaciers have been retreating for at least the last several decades. Aerial photos indicate the Mitchell Glacier has retreated more than 1 km laterally and perhaps several hundred metres vertically since 1991.

The Property is drained by Sulphurets and Mitchell Creek watersheds that empty into the Unuk River, which flows westward to the Pacific Ocean through Alaska. The tree line lies at about 1,240 masl, below which a mature forest of mostly hemlock and balsam fir abruptly develops. Fish are not known to inhabit the Sulphurets and Mitchell watersheds. Large wildlife such as deer, moose, and caribou are rare due to the rugged topography and restricted access; however, bears and mountain goats are common.

The climate is generally that of a temperate or northern coastal rainforest, with sub-arctic conditions at high elevations. Precipitation is high with annual rainfall and snowfall totals estimated to be somewhere between the historical averages for the Eskay Creek Mine and Stewart, BC. These range from 801 mm to 1,295 mm of rain and 572 cm to 1,098 cm of snow, respectively (data to 2005). The length of the snow-free season varies from about May through November at lower elevations and from July through September at higher elevations. Exploration activities have typically been carried out from late May into November. It is envisioned that operations would be conducted throughout the year with assets required for snow removal.

Access to the Property is via helicopter. Three staging areas for mobilizing crews and equipment were used. These are:

| 1. | An area located at kilometre 54 on the private Eskay Creek Mine Road, which is about 25 km to the north-northwest of the Property.

| 2. | Along the public Granduc Road, which is located about 35 km to the south-southeast of the Property, which in turn is about 40 km north of the town of Stewart, BC. A section of this road passes through Alaska and the town of Hyder. This area has not been utilized since 2011.

| 3. | The Bell II Lodge, on Highway 37, 40 km east-northeast.

Stewart, a town of approximately 400 inhabitants, is the closest population center to the Property. It is connected to the provincial highway system via paved, all weather Highway 37A. The larger population centers of Prince Rupert, Terrace, Kitimat, and Smithers, with a total population of about 36,000, are located approximately 270 km to the southeast.

Seabridge Gold Inc. | 5-1 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Multiple deep-water loading facilities for bulk shipping mineral concentrates and freight handling exist in Stewart, one of which is currently utilized by the Red Chris Mine. Historically, they have been used by several other mines in northern, BC. The nearest railway is the CNR Yellowhead route, which is located approximately 220 km southeast of the Property. This line runs east-west, and can deliver concentrate to deep-water ports near Prince Rupert and Vancouver, BC.

The Property lies on Crown land; therefore, all surface and access rights are granted by the Mineral Tenure Act, the Mining Right of Way Act, and the Mining Rights Amendment Act. There are no settlements or privately owned land in this area; there is limited commercial recreational activity in the form of helicopter skiing and guided fishing adventures. The closest power transmission lines run along the Highway 37, 40 km east of the Property, and along the 37A corridor to Stewart, approximately 50 km southeast.

Pretium is currently mining their high-grade underground Brucejack deposit that is located about 5 km east of the Property. The Brucejack mine site is accessible via a year-round, all weather road branching from Highway 37.

Additional details on availability of power, water, waste rock and tailings storage areas and surface rights to support mine operations are discussed in the relevant sections of this Report.

Seabridge Gold Inc. | 5-2 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 6.0 | History

| 6.1 | Exploration History

The modern exploration history of the Property began in the 1960s, with brief programs conducted by Newmont Exploration of Canada Ltd. (Newmont), Granduc, Phelps Dodge, and the Meridian Syndicate. All of these programs were focused towards gold exploration. Various explorers were attracted to this area due to the numerous large, prominent pyritic gossans that are exposed in alpine areas. There is evidence that prospectors were active in the area prior to 1935. Several short hole, reconnaissance level drilling programs were undertaken between 1969 and 1991. The Sulphurets Zone was first drilled by Granduc Mines in 1968, Kerr by Brinco Ltd. in 1985, Mitchell Creek by Newhawk Gold Mines Ltd. (Newhawk) in 1991, and Iron Cap by Esso Minerals in 1980.

In 1989, Placer Dome (Placer) acquired a 100% interest in the Kerr deposit from Western Canadian Mines; in the following year, they acquired the adjacent Sulphurets Property from Newhawk. The Sulphurets Property also hosts the Mitchell Creek deposit and other mineral occurrences. In 2000, Seabridge acquired a 100% interest from Placer in both the Kerr and Sulphurets properties, subject to capped royalties.

There is no recorded mineral production, nor evidence of it, from the Property. Immediately west of the Property, small-scale placer gold mining has occurred in the Sulphurets and Mitchell creeks.

Table 6.1 summarizes the more recent exploration history of the Property.

Seabridge Gold Inc. | 6-1 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Table 6.1 Exploration Summary of the Kerr Zone

Year | Activity
1982-1883 | “Alpha JV” began prospecting and soil geochem surveys of the Kerr gossan focusing on gold
1984-1985 | Brinco optioned the Kerr project, completed some geologic surveys and drilled 3 holes
1987-1989 | Western Canadian Mines optioned Kerr and completed 59 drill holes and recognized Cu-Au porphyry
1989 | Placer acquires Kerr property
1990-1992 | Placer began delineation drilling of Kerr deposit at 50 m centers by drilling 83 holes totalling 16,414 m
1992-1996 | Placer estimated resources (pre NI 43-101), met testwork, and scoping studies
1996-2000 | Project was dormant
2000 | Seabridge acquired a 100% interest in Kerr from Placer Dome
2002 | Noranda Inc. acquired an option from Seabridge with the right to earn up to a 65% interest in Kerr
2003-2004 | Noranda Inc. undertook various exploration surveys
2006 | Seabridge purchased Falconbridge (formerly Noranda) option
2009 | Seabridge drilled 7 holes totaling about 1,159 m, conducted metallurgical testing, and permit work
2010 | Seabridge drilled 4 holes totaling about 1,453 m, conducted metallurgical testing, and permit work
2011 | Seabridge drilled 9 resource definition holes totaling about 2,631 m, continued with preliminary feasibility studies
2012 | Seabridge drilled 5 exploration holes totalling 3,731 m
2013 | Seabridge drilled 29 resource definition holes totalling 23,844 m, completed induced polarization and down-hole geophysical surveys
2014 | Seabridge drilled 16 resource definition holes totalling 15,909 m, completed magneto-telluric and gravity surveys
2015 | Seabridge drilled 5 resource definition holes and 3 geotechnical holes totalling 6,629 m
2016 | Seabridge drilled 5 resource definition holes totalling 7,119 m
2017 | No exploration
2018 | One hole drilled totalling 501 m
2019 | Completed airborne ZTEM geophysical survey

Table 6.2 summarizes the exploration history of the Sulphurets, Mitchell, and Iron Cap zones.

Seabridge Gold Inc. | 6-2 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Table 6.2 Exploration Summary of the Sulphurets, Mitchell, Iron Cap Zones, and other Exploration Targets

Year | Activity
1880-1933 | Limited placer gold exploration and mining
1935-1959 | Placer gold prospecting and staking of mining claims
1959-1960 | Newmont and Granduc conducted surveys including airborne mag. Sulphurets and Iron Cap Au zones discovered. D. Ross, S. Bishop and W. Dawson prospected and stake claims in area
1961-1968 | Granduc Mines conducted geologic/geochem surveys, drilled 9 holes into Sulphurets zone. Ross-Bishop-Dawson claims were optioned by Phelps Dodge in 1962, Meridian Syndicate in 1965, and Granduc in 1968
1963 | R. Kirkham completed a M.Sc. thesis on the geology of Mitchell and Sulphurets areas
1981 | T. Simpson completed a M.Sc. thesis on the geology of the Sulphurets gold zone
1971-1977 | Granduc Mines conducted additional exploration surveys targeting molybdenum and drilled 6 holes into Snowfield zone (Bruceside)
1979-1984 | Esso Minerals optioned Sulphurets property and completed early stage exploration including drilling 14 holes (2,275 m)
1985-1991 | Granduc optioned Sulphurets to Lacana (later Corona) and Newhawk. Lacana-Newhawk JV spends about Cdn$21M developing West Zone and other smaller precious metal veins on Bruceside property. Drilled 11 holes at Sulphurets. Homestake undertook exploration after acquiring Corona
1991 | Arbee prospect was optioned by Newhawk from D. Ross
1992 | Arbee prospect was optioned by Placer from Newhawk
1991-1992 | Newhawk commissioned AB geophysical survey over Sulphurets. Newhawk subdivided Suphurets property into Sulphside and Bruceside. Placer acquired Sulphside (Sulphurets, Mitchell, Iron Cap, and other prospects)
1992 | Placer undertook delineation drilling of Sulphurets deposit at 50 meter centers (23 holes)
1993 | J. Margolis completed a PhD thesis on the Sulphurets district. Newhawk-Corona drilled 3 holes in the Snowfields and Josephine zones east of Sulphurets
1992-1996 | Placer completed geologic modeling, resource estimation (pre NI 43-101), preliminary met testwork, and scoping studies
1999 | Silver Standard Resources acquired Newhawk.
1996-2000 | Sulphurets project was dormant
2000 | Seabridge acquired a 100% interest in the Sulphurets/Mitchell properties from Placer.
2002 | Noranda Inc. acquired an option to earn up to 65% from Seabridge
2003-2004 | Noranda Inc. undertook various exploration surveys
2005 | Falconbridge Ltd. (formerly Noranda) completed 4,092 m of diamond drilling in 16 holes
2006 | Seabridge purchased Falconbridge’s option and drilled 29 holes totaling about 9,129 m at the Sulphurets and Mitchell zones
2007 | Seabridge purchased Arbee prospect from D. Ross and drilled 37 holes totaling 15,650 m
table continues…

Seabridge Gold Inc. | 6-3 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Year | Activity
2008 | Seabridge purchased Arbee prospect from D. Ross, drilled 40 holes totaling 17,328 m, started metallurgical testing, obtained new topographic data, and initiated permit related activities
2009 | Seabridge drilled 44 holes totalling 11,844 m (resource definition, geotechnical and water monitoring), conducted metallurgical testing, and intensified permit data collection
2010 | Seabridge drilled 86 holes totaling 26,756 m (resource definition and geotechnical), conducted metallurgical testing, and intensified permit data collection
2011 | Seabridge drilled 54 resource definition holes totalling 18,087 m, continued prefeasibility work, and completed magneto-telluring geophysical surveys
2012 | Seabridge drilled 40 holes totalling 18,015 m including 15 holes from other nearby targets (e.g. Camp Zone and Icefield)
2013 | Seabridge drilled 11 holes totalling 8,857 m and completed induced polarization and down-hole geophysical surveys
2014 | Seabridge drilled 13 holes totalling 13,605 m and completed magneto-telluring and gravity geophysical surveys
2015 | Seabridge drilled 3 holes totalling 4,395 m and completed airborne magnetic geophysical surveys
2016 | Seabridge drilled one resource definition hole totalling 1,038 m
2017 | Seabridge drilled 12 resource definition holes totalling 11,077 m
2018 | Seabridge drilled 42 holes totalling 25,848 m
2019 | Seabridge drilled 26 resource definition holes totalling 6,121 m and completed airborne ZTEM and 3D induced polarization surveys

| 6.2 | Historical Resource Estimates

RMI is unaware of any publicly disclosed historical Mineral Resource estimates for the KSM deposits prior to Seabridge’s entry into the district. RMI has prepared Mineral Resources consistent with the standards in NI 43-101 for the Kerr, Sulphurets, Mitchell, and Iron Cap zones (Lechner, 2007; Lechner 2008a; Lechner, 2008b; Lechner, 2009; Lechner, 2010; Lechner, 2011; Lechner, 2014: Lechner, 2016). Those Mineral Resources have been replaced and updated with this Report as discussed in Section 14.0.

| 6.3 | History of Production

There is no known production from the Kerr, Sulphurets, Mitchell, or Iron Cap deposits.

Seabridge Gold Inc. | 6-4 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 7.0 | GEOLOGICAL SETTING AND MINERALIZATION

| 7.1 | Geological Setting

The KSM porphyry Au-Cu-Mo district is located in northwestern British Columbia, within the Intermontane belt of the Canadian Cordillera. Encompassing portions of British Columbia, the Yukon, and Alaska, the Canadian Cordillera is a highly fertile metallogenetic domain that features a suite of ore bodies ranging in age from 1.6 billion to less than 20 million years old (Nelson and Colpron, 2007). The district lies within “Stikinia”, a long-lived volcanic island-arc terrane that was accreted onto the Paleozoic basement of the North American continental margin in the Middle Jurassic. Stikinia is the largest of several fault bounded, allochthonous terranes within the Intermontane belt, which lies between the post-accretionary. Tertiary intrusives of the Coast belt and continental margin sedimentary prisms of the Foreland (Rocky Mountain) belt. In the KSM area, Stikinia is dominated by variably deformed, oceanic island arc complexes of the Triassic Stuhini and Jurassic Hazelton groups (Alldrick and Britton, 1988; Kirkham and Margolis, 1995). Back-arc basins formed eastward of the KSM Property in the Late Jurassic and Cretaceous were filled with thick accumulations of fine black clastic sediments of the Bowser Group. Folding and thrusting due to sinistral transpression in the mid-Cretaceous, related to the development of the Skeena fold and thrust belt, followed by extensional conditions generated the area’s current structural features (Febbo et al., 2015). The mid-Cretaceous deformation generated two of the most important structures in the district: the north-northeast striking, moderately west-northwest dipping Sulphurets thrust fault (STF) and Mitchell thrust fault (MTF). Remnants of Quaternary basaltic eruptions occur throughout the region.

Early Jurassic sub-volcanic intrusive complexes are common in the Stikinia terrane, and several host well-known precious- and base-metal-rich hydrothermal systems. These include copper-gold porphyry zones such as Galore Creek, Red Chris, Kemess, Mt. Milligan, and KSM. In addition, there are a number of related polymetallic zones including skarns at Premier, epithermal veins, subaqueous vein, and replacement sulphide zones at Eskay Creek, Snip, Brucejack, and Granduc. At KSM, Triassic rocks include marine sediments and intermediate volcanics of the Stuhini Group. The lowermost Stuhini Group is dominated by turbiditic argillite and sandstone, which are overlain by volcanic pillowed flows and breccias. The upper portion consists of turbidites and graded sandstones similar to the base strata. The Stuhini Group is separated by an erosional unconformity from the overlying Jurassic sediments and volcanics of the Jack Formation and Hazelton Group. The Jack Formation is comprised of fossiliferous, limey sediments, mudstones, and sandstones. The base is marked by a granodiorite and limestone cobble bearing conglomerate. Overlying the Jack Formation is the Hazelton Group, dominated by andesitic flows and breccias zoned in a volcanic chain with high paleotopographic relief. Distinct felsic welded tuff horizons of the Mount Dilworth Formation are an important stratigraphic marker in the Hazelton Group, as they are closely associated with the Eskay Creek Zone.

A variety of dykes, sills, and plugs of diorite, monzodiorite, granodiorite, monzonite, syenite, and granite are found in the area. U-Pb zircon radiometric dating indicates that the vast majority of the intrusions in the district are of Early Jurassic age (~190 – 197 ± 2 Ma; Kirkham and Margolis, 1995; MacDonald, 1993; Bridge, 1993; Febbo et al., 2019a), coinciding with the age of ore formation. Small volumes of thin post-mineral dykes, including mid-Cretaceous diabase dykes and Eocene lamprophyre dykes, are also observed within the Property. The Early Jurassic intrusions in the district are categorized into two regional suites: the calc-alkaline Texas Creek suite and the alkaline Premier suite. Texas Creek suite are intimately related to porphyry Au-Cu-Mo ore formation at the Kerr, Sulphurets, Mitchell and Iron Cap deposits, and are locally referred to as the “Sulphurets suite intrusions.” The Premier suite intrusions are most abundant in the hanging wall of the STF in the district, and comprise sills and plugs of typically coarse-grained porphyritic monzonite to syenite. Figure 7.1 is a generalized geologic map of the KSM district showing lithology, major structures, and mineralized zones. Drill hole locations are shown for all KSM deposits in Section 10.

Seabridge Gold Inc. | 7-1 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Figure 7.1 Geology of the KSM District

Source: (Seabridge, 2019)

Seabridge Gold Inc. | 7-2 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 7.2 | Mineralization

| 7.2.1 | Kerr Zone

The Kerr Zone is centered on an Early Jurassic, north-south trending, steep westerly dipping, tabular intrusive complex. Drilling demonstrates that the Kerr Zone has an extent of 2,400 m along strike, a width of roughly 800 m, and a vertical extent of at least 2,200 m. The flattened and elongated modern morphology of Kerr is highly unusual for porphyry copper-gold deposits, which typically display roughly cylindrical morphologies. The flattened morphology and relatively elevated copper:gold ratio of the Kerr Zone distinguishes Kerr from the other deposits in the district. The surface expression of the Kerr deposit is a large and elongated, northerly trending, pyritic gossan, primarily exposed in a cirque on the northern flank of Kerr peak.

The Kerr intrusive complex is composed of a suite of northerly-striking and steeply west-dipping dykes and intrusions emplaced into a sequence of rhythmically bedded siltstones, sandstones, conglomerates, and debris flows belonging to the Lower Jurassic Hazelton Group. Wall rocks adjacent to the intrusions have been hornfelsed and hydrothermally altered, but generally contain marginal metal grades. The complex is composed of an east and west limb separated by a thin wedge of intensely altered sedimentary wall rock. The west limb is up to 500 m thick, and the east limb is up to 300 m thick.

Mineralization at Kerr was associated with the sequential emplacement of numerous syn-mineral plagioclase-hornblende-(biotite-K-feldspar-apatite)-phyric diorite to monzodiorite dykes, which display a porphyritic texture with 30-40 vol.% phenocrysts up to 5 mm in length. Dykes are typically several meters to several tens of meters wide. Three distinct phases of these intrusions are recognized based upon typical metal content, quartz vein abundance, degree of hydrothermal alteration, and presence of xenoliths of earlier phases. The earliest syn-mineral phase (“P1”) is characterized by strong gold and copper mineralization and a high abundance (50-90 vol.%) of quartz veins. The second phase of syn-mineral diorite intrusions (“P2”) make up the bulk of the Kerr intrusive complex and feature variable mineralization and quartz vein abundance. The P2 intrusions envelop and truncate portions of the P1 intrusions, and are observed to contain occasional P1 xenoliths. Finally, late syn-mineral to post-mineral dykes (“P3”), with weak to no mineralization, cut the earlier phases. Volumetrically minor post-mineral dkes, with true thicknesses typically on the order of a meter, are also observed at the Kerr deposit. These include a set of unmineralized, K-feldspar megacrystic, plagioclase-hornblende-phyric porphyritic dykes, aphanitic diabase dykes, and amygdaloidal lamprophyre dykes. Current geochronological data for the Kerr deposit includes a U-Pb zircon age of 197 ± 3 Ma for a syn-mineral “syenodiorite” intrusion, as well as a slightly younger U-Pb zircon age of 195 ± 1.5 Ma for a late syn-mineral K-feldspar megacrystic porphyry dyke (Bridge, 1993).

Small volumes of syn-mineral hydrothermal breccias are also found at Kerr, with subangular to angular clasts of local intrusions or wall rock and rock flour matrices. Hydrothermal breccias with abundant anhydrite and quartz vein fragments (“QABX”) are also observed; these QABX zones are commonly overprinted with strong sericitic alteration and a high-sulphidation bornite – pyrite ± enargite mineral assemblage, and can contain strong copper mineralization. The “Bornite Breccia” zone at Mitchell is analogous to the QABX zones at Kerr.

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Hydrothermal alteration at the Kerr deposit is consistent with that described for many calc-alkaline porphyry deposits, with zones of propylitic, potassic, sericite-chlorite, and sericitic alteration, as well as small zones of remnant advanced argillic alteration.

Post-mineral deformation and faulting, mainly related to the mid-Cretaceous development of the Skeena fold and thrust belt, has significantly impacted the Kerr deposit. Efforts to thoroughly document and reconstruct the geology of the Kerr deposit are hindered by intense hydrothermal alteration and extensive post-mineralization deformation. Large portions of Kerr are affected by texturally destructive sericitic alteration that contains a pervasive foliation, rendering protolith identification difficult, and the occurrence of numerous post-mineral faults further obscure the original spatial relationships within the deposit.

Pervasive foliation and vein deformation resulting from mid-Cretaceous deformation is observed throughout the Kerr deposit. However, as noted at the Mitchell deposit (e.g., Febbo et al., 2015), foliation is strongest in rocks characterized by alteration assemblages with low rock strengths, such as zones with strong sericitic alteration. East-west shortening related to mid-Cretaceous deformation likely contributed to the unusually flattened morphology of the Kerr deposit. Furthermore, the strongest sericitic alteration occurs at the northern end of Kerr, at a lower elevation than the south side of Kerr. Sericitic alteration is typically formed at relatively shallow depths in porphyry copper systems, atop deep potassic alteration. The current emplacement of alteration assemblages suggests tilting or complex structural modification of the system.

The dominant copper mineral is chalcopyrite, which typically occurs as isolated grains about 0.2 mm to 2 mm across, disseminated and clustered in quartz veins, fractures, and surrounding haloes. Bornite is present almost exclusively in the north half of the east leg, within a QABX zone containing >50% crackled quartz veins, and is accompanied by coarse grained chalcopyrite and minor tennantite. Tennantite-tetrahedrite is rare, but widely distributed in late quartz-carbonate veins, mostly in wall rocks, along with minor sphalerite, rare galena, and arsenopyrite. Dark, arseniferous pyrite is associated with these minerals. Molybdenite is a very minor constituent, and Kerr contains significantly lower overall Mo grades than the other deposits in the district. Visible gold has not been observed except under microscopic examinations, where it is observed as less than 100 µm inclusions within sulphides, mainly chalcopyrite, and sulphide grain boundaries. High-grade mineralization at Kerr is primarily associated with early potassic alteration and “A-type” and “B-type” quartz veins (using the terminology of Gustafson and Hunt, 1975). At shallower levels in the system, both copper and gold grades are highly correlated with the density of A- and B-type veining. At depth, however, mineralization does not correlate as strongly with quartz vein density, and is of a more disseminated style.

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A lithological map of the Kerr deposit is shown in Figure 7.2. Representative cross sections through the south and the north of the Kerr deposit, showing lithological units, quartz vein abundance, gold grade, and copper grade, are shown in Figure 7.3 and Figure 7.4, respectively.

Figure 7.2 Geological Map of the Kerr Deposit. Thin dykes are not displayed.

Source: (Seabridge, 2019)

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Figure 7.3 Four panels of the same vertical E-W section through the Kerr deposit, at 6,258,650N, showing different properties of the deposit.

Source: (Seabridge, 2019)

Notes: Location of section is shown on Figure 7.3: a) Simplified lithology, with important intrusive phases (P1, P2, P3, P4, P6), hydrothermal breccia with anhydrite and quartz vein clasts (QABX), hydrothermal breccias alongside P2 intrusions (IBX), and undifferentiated wall rock (WR) shown. Drill hole traces, within ±100 m of the section, are displayed as black lines. Units are described in the text. The intrusion labels refer to early intrusions with high quartz vein densities (50-90 vol.%; P1), syn-mineral plagioclase-hornblende-phyric intrusions (P2), post-mineral plagioclase-hornblende-phyric dykes (P3), post-mineral K-feldspar megacrystic dykes (P4), and post-mineral lamprophyre dykes (P6). b) Isolines of total volume percent quartz vein abundance as logged in drill core. c) Gold grade, from the Seabridge Gold EOY 2016 block model of Kerr. d) Copper grade, from the EOY 2016 block model of Kerr.

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Figure 7.4 Four panels of the same vertical E-W section through the Kerr deposit, at 6,259,650N, showing different properties of the deposit.

Source: (Seabridge, 2019)

Notes: Location of section is shown on Figure 7.4: a) Simplified lithology, with important intrusive phases (P1, P2, P3, P4, P6), hydrothermal breccia with anhydrite and quartz vein clasts (QABX), and undifferentiated wall rock (WR) shown. Drill hole traces, within ±100 m of the section, are displayed as black lines. Units are described in the text. The intrusion labels refer to early intrusions with high quartz vein densities (50-90 vol.%; P1), syn-mineral plagioclase-hornblende-phyric intrusions (P2), post-mineral plagioclase-hornblende-phyric dykes (P3), post-mineral K-feldspar megacrystic dykes (P4), and post-mineral lamprophyre dykes (P6). b) Isolines of total volume percent quartz vein abundance as logged in drill core. c) Gold grade, from the Seabridge Gold EOY 2016 block model of Kerr. d) Copper grade, from the EOY 2016 block model of Kerr.

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| 7.2.2 | Sulphurets Zone

The Sulphurets Zone is situated between the Kerr and Mitchell deposits, immediately to the north of the valley hosting Sulphurets Lake. Sulphurets was historically subdivided into several adjacent mineralized zones, including the Raewyn Copper-Gold, Breccia Gold, Main Copper, and Canyon zones. As defined by the resource cut-off (NSR Cdn$9.00), the main body of the Sulphurets deposit has a lensoidal geometry, dipping approximately 30 degrees northwest. It has a horizontal extent of 2,200 m, down dip extent of 550 m, and true thickness of up to 330 m. Copper and gold grades gradually diminish towards the limits of the known resource area, but the extent of anomalous copper and gold grades has not been completely delineated down dip or in the southwest direction.

Sulphurets is a structurally complex deposit, intersected by numerous east-vergent faults associated with the mid-Cretaceous Skeena fold and thrust belt. The deposit is composed of stacked thrust fault panels made up of Upper Triassic Stuhini Group and Lower Jurassic Hazelton Group volcano-sedimentary strata intruded by a number of dykes and stocks. The thrust faults and most dykes and intrusions within the Sulphurets Zone strike southwest-northeast and dip at shallow to moderate angles towards the northwest. Three main fault panels make up the deposit, each featuring a distinct set of intrusions and style of mineralization.

The uppermost fault panel (“Upper Panel”) is located in the hanging wall of the STF. The most voluminous plutonic rocks in the Upper Panel are Premier suite monzonites and syenites, including a large intrusive body, dipping at roughly 40° – 50° degrees towards the northwest, as well as several thinner dykes. The Premier suite intrusions are typically pinkish to reddish K-feldspar-plagioclase-phyric porphyries with abundant coarse, tabular K-feldspar phenocrysts, commonly displaying concentric zoning. A U-Pb zircon radiometric age of 191.8 +6.5/-1.0 Ma was obtained for a Premier “feldspar porphyry” from the Upper Panel (MacDonald, 1993). Most of the Premier intrusions in the Upper Panel display potassic alteration and reddish hematite dusting, but are nearly barren of sulphides. Hydrothermal breccias and strongly altered wall rocks peripheral to the intrusions host the bulk of mineralization in the Upper Panel. First discovered in the 1930s, this mineralized zone was originally referred to as the Main Copper Zone. Alteration in the breccias and wall rock ranges from a dark purplish-brown proximal potassic assemblage with hydrothermal K-feldspar, fine-grained biotite, magnetite, and a relatively high chalcopyrite:pyrite ratio, to a chlorite – pyrite-dominant alteration, to a distal chlorite – epidote – calcite ± magnetite ± hematite propylitic assemblage.

The Upper Panel structurally overlies the thin “Middle Panel,” which is bounded by the STF and the Raewyn fault (RF), a minor thrust fault that terminates against the STF on the northeastern side of the deposit. The Middle Panel features Texas Creek (“Sulphurets”) suite plagioclase-hornblende-phyric diorite intrusions and small volumes of hydrothermal breccias, and hosts very little of the mineralization at Sulphurets. Hydrothermal alteration assemblages are limited to chlorite-dominant alteration in the diorite intrusions, and to weak hornfels ± chlorite in the sedimentary wall rocks and hydrothermal breccias that occur along the margins of diorite intrusions.

Finally, the “Lower Panel,” which occurs beneath the Middle Panel, hosts the bulk of the mineralization at Sulphurets. The Lower Panel is thought to be bounded at depth by the MTF, though the MTF is only intercepted at depth in a handful of deep drill holes. This panel was historically divided into several discrete mineralized zones, including, from southwest to northeast: the Canyon zone, the Breccia Gold zone, and the Raewyn Copper-Gold zone (Fowler and Wells, 1995). The majority of the gold and copper mineralization in the Lower Panel is found in the Breccia Gold and Raewyn Copper-Gold zones. The Raewyn Copper-Gold zone is primarily hosted in crackled and hornfelsed sedimentary wall rock partially affected by distinctive dark brown, biotite-rich potassic alteration. It has an apparent north-northeast strike and dips about 45° to the northwest, with approximate true dimensions of 1,000 m in strike, 550 m down dip, and up to 250 m in thickness. The Raewyn Copper-Gold zone remains open down dip and along strike to the northeast at depth. The Breccia Gold zone has a pipe-like geometry, approximately 100 m in diameter, and plunging ~50° to the west-northwest. It is thought to be hydrothermal in original, and features tourmaline-rich alteration, strong gold grades, and abundant pyrite and tennantite, but relatively low copper grades. Sedimentary wall rock from the Lower Panel features silica addition as well as potassic alteration, as evidenced through whole rock geochemistry results comparing wall rock from the Lower Panel and the comparatively weakly altered sedimentary wall rock of the Middle Panel.

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The Lower Panel features two main types of Texas Creek suite intrusions: subporphyritic plagioclase-hornblende-phyric diorite intrusions, and porphyritic plagioclase-hornblende-K-feldspar-phyric monzodiorite dykes. The diorites typically feature 30-40 vol.% tabular plagioclase and hornblende phenocrysts, up to ~4 mm in length, in a fine-grained groundmass. Similar diorite intrusions are found at the Kerr, Mitchell, Snowfield, and Iron Cap deposits, where they are intimately associated with porphyry mineralization. However, the diorites at Sulphurets lack abundant porphyry style quartz vein stockworks, have lower pyrite contents, potassic and sericitic alteration facies are subordinate to chlorite-pyrite alteration, and their associated Cu-Au resources are much smaller. A significant proportion of the mineralization at Mitchell, Kerr, and Iron Cap is hosted within syn-mineral diorite intrusions, whereas at Sulphurets, mineralization typically occurs in the wall rocks immediately adjacent to, above and between the fingers of diorite.

Late, discontinuous veins, breccias, shear fillings, and patchy replacements similar to those in the Upper Panel occur throughout the Lower Panel fault block. A higher density of these late mineralized structures is reflected in average arsenic, lead, antimony and zinc concentrations being roughly two to three times those of the Upper Panel. Scattered, sub-meter to centimeter scale quartz and sulphide veins, breccia veins, shear fillings, patchy discontinuous clots or replacements and disseminations occur in all rock types except late dykes, but are most abundant in brecciated and crackled zones. They include coarse pyrite, carbonate, chlorite, quartz, occasionally with minor chalcopyrite, sphalerite, arsenopyrite or galena, and traces of tennantite. Native gold is rare but has been observed as sub-millimeter blebs and stringers in quartz-carbonate-sulphide veins, and chlorite-pyrite-carbonate breccia matrix fillings.

A small number of very thin, volumetrically insignificant post-mineral dykes with meter-scale widths cut all other lithologies at Sulphurets. These include northwesterly-dipping, fine-grained diabase dykes with aphanitic chilled margins and pervasive chlorite alteration, and two fine-grained, unaltered, black lamprophyre dykes dipping west-northwest. Similar post-mineral dyke sets are observed in the Kerr, Mitchell, and Iron Cap zones.

Brittle fracturing is widespread at Sulphurets, and multiple fault sets, ranging from shallow- to steeply-dipping, are found throughout the deposit. In addition to the east-vergent thrust faults, numerous small, north to northeasterly striking, steeply westerly dipping faults, fracture zones and shears cut through the district. Subvertical penetrative cleavage is found throughout much of the district, and is especially well developed in alteration facies rich in phyllosilicates and clays (Kirkham, 1963; Margolis, 1993; Febbo et al., 1995). At Sulphurets, the absence of significant rheologically weak sericitic alteration facies has resulted in the absence of the strong penetrative cleavage present at the Kerr, Mitchell, and Snowfield deposits. At Sulphurets, a northeasterly-striking, steeply dipping foliation is only observed in discrete areas with sericitic alteration (e.g., Fowler and Wells, 1995).

A geological map of the Sulphurets deposit, which also displays the extent of near-surface gold and copper mineralized zones, is found in Figure 7.5. Representative cross sections through the southwest and the centre of the Sulphurets deposit, showing lithological units, alteration assemblages, gold grade, copper grade, molybdenum grade, and silver grade are shown in Figure 7.6 and Figure 7.7, respectively.

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Figure 7.5 Map of the Sulphurets deposit

Source: (Seabridge, 2019)

Notes: a) Geology map, showing principal intrusions, dykes, and hydrothermal breccias. Labeled faults include the Sulphurets Thrust Fault (STF), Raewyn Fault (RF), and the approximate trace of the Mitchell Thrust Fault (MTF). b) Map of the principal gold and copper mineralized zones within the Sulphurets deposit, along with the historical names of these zones. The locations of the major fault panels discussed in the text – the Upper, Middle, and Lower Panels – are also indicated.

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Figure 7.6 Vertical cross-section through the Sulphurets deposit, looking ENE

Source: (Seabridge, 2019)

Notes: The traces of the Sulphurets Thrust Fault (STF), Raewyn Fault (RF), and Mitchell Thrust Fault (MTF) are shown. a) Simplified geology of the Sulphurets deposit, showing undifferentiated sedimentary wall rock (WR) and principal intrusions and breccias: diorite (DR), monzodiorite intrusions (MZD), hydrothermal breccia (BX), and diabase dyke (DB). See text for unit descriptions. The traces of drill holes occurring within ± 100 m of the section are shown as black lines. b) Simplified diagram of hydrothermal alteration zoning within the Sulphurets deposit, illustrating the extent of pervasive chlorite (CHL), propylitic (PRO), hornfels (HF), and hornfels with tourmaline (TO) alteration zones, as well as areas with quartz-sericite-pyrite overprint (SER). c) Ag grades from the Seabridge Gold EOY 2018 block model; d) Au grades from the EOY 2018 block model; e) Cu grades from the EOY 2018 block model; f) Mo grades from the EOY 2018 block model.

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Figure 7.7 Vertical cross-section through the Sulphurets deposit, looking NNE

Source: (Seabridge, 2019)

Notes: The traces of the Sulphurets Thrust Fault (STF), Raewyn Fault (RF), and Mitchell Thrust Fault (MTF) are shown. a) Simplified geology of the Sulphurets deposit, showing undifferentiated sedimentary wall rock (WR) and principal intrusions and breccias: diorite (DR), monzonite (MZ), hydrothermal breccia of Middle Panel (BX1), Sulphurets Gold Breccia (BX2), brecciated margin of monzonite intrusions (BX3), and lamprophyre dyke (LM). See text for unit descriptions. The traces of drill holes occurring within ±100 m of the section are shown as black lines. b) Simplified diagram of hydrothermal alteration zoning within the Sulphurets deposit, illustrating the extent of pervasive chlorite (CHL), propylitic (PRO), hornfels (HF), hornfels with tourmaline (TO), hornfels with biotite (BI), potassic (POT; hydrothermal K-feldspar ± biotite ± magnetite), and mixed potassic (POTM) alteration zones. c) Ag grades from the Seabridge Gold EOY 2018 block model; d) Au grades from the EOY 2018 block model; e) Cu grades from the EOY 2018 block model; f) Mo grades from the EOY 2018 block model.

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| 7.2.3 | Mitchell Zone

The Mitchell Zone crops out in Mitchell Valley, through an erosional window exposing the footwall of the MTF. The zone is a roughly cylindrical gold-copper-molybdenum deposit, dipping ~60° to the northwest, with approximate true dimensions of 1,600 m in strike, 1,500 m down dip, and up to 850 m in thickness. It remains open down dip and along strike to the northeast at depth.

Recent glacial melt back has provided exceptional surface exposure of a gold-copper-molybdenum porphyry system. A zone of intense quartz and sulphide veining (“Sheeted Quartz Vein Zone”) forms resistant bluffs in Mitchell Valley. However, the higher-grade core area is mostly covered by talus and moraine west of the bluffs. Active oxidation and leaching of sulphides has produced prominent gossans and extensive copper sulphate precipitates at the surface.

Mineralization at the Mitchell deposit is genetically and spatially linked to the Early Jurassic Mitchell intrusive complex, which is composed of Sulphurets (Texas Creek) suite diorite, monzodiorite, and granodiorite stocks and dykes. The intrusive complex cuts sedimentary and volcanic rocks of the Upper Triassic Stuhini Group and sandstones, conglomerates, and andesitic rocks of the Lower Jurassic Jack Formation (basal Hazelton Group). The Mitchell complex has been subdivided into three major intrusive phases. “Phase 1” includes voluminous pre- to early-mineral plagioclase-hornblende diorite porphyry intrusions (196 ± 2.9 Ma and 189 ± 2.8 Ma; U-Pb zircon, Febbo et al., 2019) and a suite of westerly-striking and northerly-dipping monzodiorite dykes. “Phase 2” includes a syn-mineral granodiorite plug within the centre of the deposit (192 ± 2.8 Ma; U-Pb zircon, Febbo et al., 2019), which features abundant xenoliths of quartz veins and Phase 1 intrusions, as well as multiple westerly-striking and northerly-dipping granodiorite dykes in the centre and western half of the complex. Following Phase 2 magmatism, a suite of late-mineral breccia dykes was emplaced. Finally, “Phase 3” is made up of a small diorite plug, measuring 50 m by 125 m, which cuts the breccia dykes.

The successive intrusive phases were accompanied by the development of different hydrothermal alteration assemblages, veining and mineralization within the Mitchell deposit. Hosted by Phase 1 plutons, Stage 1 sheeted quartz – chalcopyrite – pyrite ± magnetite ± K-feldspar ± chlorite veins and stockworks contain most of the copper-gold mineralization, and are accompanied by proximal potassic to calc-potassic alteration and peripheral propylitic alteration. Stage 2 quartz – pyrite – chalcopyrite – molybdenite veining produced a molybdenum-enriched halo around the centre of the deposit (190.3 ±0.8 Ma; Re-Os molybdenite). The molybdenum halo is associated with sericite-pyrite alteration and is temporally related to a Phase 2 stock that outcrops central to the halo. Stage 3 consists of anhydrite veining, poorly mineralized massive pyrite veins, and shallow quartz – pyrophyllite advanced argillic alteration. Whereas the predominant copper-bearing mineral at Mitchell is chalcopyrite, a discrete bornite-bearing zone is found near the centre of the Mitchell Zone towards the hanging wall side. This “Bornite Breccia” was only intersected in three holes (including one interval of 86 m with 1.42% copper and 0.23 g/t gold), and the interpreted dimensions are about 400 m long down dip, 60 m thick, and 250 m along strike. Its geometry roughly aligns with the northwest plunging trend of the Mitchell deposit. The breccia is composed of a chaotic, swirly mix of crackled and milled light grey quartz, anhydrite and clay, with disseminated and interstitial pyrite, chalcopyrite, bornite, and minor tennantite and molybdenite. In deeper intersects the breccia transitions to a mostly quartz, anhydrite, pyrite, and chalcopyrite hosting structure with only traces of bornite. The breccia body is interpreted to be related to structurally controlled, late advanced argillic alteration. A similar breccia zone, with both bornite and abundant anhydrite, is found in the Kerr deposit.

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The Mitchell deposit features many characteristics typical of gold-enriched calc-alkaline porphyry copper deposits: calc-alkaline syn-mineral intrusions; hydrothermal alteration assemblages that include deep potassic, peripheral propylitic, and extensive shallow sericitic alteration; and abundant quartz veining. Metals, chiefly gold and copper (in terms of economic value), are generally at low concentrations, finely disseminated, stockwork or sheeted veinlet controlled, and pervasively dispersed over dimensions of hundreds of metres. Grades diminish slowly over large distances; sub-economic grades are encountered at distances of several hundreds of metres beyond the interpreted centre of the system. This is distinct from the Sulphurets and Kerr zones, where there are more abrupt breaks in grade due to higher structural complexity and juxtaposition of weak and moderate grade domains by faulting, both syn-mineral structures controlling breccia contacts, and post-mineral faulting and displacements.

A small portion of the Mitchell Resource (less than 2%) is found in the hanging wall of the MTF, where disseminated and veinlet chalcopyrite occur in hydrothermal breccias, intrusion breccias, magnetite skarn and hornfels altered sediments and volcanics adjacent to non-mineralized Premier suite porphyritic monzonite to syenite intrusions 193.9 ± 0.5 Ma (Kirkham and Margolis, 1995). This portion of the deposit is more typical of alkalic porphyry deposits, with sodic-potassic alteration, relatively low quartz veining, and K-feldspar – magnetite – albite – pyrite – chalcopyrite – specularite veins. This style of mineralization is similar to that observed in the hanging wall of the STF at the Sulphurets deposit (Main Copper Zone).

The deposit was deformed during development of the mid-Cretaceous Skeena fold and thrust belt (SFTB), transforming portions of the deposit into intensely foliated, mylonitic zones, with deformed and flattened quartz veinlets. The main foliation generally strikes to the east-southeast, and dips moderately to the north. The intensity of the foliation and shortening varies within the boundaries of the Mitchell Zone, and is directly related to hydrothermal alteration assemblages. Sections of the deposit featuring rheologically weak rocks, such as strong sericitic alteration, exhibit the most shortening and deformation (Febbo et al., 2015). Furthermore, during SFTB development, Mitchell was severed along the MTF. This offset portion now lies approximately 1,600 m to the east-southeast in the hanging wall of the MTF (Snowfield Zone; Febbo, et al., 2015). Deep drill holes have also indicated the presence of a roughly 50 m thick, banded, mylonitic shear zone that may offset the base of the Mitchell deposit, termed the Basal Shear Zone (BSZ). The BSZ dips to the northwest and appears to parallel the MTF. As the Bornite Breccia and BSZ may have structurally offset portions of the Mitchell Zone, potential remains for additional mineralization to be discovered.

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A geological map of the deposit is presented in Figure 7.8. A representative cross section and 500 masl level plan of the Mitchell deposit, showing lithological units, alteration assemblages, quartz vein abundance contours, gold grade, copper grade, and molybdenum grade, are shown in Figure 7.9 and Figure 7.10, respectively.

Figure 7.8 Geology Map of the Mitchell Deposit

Source: (Seabridge, 2019)

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Figure 7.9 Vertical section through the Mitchell deposit

Source: (Seabridge, 2019)

Notes: Location is shown as A-A’ Figure 7.9. The traces of the Mitchell thrust fault (MTF) and Mitchell basal shear zone (BSZ) are shown on each panel. a) Simplified geology of the Mitchell deposit, showing undifferentiated wall rock (gray), the Phase 1 diorite porphyry (pink), the sheeted vein body (red), Phase 2 granodiorite porphyry dykes (darker pink), and post-mineral dolerite dykes (dark green). The traces of drill holes occurring within ± 50 m of the section are shown as black lines. b) Simplified diagram of hydrothermal alteration zoning within the Mitchell deposit, illustrating the extent of potassic alteration, moderate and strong sericitic alteration (see text), as well as the occurrence of hydrothermal anhydrite; c) contours of logged volume percent of total quartz veins within Mitchell drill core; d) Au grades from the Mitchell block model, showing clear concentric zoning; e) block model Cu grades, showing a concentric zoning pattern similar to that of Au; and f) block model Mo grades, showing a zoning pattern with highest Mo grades rimming the central zone of high Au and Cu.

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Figure 7.10 Plan view of the Mitchell deposit at 500 masl

Source: (Seabridge, 2019)

Notes: Location is shown as A-A’ Figure 7.10: a) simplified geology of the Mitchell deposit, showing undifferentiated wall rock (brown), the Phase 1 diorite porphyry (pink), the sheeted vein body (red), Phase 2 granodiorite porphyry dykes (darker pink), quartz-anhydrite-bornite hydrothermal breccia zone (orange), post-mineral dolerite dykes (dark green); drill hole traces in a 50 m envelope are also shown as black lines; b) simplified diagram of hydrothermal alteration zoning within the Mitchell deposit, illustrating the extent of potassic alteration, moderate and strong sericitic alteration, as well as the occurrence of hydrothermal anhydrite; c) contours of logged volume percent of total quartz veins within Mitchell drill core; d) Au grades from the Mitchell block model, showing clear concentric zoning; e) block model Cu grades, showing a concentric zoning pattern similar to that of Au; and f) block model Mo grades, showing a zoning pattern with highest Mo grades rimming the central zone of high Au and Cu.

Seabridge Gold Inc. | 7-17 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 7.2.4 | Iron Cap Zone

The Iron Cap deposit is the northernmost porphyry gold-copper-molybdenum deposit in the KSM district, and occurs structurally above the Mitchell deposit, in the panel of rocks between the MTF and STF. It is now known to be hosted by an Early Jurassic intrusive complex that is roughly contemporaneous with those in the Mitchell and Kerr zones. The Iron Cap mineralized zone forms a tabular body striking roughly north-south and dipping ~60° to the west, with approximate true dimensions of 1,500 m in strike, 1,500 m down dip, and up to 800 m in thickness. Mineralization remains open down dip.

Only the southern and eastern margins of the Iron Cap deposit outcrop at surface, where weak to moderate gold + copper ± molybdenum mineralization is primarily hosted in sedimentary strata of the Hazelton Group. It is now known that Iron Cap glacier covers the central and northern portions of the deposit, and that the western side of the deposit is concealed beneath the largely barren hanging wall of the STF. Drilling campaigns during the 2016, 2017, and 2018 field seasons, which included drill holes that collared on the glacier and on the ridge to the west of Iron Cap, significantly expanded the known resource at Iron Cap.

The Early Jurassic Iron Cap intrusive complex is composed of multiple intrusion and breccia phases. One of the earliest phases is a pre-mineral, medium-grained plagioclase-hornblende-phyric diorite with porphyritic texture found on the southeast side of the complex (“P2 East”). A second plagioclase-hornblende-phyric diorite phase, finer-grained than the pre-mineral diorite, is located in the northwest of the complex (“P2 West”). The P2 West phase is thought to be syn-mineral, as it hosts significant volumes of early quartz-sulphide veins (A-type and B-type veins, based on the vein classification scheme of Gustafson and Hunt, 1975), and because this diorite phase is spatially associated with some of the highest gold and copper grades observed at Iron Cap. Syn-mineral, plagioclase-K-feldspar-hornblende(-quartz)-phyric monzonite intrusions (“P3 East”) occur in the central to eastern parts of the Iron Cap intrusive complex. The P3 East monzonites are typically medium-grained, with seriate texture, and include a number of thin dykes as well as a ~100 m thick tabular intrusion striking roughly north through the centre of the Iron Cap deposit, and dipping ~60° west. As with the syn-mineral diorite, portions of the P3 East monzonite host significant volumes of early quartz-sulphide A-type and B-type veins and strong gold-copper mineralization. Finally, a suite of weakly-mineralized monzonite dykes and intrusions, coarser-grained than the central syn-mineral monzonite, occurs on the western side of the intrusive complex (“P3 West”). A radiometric U-Pb zircon age of 195.4 ± 2.1 Ma was obtained for the P3 East monzonite, which is in line with the other radiometric ages obtained for porphyry emplacement and mineralization within the Sulphurets district (190.3 ± 0.8 Ma to 197 ± 3 Ma; Bridge, 1993; Margolis, 1993; Febbo et al., 2019a; Febbo et al., 2019b). Unfortunately, U-Pb zircon age dating is not possible for the diorite intrusive phases at Iron Cap, as they do not contain zircon as an accessory mineral. The relative ages of the P2 West, P3 East, and P3 West intrusions are also unclear, due to the lack of cross-cutting relationships, but they are assumed to be roughly contemporaneous.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Mineralized hydrothermal breccias are significantly more abundant and voluminous at Iron Cap than at the Mitchell, Sulphurets, or Kerr deposits. Tabular, steeply dipping and northerly striking mineralized hydrothermal breccia bodies, ranging from ~10 m to ~250 m wide, occur within central Iron Cap. The breccias are matrix-supported and contain poorly-sorted subangular clasts. The clasts within the breccias are commonly too altered to permit protolith identification, but clasts of P3 monzonite are sometimes observed. Jigsaw hydrothermal breccias are also common along the margins of the P2 East intrusion, with subangular to angular monomict clasts of the P2 East intrusion in a hydrothermally-altered rock flour matrix.

Post-mineral intrusions in the Iron Cap intrusive complex are limited to a small number of volumetrically insignificant dykes. K-feldpar-phyric monzodiorite dykes have been identified in the northern half of Iron Cap. These plagioclase-K-feldspar-hornblende-phyric dykes contain plagioclase and hornblende phenocrysts up to 8 mm and K-feldspar phenocrysts up to 2 cm, and closely resemble the post-mineral K-feldspar megacrystic dykes observed at the Kerr deposit. In the Sulphurets district, these dykes are thought to represent the last pulse of the Early Jurassic porphyry magmatic system. Rare aphanitic, mafic to ultramafic, dark green to blackish post-mineral dykes with meter-scale thicknesses also occur. These dykes are common throughout the rest of the Sulphurets district and may have been emplaced during Cretaceous deformation.

The Iron Cap deposit features hydrothermal alteration assemblages typical of porphyry Cu-Au deposits, including deep, central potassic alteration with an assemblage of K-feldspar – magnetite ± biotite (now chloritized), weak peripheral propylitic alteration with an assemblage of chlorite – epidote – carbonate ± hematite ± magnetite, and sericitic alteration. Potassic alteration is most commonly observed in the P2 West, P3 East and P3 West intrusions, as well as the deep hydrothermal breccias. Sericitic alteration zones, featuring an assemblage of fine-grained muscovite or illite – quartz – pyrite, are concentrated in the upper half of the deposit as well as within the hydrothermal breccias. The intensity of the sericitic alteration varies from the replacement of felsic minerals, to the partial replacement of mafic minerals, to the complete and texturally destructive replacement of all primary felsic and mafic phenocrysts. Strong sericitic alteration obliterates most original textures, making protolith identification difficult in places. Though present in overall minor quantities, the greatest abundances of galena, sphalerite, and silver or arsenic sulphosalts are found in the upper half of the Iron Cap Zone, and may be a consequence of telescoping or downward migration of the hydrothermal system. Hydrothermal anhydrite, which occurs abundantly in certain parts of the Kerr and Mitchell deposits, is conspicuously absent at Iron Cap.

Total quartz vein abundances at Iron Cap typically range from <1 vol.% to ~5 vol.%. However, a handful of discrete, narrow zones at Iron Cap feature notably high (>10 vol.%) abundances of sheeted A-veins, which correspond to zones of particularly elevated gold and copper grades. Such zones are notably observed within the P3 East and P2 West intrusive phase. The vein types present at Iron Cap include: early quartz – chalcopyrite ± magnetite “A-veins”, magnetite ± chalcopyrite veins with selvedges of pink hydrothermal K-feldspar ± magnetite; quartz – chalcopyrite ± molybdenite ± pyrite “B-veins,” with a distinctive sulphide-bearing central suture surrounded by quartz; gray quartz veinlets with a variable sulphide-sulphosalt assemblage of pyrite ± tennantite ± tetrahedrite ± chalcopyrite ± sphalerite, which crosscut the earlier A-veins and B-veins; pyritic “D-veins” with sericite-pyrite selvedges; and post-mineral quartz – carbonate ± chlorite ± sphalerite ± galena ± chalcopyrite ± tennantite ± tetrahedrite veins, which are thought to be associated with mid-Cretaceous deformation. High silver values are generally associated with presence of galena and sphalerite.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Microscopic examinations of polished thin sections confirm that Iron Cap was also subjected to post-mineral deformation, as evidenced by widespread mylonitic textures.

The Iron Cap Zone terminates at the south along the north-dipping Iron Cap Fault (ICF). South of the fault, hornfelsed sediments are mineralized with marginal gold and copper grades similar to intervals above the MTF at Mitchell. A few holes through this area contain higher than average molybdenum grades, including in interval of 133 m with 0.10% molybdenum.

A geological map of the deposit is presented in Figure 7.11. A representative cross section and 1200 masl level plan of the Iron Cap deposit, showing lithological units, alteration assemblages, quartz vein abundance contours, gold grade, copper grade, and molybdenum grade, are shown in Figure 7.12 and Figure 7.13, respectively.

Figure 7.11 Iron Cap geology map. The traces of the Sulphurets Thrust Fault (STF), Johnstone Fault (JF), and Iron Cap Fault (ICF) are shown.

Source: (Seabridge, 2019)

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Figure 7.12 Vertical cross-section through the Iron Cap deposit, looking NNE

Source: (Seabridge, 2019)

Notes: Location of section is shown on Figure 7.12. The traces of the Sulphurets Thrust Fault (STF) and the Iron Cap Fault (ICF) are shown. a) Simplified geology of the deposit, showing undifferentiated wall rock (WR) and the principal intrusions and breccias of the Iron Cap intrusive complex: P2 East (P2E), P3 West (P3W), P3 East (P3E), and hydrothermal breccias (BX). See text for unit descriptions. The traces of drill holes occurring within ±100 m of the section are shown as black lines. b) Hydrothermal alteration zoning illustrating the extent of potassic (POT.), moderate sericitic (MOD. SER.) and strong sericitic (STR. SER.) alteration assemblages (see text), as well as the extent of arsenic concentrations >50 ppm; c) contours of logged volume percent of total quartz veins within drill core; d) Au grades from the Seabridge Gold EOY 2018 block model; e) Cu grades from the Seabridge Gold EOY 2018 block model; and f) Mo grades from the Seabridge Gold EOY 2018 block model.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Figure 7.13 Plan view through the Iron Cap deposit at 1200 m elevation

Source: (Seabridge, 2019)

Notes: The traces of the Sulphurets Thrust Fault (STF), Johnstone Fault (JF), and Iron Cap Fault (ICF) are shown. a) Simplified geology of the deposit, showing undifferentiated wall rock (WR) and the principal intrusions and breccias of the Iron Cap intrusive complex: P2 East (P2E), P2 West (P2W), P3 West (P3W), P3 East (P3E), hydrothermal breccias (BX) and post-mineral K-feldspar megacrystic dykes (P4). See text for unit descriptions. The traces of drill holes occurring within ±100 m of the section are shown as black lines. b) Hydrothermal alteration zoning illustrating the extent of potassic (POT.), moderate sericitic (MOD. SER.) and strong sericitic (STR. SER.) alteration assemblages (see text), as well as the extent of arsenic concentrations >50 ppm; c) contours of logged volume percent of total quartz veins within drill core; d) Au grades from the Seabridge Gold EOY 2018 block model; e) Cu grades from the Seabridge Gold EOY 2018 block model; and f) Mo grades from the Seabridge Gold EOY 2018 block model.

Seabridge Gold Inc. | 7-22 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 8.0 | DEPOSIT TYPES

Porphyry copper ± gold ± molybdenum systems are defined as large volumes (~10 km3 – 100 km3) of mineralized and hydrothermally altered rocks centered on porphyry intrusions (Sillitoe, 2010). Porphyry systems are generated at convergent plate margins and along magmatic arcs above subduction zones. These systems typically span the uppermost ~4 km of the crust (Singer et al., 2008), with their stocks and dykes connecting downwards to deeper parental magma chambers that double as the source of high-temperature, high-pressure metalliferous fluids for the system (e.g., Cloos, 2001). The parental magmas are water-rich, highly oxidized, and exceptionally sulphur-enriched, and range in composition from granite to diorite to monzonite (e.g., Sillitoe, 2010). A metal-rich aqueous phase is released from cupolas atop the parental chambers during cooling and fractionation, and is episodically transported upward along with porphyry dykes. As they rise, the magmas and fluids undergo drastic temperature, pressure, and chemical changes that result in the precipitation of metals and formation of distinctive hydrothermal mineral assemblages in an upward and outwardly zoned pattern.

The KSM deposits display many diagnostic features of porphyry Cu ± Au ± Mo systems. The deposits are centered on intrusive complexes composed of Early Jurassic, Texas Creek suite porphyry stocks and dykes, like the Kerr, Mitchell, and Iron Cap Zones which display the typical lateral and vertical zoning sequence of alteration assemblages observed in many porphyry systems: deep central potassic alteration, peripheral propylitic alteration, and shallow sericitic alteration. Mineralization is associated with quartz veinlet stockworks and sheeted quartz veinlet arrays, with vein density decreasing in later phases. Host rocks may be mineralized for up to several hundred meters from the intrusions. The Kerr and Mitchell zones also feature small remnants of advanced argillic alteration. The structurally complex Sulphurets deposit does not feature the same clear alteration zoning patterns observed at the other three deposits, due to its dismembered and fragmental nature. However, the Lower Panel fault block at Sulphurets, which hosts the bulk of the mineralization, features potassic alteration and mineralization typical of porphyry Cu ± Au ± Mo systems.

Other large gold-enriched porphyry copper deposits, comparable to those in the KSM district, include Pebble (Alaska), Bingham (Utah), Grasberg (Indonesia), and Galore Creek (British Columbia). The KSM district porphyry Au-Cu-Mo deposits feature large volumes of plagioclase-hornblende-phyric diorite that is intimately associated with mineralization and smaller volumes of more evolved magmas, including granodiorite and monzodiorite dykes. Diorite is at the mafic end of the compositional spectrum of magmas commonly associated with porphyry Cu ± Au ± Mo formation. Another example of a gold-enriched porphyry copper deposit with syn-mineral diorite intrusions is Cerro Casale, Chile (e.g., Palacios et al., 2001).

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Porphyry Cu ± Au ± Mo systems are subdivided into two categories: calc-alkaline and alkaline Cu ± Au ± Mo deposits. The KSM district deposits mainly display features that are characteristic of calc-alkaline porphyry Cu ± Au ± Mo deposits: high quartz vein volumes, strong pyrite mineralization, strong sericitic alteration, and an association with calc-alkaline intrusions. However, discrete zones within the district, restricted to fault panels that have been thrust over the main deposits, display characteristics consistent with alkalic style porphyry mineralization. These zones include the hanging wall of the MTF above the main Mitchell deposit, as well as the Upper Panel fault block of the Sulphurets deposit in the hanging wall of the STF. These alkalic-style zones contain largely barren porphyritic monzonite to syenite intrusions with potassic alteration and reddish hematite dusting, surrounded by volcano-sedimentary wall rock and hydrothermal breccias that contain weak to moderate copper-gold mineralization. The alkalic-style mineralized zones represent only a very minor portion of the total district resource.

The principal sulphides observed in the KSM deposits are pyrite and chalcopyrite, with minor molybdenite, and trace amounts of sphalerite, galena, tennantite-tetrahedrite, bornite, enargite, and arsenopyrite. Magnetite and hematitized magnetite are common, especially in deeper parts of the deposits, and hydrothermal anhydrite is common in certain areas of the Kerr and Mitchell deposits. Native gold is rarely observed macroscopically, and primarily occurs as microscopic clusters at sulphide grain boundaries or as inclusions. All mineralization is hypogene, except for a small remnant of preserved supergene mineralization at the upper limits of the Kerr deposit where chalcocite coatings on pyrite and chalcopyrite have been observed, and at the Main Copper (Sulphurets) occurrence where a remnant of leached capping and partial oxide mineralization is preserved at the highest elevations.

Seabridge Gold Inc. | 8-2 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 9.0 | exploration

This section summarizes Seabridge’s 2011 through 2019 non-drilling exploration programs at KSM. Since 2006, the majority of Seabridge’s exploration activities have been focused on various drilling programs that were designed to test recognized mineralized zones as well as several exploration targets. The following sections summarize by year various non-drilling exploration activities that Seabridge has completed.

| 9.1 | 2011 Geophysical Exploration Program

A Spartan MT (magneto telluric) survey was carried out by Quantec Geoscience Limited, over a period of 37 days from July 21, 2011 to August 26, 2011. A total of 175 MT Logger sites at an average spacing of 500 m were completed covering the KSM property from Kerr to Iron Cap zones and tunnel route north of the KSM grid. The data were collected over a frequency range of 320 Hz to 0.001 Hz with variable site spacing. The exploration objectives of the Spartan MT survey at the KSM Property were to detect mineralization and/or associated alteration zones, establish an understanding of the geological system and fluid pathways to great depth, and generate a 3D geophysical model by MT data to integrate with KSM geological model.

| 9.1.1 | Results of 2011 Geophysical Program

The 3D inversion results indicate that the subsurface resistivity, from the surface to a depth of approximately 2 km, varies over a range of 100 Ωm to 5000 Ωm. The southern half of the KSM property reveals a large conductive body with a thickness of more than 2 km, interpreted to be a large magmatic intrusion and favourable for porphyry system. The central and northern parts of the KSM Property show relatively shallow conductive zones with thickness of approximately 500 m and are associated with alteration zones, which are the results of large scale thrust faulting in the area. The resistivity model also helped to map a number of sub-parallel faults and lineaments.

| 9.2 | 2013 Geophysical Exploration Program

SJ Geophysics Ltd. was contracted to acquire geophysical data in several boreholes at KSM for Seabridge. Borehole total magnetic field, induced polarization (IP), and time domain electromagnetic (TDEM) techniques utilizing SJ Geophysics’ borehole Volterra-EM system were conducted on five drill holes at the Kerr, McQuillan, and Iron Cap zones. The objectives were to determine if the known deposit zones have an electromagnetic and/or magnetic signature and to identify if there is a relation between quantities of magnetite and gold-copper mineralization.

Seabridge Gold Inc. | 9-1 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 9.2.1 | Results of 2013 Geophysical Program

The EM data did not identify any strong conductors in the holes surveyed. Based on the current holes surveyed, there does not appear to be an EM response to the deposit. In borehole K-13-24, a weak off-hole EM anomaly was identified that is believed to correlate with a low resistivity zone. Unfortunately, IP data were not collected in this hole, so this cannot be verified. In drill hole K-13-25, chargeability was found to increase with depth, and more chargeable rocks are located towards the east of the drill hole.

In borehole MQ-13-06, the total field magnetic data show three major magnetic units are present. Two off-hole EM anomalies were identified, with the larger one being located at approximately 850 m and to the southeast of the drill hole, mostly likely within 50 m to 100 m. Again, this conductor was poorly coupled. It may be better to use smaller, more focused loops in the future.

In the Iron Cap zone, boreholes IC-13-48 and IC-13-49 showed very good agreement between each other. From the magnetic data, the boundaries of a relatively constant magnetic unit were identified. This unit correlates with a low resistivity and high chargeability unit based on the IP data. It is believed that this may be a mineralized zone. The IP data from IC-13-48 indicates that more chargeable rocks are located towards the east and south of the drill hole.

S.J.V. Consultants Ltd. (SJV) was retained to review and interpret all historical geophysical data with the intention of identifying characteristic signatures that might be related to observed mineralization and geology. The IP data were reviewed and input to the UBC 2D resistivity and IP inversion algorithms, with the aim of improving the available models by using more modern algorithms as well as a finer cell size. The output models show some significant differences from the previous results and also have significantly better resolution of features at depth, as a result of the finer cell size.

The Kerr deposit overlies a resistive body but is itself hosted within a moderately conductive zone, close to the background level. There are no available IP or magnetic data for this area. The Sulphurets deposit also sits on the margins of a resistive body. The more detailed IP resistivity identified a narrow, dipping conductive zone that correlates very well with the actual mineralization. The magnetic susceptibility in this area is moderate. The Mitchell deposit is located along a conductive feature that correlates well with Mitchell creek. This is interpreted to represent a fault zone that may have provided a pathway for mineralizing fluids and/or intrusions. Mitchell sits within a clear magnetic low, which is interpreted to represent magnetite-destructive alteration. The Iron Cap deposit is located within an area of high chargeability, but was not well defined on either the MT or IP resistivity data. The mineralization sits on the boundary of a large magnetic high that has significant roots in the 3D inversion.

The most direct correlation between the geophysics and the mineralization occurred at the Mitchell deposit, which is hosted by a prominent magnetic low, interpreted to be a result of magnetite-destructive alteration. Other magnetic lows within the property are therefore considered high priority targets. The 2D IP provides better resolution of the resistivity at relatively shallow depths than the MT survey, which is focused on deep exploration. Interpretation of the IP was somewhat limited as the survey covers only the northern portions of the KSM area and the lines are spaced too far apart to allow good correlation of features between lines. The relationship of the mineralization to the IP resistivity and chargeability was variable and complicated. Low resistivity features within high chargeability zones could be correlated to the mineralization at the Sulphurets deposit and are therefore recommended as possible future targets.

Seabridge Gold Inc. | 9-2 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 9.3 | 2014 Geophysical Exploration Program

Quantec Geoscience Ltd. was contracted to perform Spartan MT and AMT (audio magneto telluric), gravity, and IP surveys over the Property. The MT/AMT survey was designed to expand coverage of the 2011 survey and selectively infill areas at a higher resolution with readings at approximately 250 m spacings. Gravity readings were collected at the same survey sites. Two east-west lines of IP surveying were made, one in Sulphurets valley and one in Mitchell valley.

| 9.3.1 | Results of 2014 Geophysical Programs

The gravity survey utilized a L&R Model G743 Base station, and the Geosoft Reduction processing platform with density of 2.65 g/cc. Corrections were made for latitude, terrain (regional and local), free-air anomaly calculation, and bouguer anomaly calculation. There were a total of 154 readings unevenly distributed over an area of about 75 skm2. The strong correlation between high readings and high topography suggests the topography is too extreme and the model too coarse to resolve and interpret the data correctly.

The Spartan MT survey utilized the same equipment and parameters as the survey done in 2011. A total of 49 readings produced acceptable results, and 23 readings were rejected due to excessive “noise”. Attempts to obtain readings on permanent snow and ice were unsuccessful, so the area covered was not expanded significantly from 2011.

IP surveying was performed on east-west lines of about 5 km each in the bottom of Sulphurets and Mitchell valleys, to reach the minimum elevation penetration of the survey. In Sulphurets valley, a strong chargeability response occurs east of the trace of the Sulphurets thrust fault and reflects continuity of pyritic alteration between Kerr and Sulphurets deposit. Similarly, in Mitchell valley high chargeability responses occur east of the STF trace.

| 9.4 | 2015 Geophysical Exploration Program

Precision Geosurveys Inc. was contracted to run a helicopter supported magnetic and radiometric survey covering almost all of the KSM claims, to improve resolution of the subsurface geological model in undrilled areas. The magnetic sensors are flown in a non-magnetic and non-conductive nose stinger configuration with 3D compensation, which allows the survey to be safely flown at reduced terrain clearance to minimize noise, improve resolution, and reduce the need for complex corrections to the data. The survey covered an area of about 1,600 km2. at a flight line spacing of 100 m. KSM mineralized areas are reflected by low susceptibilities where overprinting phyllic alteration has converted magnetite to pyrite, and high susceptibilities where potassic alteration with magnetite is still preserved.

Seabridge Gold Inc. | 9-3 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 9.5 | 2019 Geophysical Program

During 2019, Geotech Ltd. carried out a helicopter-borne geophysical survey which included a caesium Z-axis tipper magnetometer, which is known as a Z-Axis Tipper Electromagnetic (ZTEM) system, a GPS navigation system and a radar altimeter. In a ZTEM survey, a single vertical-dipole air-core receiver coil is flown over the survey area in a grid pattern, similar to regional airborne EM surveys. Two orthogonal, ferrite-core horizontal sensors are placed close to the survey site to measure the horizontal EM reference fields. Data from the three sensors are used to obtain the Tzx and Tzy tipper components at six frequencies in the 30 Hz to 720 Hz band. The ZTEM is useful in mapping geology using resistivity contrasts and magnetometer data provides additional information on geology using magnetic susceptibility contrasts. The survey covered an area of 190 km2. with a total of 1,000 line km flown at a 200 m spacing in an east-west direction. The survey was carried out between July 25 to September 8, with 23 days lost to weather, 10 due to equipment malfunctions, and 9 survey days flown.

In July and August, Dias Geophysical Ltd. was contracted to perform an IP survey over part of the Sulphurets zone area extending westward, to assist in mapping sulphide distributions and orientations in the area of drill testing, covering 3 km2. The survey utilized a continuously rolling distributed array pole-dipole using the common voltage reference (CVR) method, using a 200 m line spacing and a receiver spacing of 100 m. This approach has inherent flexibility in calculating multi-scale dipoles from the acquired data. The survey data can be processed in the conventional fashion, in which 50 m dipoles are calculated for the entire survey. If the signal to noise ratio is found to be too low the data can be reprocessed to increase the dipole separation to 200 m or 400 m to increase the ratio. The southeast corner of the survey covered a portion of the Sulphurets zone beneath the STF, and high chargeability features correlate very well with the modelled sulphide distribution previously determined from drilling. The exploration area targeted in 2019 above the STF displayed more limited and discontinuous high chargeability areas, which drilling confirmed were due to high concentrations of pyrite with minor chalcopyrite in veins, disseminations and patchy replacements associated with skarn alteration near intrusive/wall rock contact zones.

Mira Geoscience was contracted to undertake a regional district scale integrated model and targeting exercise for KSM, including the compilation of all digital exploration data and construction of an integrated 3D geological and structural model, as a basis for constrained inversion of magnetics, MT and IP surveys data. Modelling is based on available historical data, literature, geology, structures, geochemistry and geophysics datasets provided by Seabridge. The integrated 3D model was used to produce a set of targets based on a porphyry deposit concept involving both knowledge driven, and data driven (machine learning) analysis of the data and model.

Deep penetrating geophysical techniques were employed to improve resolution on targets and generate discrete zones for testing in the future. New ZTEM surveys and 3D IP surveys were completed. Data was integrated into a digital 3D earth model by Mira Geoscience. These results are now being integrated with historical MT surveys, airborne high-resolution magnetic survey, bore hole geophysical surveys and geological mapping. Geophysical profiles indicate that these targets can be tested from surface but would likely be evaluated as bulk underground opportunities.

Seabridge Gold Inc. | 9-4 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 10.0 | Drilling

| 10.1 | Introduction

Drilling methods, procedures, extent of drilling, and relevant results for the Property were described in previous NI 43-101 Technical Reports prepared by the QP responsible for this section of this Report (Lechner 2007; 2008a; 2008b; 2009; 2010; 2011; 2012; 2014, and 2016). Plan maps and representative drill hole cross sections through the mineral deposits are shown in Figures 7.2, 7.5, 7.8, 7.11, 14.4 to 14.7, 14.12 to 14.15, 14.20 to 14.23, and 14.28 to 14.31 to illustrate a summary and interpretation of the drilling results.

The majority of KSM drilling information that is stored in the end-of-year 2019 AcQuire database was collected by Seabridge (86%). Seabridge has conducted annual drilling campaigns at KSM beginning in 2006. The remaining 14% of the drilling data were collected by Placer (8%) and Falconbridge/Noranda (about 1%), with the remainder collected by six other companies (5%). The 2005 Falconbridge drill campaign was conducted as a joint venture with Seabridge. A summary of all KSM drill hole data organized by year is shown in Table 10.1. The majority of the 742 core holes shown in Table 10.1 were used to estimate Mineral Resources disclosed in this Report, but some of the data tested several non-resource targets in the KSM Property. Table 10.2 summarizes drilling data by company through 2018. The companies listed in Table 10.2 have been arranged in approximate chronological order starting with Esso Minerals in the 1960s. Minor core drilling was completed at KSM by several companies in the early 1980s, but ramped up significantly in the late 1980s and early 1990s by Placer Dome. Seabridge systematically added to the KSM drill hole data after their entry into the district in 2000, with annual drill campaigns beginning in 2005.

Seabridge Gold Inc. | 10-1 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Table 10.1 KSM Historic Drilling Through 2019

Table 10.2 Historic Drilling by Company Through 2018

Seabridge Gold Inc. | 10-2 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Table 10.3 summarizes drilling at KSM by deposit and company through 2018. The data shown in Table 10.3 are the drill holes that were used to estimate Mineral Resources that are the subject of this Technical Report.

Table 10.3 KSM Drill Hole Summary by Area and Company Through 2018

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| 10.2 | Type and Extent of Drilling

Seabridge contracted Boart Longyear to complete the initial 2006 core drilling program. Since then, Hy-Tech Drilling Ltd. from Smithers, BC has completed all of Seabridge’s Mineral Resource definition core drilling at KSM through 2018 using their own manufactured Tech-5000 Fly Rigs. Drilling was completed using either HQ or NQ tools. Helicopter support for Seabridge’s KSM exploration programs has been provided by Lakelse Air Ltd. from Terrace, BC since 2007. This long-standing relationship with local drilling and air support contractors has allowed for a continually growing understanding about local drilling conditions. The drilling operations were conducted from the Sulphurets Creek camp which is located northwest of the Kerr deposit.

In 2019 Seabridge contracted with Driftwood Diamond Drilling out of Smithers, BC to complete a 26-hole diamond core drilling program at the Sulphurets deposit. Those holes were not used to estimate Mineral Resources that are the subject of this Report because the results were not available at the time this report was written.

Figure 10.1 is a drill hole location map for the entire KSM district, showing all of the drilling data that were available to estimate Mineral Resources that are the subject of this Report (drilling through 2018). The drill holes are colour coded (blue represents non-Seabridge and red represents Seabridge drilling). Detailed drill hole location maps are presented in Figures 10.2 to 10.5 for the Kerr, Sulphurets, Mitchell, and Iron Cap deposits, respectively. Figures 10.2 to 10.5 also show the outline of conceptual resource pits and resource block cave that define the Mineral Resources for the Property.

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Figure 10.1 KSM Drill Hole Locations

Source: (RMI, 2019)

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Figure 10.2 Drill Hole Locations – Kerr Deposit

Source: (RMI, 2019)

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Drilling at the Kerr deposit has identified a mineralized area measuring roughly 2,400 m north-south by 800 m east-west, and about 2,200 m vertically. The drill hole spacing in the upper open pit resource area is approximately 50 m to 75 m. Drill hole spacing through the block cave resource, which has been classified as nearly all Inferred material, ranges between 100 m to 200 m. Representative drill hole cross sections and level maps through the Kerr deposit are shown in Figures 14.4 to 14.7.

Figure 10.3 Drill Hole Locations – Sulphurets Deposit

Source: (RMI, 2019)

Drilling at the Sulphurets deposit has identified a mineralized area measuring roughly 2,200 m northeast-southwest by 550 m northwest-southeast, and about 330 m vertically. The drill hole spacing in the open pit resource area ranges between 50 m to 75 m. Figures 14.12 to 14.15 show representative drill hole cross sections and level maps through the Sulphurets deposit.

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Figure 10.4 Drill Hole Locations – Mitchell Deposit

Source: (RMI, 2019)

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Drilling at the Mitchell deposit has identified a mineralized area measuring roughly 1,600 m east-west by 1,500 m down-dip, and 850 m thick. The drill hole spacing in the upper open pit resource area is approximately 75 m to 100 m. Drill hole spacing through the block cave resource, which has been classified predominantly as Inferred material, ranges between 100 m to 200 m. Representative drill hole cross sections and level maps are shown in Figures 14.20 to 14.23.

Figure 10.5 Drill Hole Locations – Iron Cap Deposit

Source: (RMI, 2019)

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Drilling at the Iron Cap deposit has identified a mineralized area measuring roughly 1,500 m northeast-southwest, by 1,500 m northwest-southeast, and about 850 m thick. The drill hole spacing in the upper block cave resource shapes ranges from 70 m to 75 m. Drill hole spacing through the lower block cave resource, which has been classified predominantly as Inferred material, ranges from 100 m o 200 m. Representative drill hole cross sections and level maps through the Iron Cap deposit are shown in Figures 14.28 to 14.31.

| 10.3 | Drilling Procedures

As mentioned in Section 10.1, all of the drilling data collected between 2007 and 2018 was completed by one drill contractor, Hy-Tech Drilling Ltd. located in Smithers, BC. The drilling was completed using Hy-Tech’s Tech-5000 Fly Rigs utilizing HQ, NQ, BQ, and AQ rods. Nearly all of the helicopter support was provided by Lakelse Air Ltd. using Eurocopter A-Star machines. Drilling operations were conducted from the Sulphurets Creek camp, which is located northwest of the Kerr deposit.

Seabridge used directional drilling and conventional wedging methods for a portion of some of their deeper drilling programs at Kerr, Iron Cap, and Mitchell. Tech Directional Services Inc. from Ontario, Canada were contracted to provide directional drilling services using DeviDrill equipment. DeviDrill uses a steerable wireline core barrel that allows a “daughter” hole to be wedged off of a “mother” hole and vectored towards a target zone with reasonable accuracy. Small diameter core (AQ) was retrieved during the crucial turn away from the mother hole so minimal data was lost. Bearing and inclination data were collected using a miniature electronic single-shot survey tool (DeviTool PeeWee) that is designed to pass through the DeviDrill bit. Information regarding this drilling method can be found at http://www.techdirectional.com/. Drill holes with a letter designation after the hole number represent wedged drill holes that utilized the directional drilling method. A total of 34 daughter holes were wedged off mother holes totalling approximately 25,253 m.

Drill core was placed into wooden core boxes by the drill contractor at the rig and delivered twice daily by helicopter from the rigs to Seabridge’s Sulphurets Creek camp. An inventory of the core was completed by Seabridge geologists, which included a review of core condition, a check of run block depths, and generation of a quick down-hole lithologic log.

The drill core was typically scanned for various base metal quantities using a Niton hand held x-ray fluorescence (XRF) analyzer prior to cleaning the core. Seabridge has determined that a factor of 2.0 to 2.2 times the Niton copper reading closely approximates the assayed copper content percentage. The Niton readings are primarily used to alert/train the logging geologist about apparent mineralized intersections. That data was written on the core with wax markers and are visible on core photos. Magnetic susceptibility was also recorded for each drill hole using a handheld device. The mag readings were exported from the device as .csv files, but the data is not currently being used.

After cleaning, the core was logged for lithology, alteration, structure, and oxidation state onto paper logs by Seabridge geologists. That information was later entered into Microsoft Excel® spreadsheets by each logger. Separate paper logs were used to capture geotechnical information like core recovery, rock quality designation (RQD), and fracture count. The geotechnical logs are based on data between core run blocks.

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Assay samples were laid out by the logging geologist. Samples were primarily laid out in 2-m lengths, but were broken at distinct lithologic, alteration, or mineralization contacts. Likewise, samples were broken at core diameter changes. The sample data were hand recorded onto paper logs with hole name, from depth, to depth, and various sulphide mineral estimates.

Pieces of drill core (14 cm to 20 cm long) were marked for bulk density determination about every 100 m down-the-hole by the logging geologist by labelling the wooden core box with “SG”. Those small pieces were not cut for assay sample. Periodically a contract employee weighed the core pieces in air and water so that a bulk density could be calculated.

Prior to sawing, the drill core was photographed. Two close-up photographs were taken for each core box and the two photos were “stitched” together to create a detailed photograph. After all logging procedures were complete the core boxes were moved to the core cutting facilities located adjacent to the core logging tents.

After integrating results from all available drilling, the following observations were made for the Kerr deposit:

| ● | The Kerr deposit is centered on a north-south trending, steep westerly dipping tabular intrusive complex,

| ● | Drilling has demonstrated that the mineralized system has a strike length of 2,400 m and down-dip extents of 2,200 m,

| ● | Deep mineralization is characterized by two north-south west dipping limbs that appear to coalesce near the surface. The west limb is about 500 m thick and the east limb is about 300 m thick,

| ● | There are several intrusive phases present, with the earliest phase being a fine-grained diorite with 5% to 60% quartz-sulphide vein stockworks. This intrusive phase appears to contain the majority of economically interesting mineralization. Later, coarser grained intrusive phases envelop and sometimes invade the earlier phase.

A summary of the interpretation of the Kerr drilling results are illustrated in Figures 7.3 to 7.4 and 14.4 to 14.7.

Based on a review of available drilling data, the following observations were made for the Sulphurets deposit:

| · | The Sulphurets deposit has a lensoidal geometry, striking to the northeast and dipping to the northwest,

| · | Drilling has defined a mineralized system with a strike length of 2,200 m and down-dip extents of 550 m,

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| · | Most of the gold and copper mineralization occurs in a structural block (“lower panel”) that is bounded by the upper Raewyn fault and the lower Mitchell thrust fault,

| · | Drilling has shown that most of the mineralization in the lower panel is associated with veins, breccias, shear fillings, and patchy replacements.

A summary of the interpretation of the Sulphurets drilling results are illustrated in Figures 7.6 to 7.7 and 14.12 to 14.15.

Based on a review of available drilling data, the following observations were made for the Mitchell deposit:

| · | The Mitchell deposit is genetically and spatially related to Early Jurassic intrusions,

| · | Drilling and outcrop mapping have defined a mineralized zone measuring approximately 1,600 m by 1,500 m by 850 m,

| · | The majority of mineralization is hosted by sheeted veins and intense stockworks within potassic and propylitically altered intrusives,

| · | The Mitchell deposit has many similarities with other gold-enriched calc-alkaline porphyry copper deposits (e.g. syn-mineral intrusions and various hydrothermal alteration assemblages like deep potassic, peripheral propylic, extensive shallow sericite alteration and abundant quartz veining).

A summary of the interpretation of the Mitchell drilling results are illustrated in Figures 7.9 to 7.10 and 14.20 to 14.23.

Based on a review of available drilling data, the following observations were made for the Iron Cap deposit:

| · | The Iron Cap deposit is hosted by Early Jurassic intrusions which are bounded by the upper Sulphurets thrust fault and lower Mitchell thrust fault,

| · | Drilling has defined a tabular body that strikes north-south and dips about 60 degrees to the west and measures about 1,500 m by 1,500 m by 850 m although the deposit remains open down-dip,

| · | Pre and syn-mineralized dioritic intrusions are characterized by tabular, steeply dipping hydrothermal breccias,

| · | Post mineral intrusions are limited to volumetrically insignificant thin dykes.

A summary of the interpretation of the Iron Cap drilling results are illustrated in Figures 7.12 to 7.13 and 14.28 to 14.31.

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| 10.4 | Qp Comments Regarding Drilling And Sampling Factors

In general, core recovery for the various KSM drilling campaigns was excellent, averaging approximately 97%. RQD and core recovery were affected in some of the pre-2012 drilling in the upper portions of the Kerr deposit due to a rubble zone created by the dissolution of anhydrite veinlets in various lithologic units. Core recovery in the 2013 to 2016 drilling campaigns at Kerr was excellent at depth within the modeled anhydrite veinlet population where no dissolution has occurred.

No material drilling, sampling, or recovery issues were encountered within the mineralized portions of the other deposits within the KSM Property that were drill tested during the 2006 to 2019 campaigns. Difficulties were encountered while drilling through the Sulphurets fault zone (SFZ) at the Iron Cap deposit. Nearly all of the mineralization at Iron Cap is located beneath the SFZ. Once beneath the SFZ, core recovery improved dramatically.

In the opinion of the Qualified Person responsible for this section of this Report, there are no drilling or sampling factors that could materially impact the accuracy and reliability of the assay results associated with the 2006 to 2019 KSM drilling. Furthermore, the QP responsible for this section of this Report believes that the assays associated with the various KSM drilling campaigns are suitable to be used to estimate Mineral Resources.

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| 11.0 | Sample Preparation, Analysis and Security

| 11.1 | Introduction

This section summarizes sample preparation, analyses, and security of KSM drill samples through the 2018 drilling campaign. Prior Technical Reports provided sampling details and quality control results from various annual drilling programs. In this Technical Report, a summary of sampling and assaying over the past 12 years is provided. In general, Seabridge has employed relatively consistent sampling methods over the years with minor modifications over the past five years regarding some assaying protocols. Initially, Seabridge used Eco Tech as their primary assay laboratory from 2006 to 2011 when Eco Tech was bought out by ALS Chemex, who has acted as Seabridge’s primary assay laboratory since that time to the present. Over the years, Seabridge’s quality control protocols have included the submission of certified standard reference materials (SRMs or “standards”) blanks, and duplicate field samples. Typically, 5 % to 10 % of the assay pulps from the primary laboratory were submitted to a secondary accredited assay laboratory for check assay comparison purposes. The following sections will summarize key aspects of sample preparation, analyses, etc. covering two time periods, pre-2012 and post-2012, primarily to account for the change of primary laboratory in 2011.

| 11.2 | KSM Sample Preparation Methods and Procedures

| 11.2.1 | Statement on Sample Preparation Personnel

Labourers contracted from Tahltan Native Development Corporation conducted all initial sample preparation (sawing and bagging) and were trained by, and under the direct supervision of, geologists employed by Seabridge. Drill core and quality control samples were shipped to the primary assay laboratory’s preparation facility (either to Eco Tech’s preparation facility located in Stewart, BC or ALS Canada’s preparation facility located in Terrace, BC). The prepared samples were then shipped to the primary assay laboratory’s analytical facility (i.e. either Eco Tech’s laboratory located in Kamloops, BC or ALS Canada’s laboratory located in North Vancouver, BC).

| 11.2.2 | Sample Preparation and Dispatch

Drill core, placed in wooden core boxes by the drilling contractor, was loaded into metal baskets and delivered twice daily by helicopter from each drill rig, to Seabridge’s Sulphurets Creek camp facilities. After the core was delivered to the core logging tents, Seabridge geologists took an inventory of the core (a review of core condition, a check of run block depths, and a quick down-hole lithologic log was prepared). After the drill core had been completely logged, the sample starting/ending points were then laid out. A numbered sample tag was stapled to the wooden core box at the beginning of the sample run and the core was photographed. Core boxes were then moved to the core cutting facilities located adjacent to the core logging tents. Two Weatherhaven tents containing four saws with 14-inch diamond impregnated blades designed for rock cutting were utilized for sawing the core longitudinally. The saws were mounted on secure wooden stands at waist height. The saw blades were cooled, cleaned, and lubricated with fresh, non-recirculated water during cutting. The saw operator placed uncut core boxes on tables adjacent to the saws and cut each piece of core sequentially within each marked sample interval. The assay half of the sample was placed in a heavy-duty polythene bag along with the sample number tag. The outside of the polythene bags was marked with the sample number and then stapled shut. The remaining half of the drill core was carefully put back into the wooden core boxes as soon as cutting was complete.

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The polythene sample bags were inventoried and then approximately three polythene sample bags containing HQ diameter core (or six bags with NQ core) were placed into large polyweave (rice) shipping bags. Approximately 30 rice bags were placed into wooden crates (“totes”) for shipment. The totes were flown by helicopter to a landing pad near km 54 on the Eskay Creek Mine Road located behind a locked gate. The totes were then placed into a locked steel sea-going container that can hold approximately eight totes. Once or twice a week, Granmac Services from Stewart, BC, picked up the totes and delivered them to an ALS receiving facility located in either Stewart, BC (pre-2013) or Terrace, BC where the samples were logged into ALS’s system. From the ALS preparation laboratory, the samples were delivered to the ALS assay laboratory located in North Vancouver, BC by either ALS personnel or other commercial carriers.

| 11.2.3 | Pre-2012 Analytical Procedures

At the Eco Tech facilities in Stewart, samples were sorted and dried (if necessary), crushed through a jaw crusher and cone or roll crusher to –10 mesh, then split through a Jones riffle until a –250 g sub-sample was achieved. The sub-sample was pulverized in a ring and puck pulverizer so that 95% of the material passed a –140 mesh screen, then rolled to homogenize. The resulting pulp sample was placed in a numbered paper envelope and securely packed in cardboard boxes. These boxes were shipped via Greyhound freight services to the Eco Tech facilities located in Kamloops, BC.

At the Eco Tech’s laboratory in Kamloops, a 30 g sample size was split out from the pulp envelope and then fire assayed using appropriate fluxes. The resultant doré bead was parted and then digested with aqua regia followed by an atomic absorption (AA) finish using a Perkin Elmer AA instrument. The lower limit of detection for gold is 0.03 g/t or 0.001 oz/t. For other metals, a multi-element inductively coupled plasma (ICP) analysis was completed. For this procedure, a 0.5 g sample was digested with 3 mL mixture of hydrogen chloride, nitric acid, and water at a ratio of 3:1:2 that contained beryllium, which acts as an internal standard, for 90 minutes in a water bath at 95°C. The sample was then diluted with 10 mL of water and analyzed on a Jarrell Ash ICP unit. Eco Tech’s ICP detection limits (lower and upper) are summarized in Table 11.1.

Assay results were then collated by computer and were printed along with accompanying internal quality control data (repeats and standards). Results were printed on a laser printer and were faxed and/or mailed to appropriate Seabridge personnel. Appropriate standards and repeat samples were included on the data sheet.

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Table 11.1 ICP Detection Limits – Pre-2012 Data

Element | Lower | Upper |  | Element | Lower | Upper
Ag | 0.2 ppm | 100.0 ppm |  | Mo | 1 ppm | 10,000 ppm
Al | 0.01% | 10.00% |  | Na | 0.01% | 10.00%
As | 5 ppm | 10,000 ppm |  | Ni | 1 ppm | 10,000 ppm
Ba | 5 ppm | 10,000 ppm |  | P | 10 ppm | 10,000 ppm
Bi | 5 ppm | 10,000 ppm |  | Pb | 2 ppm | 10,000 ppm
Ca | 0.01% | 10.00% |  | Sb | 5 ppm | 10,000 ppm
Cd | 1 ppm | 10,000 ppm |  | Sn | 20 ppm | 10,000 ppm
Co | 1 ppm | 10,000 ppm |  | Sr | 1 ppm | 10,000 ppm
Cr | 1 ppm | 10,000 ppm |  | Ti | 0.01% | 10.00%
Cu | 1 ppm | 10,000 ppm |  | U | 10 ppm | 10,000 ppm
Fe | 0.01% | 10.00% |  | V | 1 ppm | 10,000 ppm
La | 10 ppm | 10,000 ppm |  | Y | 1 ppm | 10,000 ppm
Mg | 0.01% | 10.00% |  | Zn | 1 ppm | 10,000 ppm
Mn | 1 ppm | 10,000 ppm |  |  |  |

| 11.2.4 | Post-2012 Analytical Procedures

ALS served as Seabridge’s primary assay laboratory. ALS is a leading provider of assaying and analytical testing services for mining and mineral exploration companies and has no association or affiliation with Seabridge. All of ALS’s locations are International Organization for Standardization (ISO) 9001:2000 certified.

AcmeLabs served as a check assay laboratory for Seabridge. AcmeLabs is a leading geochemical and assaying facility and has no association or affiliation with Seabridge. In October 2011, AcmeLabs’ Vancouver, BC facility received ISO/International Electrotechnical Commission (IEC) 17025:2005 accreditations from the Standards Council of Canada.

At the ALS preparation facility located in Terrace, BC, samples were sorted and dried (if necessary), crushed through a jaw crusher and cone or roll crusher to 70% –2 mm using ALS protocol CRU-31. The crushed sample was then split using ALS protocol SPL-21 using a riffle splitter. A portion of the crushed sample was replaced into the polythene bag (coarse reject) and stored temporarily at the ALS facility. A portion of the crushed sample was then pulverized using ALS protocol PUL-31 using a ring and puck pulverizer, until approximately a 250 g sub-sample (pulp) was achieved with 85% passing 75 µm or better. The resulting pulp sample was placed in a numbered paper envelope and securely packed in cardboard boxes. These boxes were shipped by ALS to their assay facility located in North Vancouver, BC.

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At the ALS analytical laboratory located in North Vancouver, a 30 g sample was split out from the pulp envelope and then fire assayed using ALS protocol Au-AA23 using appropriate fluxes. The resultant doré bead was parted and then digested with aqua regia followed by an AA finish using an Agilent AA 240 Series instrument. The lower limit of detection for gold is 0.005 g/t.

For other metals, a multi-element ICP-atomic emission spectroscopy (AES) analysis was completed using ALS protocol ME-ICP41. For this procedure, an approximately 0.5 g sample was digested with aqua regia in a graphite heating block. After cooling, the resulting solution was diluted to 12.5 mL with de-ionized water, mixed and analyzed by ICP-AES using an Agilent ICP 720/730-ES Series instrument. The analytical results were corrected for inter-element spectral interferences. ALS’s ME-ICP41 lower and upper detection limits are summarized in Table 11.2 for the elements that Seabridge requested.

Table 11.2 ICP Detection Limits – Post 2012

Element | Units | Lower | Upper |  | Element | Units | Lower | Upper
Ag | ppm | 0.2 | 100 |  | Mo | ppm | 1 | 10,000
Al | % | 0.01 | 25 |  | Na | % | 0.01 | 10
As | ppm | 2 | 10,000 |  | Ni | ppm | 1 | 1,000
B | ppm | 10 | 10,000 |  | P | ppm | 10 | 1,000
Ba | ppm | 10 | 10,000 |  | Pb | ppm | 2 | 1,000
Be | ppm | 0.5 | 1,000 |  | S | % | 0.01 | 10
Bi | ppm | 2 | 10,000 |  | Sb | ppm | 2 | 1,000
Ca | % | 0.01 | 25 |  | Sc | ppm | 1 | 1,000
Cd | ppm | 0.5 | 1,000 |  | Se | ppm | 10 | 10,000
Co | ppm | 1 | 10,000 |  | Sn | ppm | 10 | 10,000
Cr | ppm | 1 | 10,000 |  | Sr | ppm | 1 | 1,000
Cu | ppm | 1 | 10,000 |  | Th | ppm | 20 | 1,000
Fe | % | 0.01 | 50 |  | Ti | % | 0.01 | 10
Ga | ppm | 10 | 10,000 |  | U | ppm | 10 | 1,000
K | % | 0.01 | 10 |  | V | ppm | 1 | 1,000
La | ppm | 10 | 10,000 |  | W | ppm | 10 | 1,000
Mg | % | 0.01 | 25 |  | Zn | ppm | 2 | 1,000
Mn | ppm | 5 | 50,000 |  |  |  |  |

| 11.3 | Summary of the Nature, Extent, and Results of Quality Control Procedures

Since 2006, Seabridge has routinely submitted certified standard reference materials (“standards”) and barren (“blanks”) with their sawn drill core samples at a frequency rate of approximately 1 standard and 1 blank for every 32 to 33 regular samples. At the conclusion of each drilling campaign, about 5% to 10% of the pulps from the primary laboratory were sent to a secondary accredited assay facility providing an independent check of the original pulps. Starting in 2007, field duplicate drill core (sawn ¼ core) samples were collected and sent to the primary assay laboratory at a frequency of about one duplicate sample for every 54 regular samples. Table 11.3 summarizes the number of control samples that were submitted by Seabridge on an annual basis.

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Table 11.3 Summary of KSM Control Samples Submitted Thru Time

In general, very few control samples assayed between 2007 and 2018 failed (i.e. were outside of +/- 3 standard deviation units of the expected value). Many of the failures were attributed to sample labeling errors (e.g. wrong standard listed in drill log). In the early years, some of the river gravel blanks had anomalous traces of copper and/or gold. The ¼ core duplicate sample results provide some insight into sample reproducibility. Table 11.4 summarizes basic statistics associated with original and ¼ duplicate samples collected at KSM between 2007 and 2018.

Table 11.4 ¼ Core Duplicate Sample Statistics (2007-2018)

Figure 11.1 show two box plots (gold on the left and copper on the right) that compare the original ¼ sample with the duplicate ¼ sample.

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Figure 11.1 Duplicate Sample Box Plots

Source: (RMI, 2019)

Figures 11.2 and 11.3 show a number of graphical comparisons between the original and duplicate ¼ samples collected at KSM from 2007 to 2018 for gold and copper, respectively.

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Figure 11.2 Gold Duplicate Sample Graphs

Source: (RMI, 2019)

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Figure 11.3 Copper Duplicate Sample Graphs

Source: (RMI, 2019)

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 11.3.1 | Pre-2012 Quality Control Procedures

Seabridge developed quality control procedures for their 2006 drilling campaign that remain similar to their current protocols. The quality control procedures included the insertion of certified standard reference materials (SRMs), blanks, and field duplicates into the regular drill core sample stream. During the first two drill campaigns (2006 to 2007), “barren” river gravels collected near Stewart, BC were used as blank material. In subsequent years, commercially available landscaping materials (crushed marble and granite) were used as blanks. Blanks were inserted into the sample stream at a rate of about one blank for every 33 regular samples. Rare anomalous blank assays for copper and/or gold were attributed to the uncontrolled nature of the blank material.

Pre-packaged certified standards, primarily from CDN Resource Laboratories Ltd. (CDN Resource), were submitted at a rate of about one standard per 33 regular core samples. These standards were selected based on the expected gold and copper grades and sample matrix makeup. In 2011, Seabridge had two custom standards prepared and certified by CDN Resource using material from the Mitchell deposit (SEA-KSM) and felsic gold bearing material from a Seabridge deposit located in the Northwest Territories (SEA-CL). Seabridge attempted to ensure that their quality control measures included the insertion of at least one sample blank and one standard within each Eco Tech laboratory batch of approximately 35 samples. The blank and pulp standards were numbered using the same number sequence that was used for the core samples and inserted into each batch shipment randomly by the geologist during the sample layout process.

Duplicate field samples consisted of ¼ core samples were collected at a frequency of about one duplicate sample for every 50 regular samples. The HQ and/or NQ core was initially sawn in half and then the original and duplicate sample were created by sawing a specific interval of the half core. The ¼ core pieces were then submitted as original and duplicate samples.

At the conclusion of the drilling/assaying season, Seabridge’s geologic staff selected about 5% to 10% of the primary laboratory assay pulps, which were then sent to an accredited secondary laboratory for “check assay” purposes.

Seabridge tried to ensure that they had at least two control samples per the primary laboratory’s fire assay batches, which typically was 33 samples (the laboratory reserved slots for their own control samples). If more than one control standard (SRM or blank) was outside of three standard deviation units of the expected value, the entire batch was deemed suspect and an order was given to the laboratory to re-assay that entire batch. If a single control sample failed and all of the regular samples were unmineralized then no action was taken. If a single standard failed and the batch contained some mineralized intervals, the control sample plus three to five samples above and below the control sample were re-analyzed. Very few batches were re-run during the period 2006 to 2011.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 11.3.2 | Post-2012 Quality Control Procedures

Seabridge’s post-2012 quality control procedures were essentially the same as those described for the pre-2012 data and have also been described in previous Technical Reports (Lechner, 2014; Tetra Tech, 2016). Standards, blanks, and field duplicates were submitted at similar frequencies as prior years. Control sample results were monitored by Seabridge personnel during each annual drilling campaign. In rare cases where more than one control sample failed, sample batches or three to five samples on either side of the failed control sample were re-run.

| 11.4 | QP’s Opinion

In the opinion of the QP responsible for this section of the Technical Report, sample security, sample preparation, analytical procedures, and QA/QC protocols/results associated with Seabridge’s 2006 to 2018 KSM drilling campaigns were adequate and consistent with standard industry practices. The QP also believes that the assays are suitable to be used to estimate Mineral Resources.

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| 12.0 | Data Verification

The QPs have taken steps to verify the data are adequate for the purposes used in the Report.

In previous KSM NI 43-101 Technical Reports, the QP responsible for Mineral Resources described various data verification measures that were undertaken for the Kerr, Sulphurets, Mitchell, and Iron Cap mineralized zones. Section 4.2(7) of Companion Policy 43-101CP allows a new technical report to include any material information documented in a previously filed technical report, to the extent that this information is still current and relevant. The 2016 PFS is considered by Seabridge and the QPs to be current and relevant, and it has been summarized into this Report by the QPs with minor changes – there have been minor changes to the Measured and Indicated Mineral Resources used in the 2016 PFS, and the text content has been summarized. As part of the data verification process, the QPs reviewed information on the Property generated since the effective date of the 2016 PFS to confirm there has been no material change to the 2016 PFS and that the Mineral Reserves and economic outcomes from the study remain current. The checks taken and the results of this data verification are presented in Section 12.3.

Section 2.3(4) of Companion Policy 43-101 CP allows disclosure of the results of a Preliminary Economic Assessment (PEA) that includes Inferred Mineral Resources after the completion of a Prefeasibility Study that establishes Mineral Reserves. The QPs also have reviewed the key technical and economic parameters used in the 2020 PEA to determine if they have any impact on the 2016 PFS and that the study and the Mineral Reserves remain current. The checks taken and the results of this data verification are presented in Section 12.4.

| 12.1 | Drill Hole Data Verification

| 12.1.1 | Assay Verification

The QP responsible for Mineral Resources for this Report has personally compared assay certificates against the Seabridge drill hole assay database on an annual basis after the 2006 drilling campaign. The entire 2006 gold and copper assay data have been checked with no errors found. In early 2007, the QP responsible for Mineral Resources for this Report undertook an assay verification of some of the older drilling records from the 1991 and 2005 campaigns. The only errors from this check were that the electronic database did not contain some copper values despite certified copper assays recorded on the Eco Tech certificates (2005 drilling). The Placer Dome drill core was assayed at their internal laboratory located in Vancouver BC. The QP responsible for Mineral Resources for this Report randomly selected a number of Placer drill holes and compared the assays stored in Seabridge’s electronic database with the individual drill hole logs that contained the inhouse assay results. No errors were noted. Assay intersections from five Seabridge drill holes were compared with nearby older holes drilled by Placer Dome. Those comparisons showed that in general, the Placer Dome gold and copper data compared reasonably well with the newer Seabridge data with the newer data showing slightly higher grades than the older data.

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Prior to 2014, at least 10% of the assay records for each annual Seabridge drilling campaign were checked. These assay comparisons typically showed error rates of about 0.1%, which is well within industry accepted standards. Since 2014, the QP responsible for Mineral Resources for this Report has conducted a 100% comparison of certified assay records against Seabridge’s electronic drill hole database. These complete checks have not revealed many errors, most of which were associated with re-run samples (e.g. over limit gold assays that were re-run). All errors were brought to Seabridge’s attention and the electronic database corrected. Details regarding various drill assay database reviews were discussed in prior Technical Reports (Lechner 2007; Lechner 2008; Lechner 2009; and Lechner 2010, Tetra Tech 2012, Lechner 2014, and Tetra Tech 2016).

Initially, most of the minor assay database errors were associated with values that had not been updated in the database (e.g. over limit re-runs, etc.) Over the years, the number of errors discovered during the assay database checks have been reduced to essentially zero because of the robustness of the AcQuire database import scripts.

| 12.1.2 | Drill Hole Logs

The QP responsible for Mineral Resources for this Report has completed numerous comparisons between drill hole core and both paper and electronic drill hole logs (lithology, alteration, mineralization, etc.) during site visits conducted in 2006 through 2019. Core recovery and RQD values were randomly checked by the QP during those site visits. It is the QP’s opinion that drill hole logging was completed in a professional manner and fairly represents the geology of the KSM deposits.

| 12.1.3 | Drill Hole Collar Locations

Drill hole collar coordinates at the KSM Property are expressed in terms of NAD83 units. The inherited Kerr and Sulphurets drill hole data were originally located in a local mine grid that was tied to NAD27 datum. Seabridge personnel located nine of the older Placer Dome drill hole collars and surveyed them with their handheld Trimble DGPS instrument in NAD83 units. Both coordinate sets for those nine drill holes were provided to Aero Geometrics who was able to derive a transformation of the historical KSM drill holes to NAD83 datum. Seabridge’s practice has been to survey their diamond core holes using the handheld Trimble DGPS unit. McGladrey & Associates has been used by Seabridge to provide more accurate locations for drill hole collars. Comparisons between Seabridge and McGladrey & Associates hole locations showed in general that there were only sub-metre differences in location which is not considered to be material.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

The terrain in the KSM area is quite steep requiring large wooden drill platforms to be constructed. The drill contractors measure drill depths starting from the platform elevation which can be several meters higher than the actual ground level where the hole actually penetrates the ground. The final drill hole collar locations are trigonometrically calculated using the drill hole inclination and amount of drill hole “stick-up” (distance between the ground and the bottom of the drill platform).

Drill hole collar locations have been compared against surface topography and in general found to fit reasonably well. It is the opinion of the QP responsible for Mineral Resources that are the subject of this Report that the drill hole collar locations are reasonable for estimating Mineral Resources.

| 12.1.4 | Down-hole Surveys

The QP responsible for Mineral Resources for this Report has completed numerous comparisons between the driller’s handwritten down-hole survey records and the Seabridge electronic database. Between the 2006 and 2018 drilling seasons very few discrepancies were discovered. Those discrepancies were researched by Seabridge’s geological staff and appropriate corrections completed. A minor number of down-hole survey records were discarded each year due to suspect azimuth or inclinations versus adjacent survey stations or surveys with high magnetic susceptibility. It is the opinion of the QP responsible for Mineral Resources for this Technical Report that the down-hole surveys are reasonable and suitable for estimating Mineral Resources.

| 12.1.5 | Quality Assurance/Quality Control

The QP responsible for Mineral Resources for this Report has been provided with QA/QC data for KSM since the initial 2006 Seabridge drilling program. This data has been analyzed with the results and conclusions presented in prior Reports. It is the opinion of the QP responsible for Mineral Resources for this Report that Seabridge’s QA/QC procedures have been adequate in assuring that the drill hole data are suitable for estimating Mineral Resources.

| 12.1.6 | Topographic Data

In 2008, McElhanney of Vancouver, BC was contracted to perform an aerial survey, and provide Seabridge with an updated accurate topographic base map of the three deposits and surrounding area. McElhanney obtained the data by conducting a helicopter-borne LiDAR survey.

The new topographic map of the district was provided to Seabridge in the UTM NAD83 coordinate system, which is the standard system for all Government of BC and industry mapping applications. Seabridge contracted Aero Geometrics of Vancouver to translate the KSM drill hole collar locations from NAD27 to NAD83 datum. Aero Geometrics used Sierra Systems Groups Inc. MAPS 3D software to perform the transformation of all collar coordinates. MAPS 3D uses the Canadian National Transformation Versions 1.1 and 2.0 for the transformation.

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The QP responsible for Mineral Resources for this Report and Seabridge noted some discrepancies in the GPS surveyed collar locations and the new LiDAR topographic surface. These differences are believed to be based on:

| ● | the fact that no transform of the Z-coordinate was considered by the Canadian National Transformation software

| ● | the inaccuracy of the initial GPS elevation

| ● | the fact that many of the holes were surveyed immediately below the drill deck and not ground level or “stick-up” differences magnified by steep terrain.

Therefore, the GPS surveyed collar elevations were used.

| 12.1.7 | Bulk Density Data

As a part of their drill core logging and sampling procedures at KSM, Seabridge has routinely collected bulk density data on drill core samples by weighing the air-dried core samples in air and water and then calculating the ratio. A total of 1,865,837,910 and 501 bulk density determinations for the basis for assigning bulk density values for the Kerr, Sulphurets, Mitchell, and Iron Cap deposits, respectively. The QP responsible for Mineral Resources for this Technical Report has reviewed Seabridge’s field procedures for obtaining bulk density determinations during various site visits and checked the mathematical formulas used to determine the bulk density values. The QP responsible for Mineral Resources for this Report believes that Seabridge’s bulk density collection procedure is reasonable and suitable for estimating Mineral Resources.

| 12.2 | Post Model Drilling Results

The four KSM deposits have not all been updated at the same time because annual drilling campaigns typically focused on one or two of the recognized mineralized zones each season resulting in model updates for just those areas. The current models of record which are the basis of Mineral Resources discussed in this Report are as follows: Kerr – drilling through 2016; Sulphurets – drilling through 2018; Mitchell – drilling through 2011; and Iron Cap – drilling through 2018.

Since those models were finalized, Seabridge has completed limited drilling adjacent to or in some cases through Mineral Resource blocks. Only one hole has been drilled at Kerr (2018) since the model was finalized. That shallow hole was drilled in the southern portion of the deposit in an area where primarily Placer Dome drilling exists. In general, the 2018 drill hole confirmed the geometry and location of mineralization but the drill hole gold grade was about 4% lower than the predicted model grade (0.23 versus 0.24 g/t) and the drill hole copper grade was lower than the model predicted (0.39% versus 0.44%). This comparison of 32 composites against the model is not considered to be a material issue.

Twenty-six holes have been drilled in 2019 at the Sulphurets deposit since the resource was formalized with drilling through the 2018 campaign. Some of the assays for the 2019 holes are still pending at the time of this writing. The 2019 Sulphurets drilling has been completed in the upper plate above the Sulphurets Thrust Fault which consists primarily of various metasedimentary units belonging to the Hazelton Group. None of the 2019 drilling has intersected estimated model blocks in the Sulphurets model of record. Several holes came close to narrow zones of estimated block grades in the upper plate but since they did not intersect resource blocks no statistical comparisons can be made. The 2019 drilling suggests that several zones of lower grade mineralization could possibly be modeled in the upper plate but this additional potential mineralization is not considered to be material.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Since the completion of the Mitchell model which was based on using data through 2011, a total of 23 core holes have been drilled within the deposit area. In 2012, 2015, and 2018 a total of two, three, and 18 holes have been drilled, respectively. The majority of the new drilling was conducted in the hanging wall of the deposit where no block grades had been estimated. Several of those holes intersected narrow zones of primarily low-grade mineralization. Only eight of the post-model holes intersected estimated model blocks from the model of record. A total of 145 drill hole composites from the newer drilling intersected model blocks with estimated grades. The drill hole gold grade was found to be 3% lower than the predicted model grade (0.59 g/t versus 0.61 g/t). The drill hole copper grade was 4% higher than the model predicted (0.19% versus 0.18%). The estimated block silver grade was 3% lower than the new drilling. Given that much of the new drilling has been conducted in areas of un-estimated block grades that may generate some potential low-grade mineralization and that there was a reasonable comparison of new drilling that intersected the model of record, it is the opinion of the QP responsible for Mineral Resources that are subject of this Report that the new drilling at the Mitchell deposit would not result in a material change to the Mitchell Mineral Resource.

| 12.3 | Verification that 2016 Prefeasibility Study Remains Current

The 2016 PFS is considered by Seabridge and the QPs to be current and relevant based on the outcome of a data verification process. As part of the data verification process, the QPs reviewed information on the KSM Property generated since the effective date of the 2016 PFS to determine whether there would be a material change to the 2016 PFS, Mineral Reserves and economic outcomes from the study, if 2016 PFS was updated with the latest information. This check is considered necessary if the 2016 PFS is to be included unchanged in the 2020 Technical Report on the KSM Property.

Also, this check on the 2016 PFS is necessary to support the disclosure required by Section 2.3(3) of NI 43-101 when the results of a PEA that includes Inferred Mineral Resources are disclosed at the same time as a PFS that establishes Mineral Reserves on the subject property. The QPs reviewed the key technical and economic parameters used in the 2020 PEA and have determined that they have no material impact on the 2016 PFS and that the 2016 PFS and the Mineral Reserves from the PFS remain current.

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| 12.3.1 | Verification Checks on the PFS Mine Quantities

The verification process for the 2020 Technical Report included comparisons of mine quantity inputs to the 2016 PFS to the most recent information available to determine if there would be a material difference to the outcomes of the 2016 PFS if the information was updated. The inputs that were verified include the following:

| ● | updated Mineral Resource models (December 2019) compared for three of the four KSM deposits. The changes that occurred with the new models were primarily with the Inferred class material which was excluded from the 2016 PFS.

| ● | process recoveries for some of the KSM deposits, and

| ● | updates to assumptions for operating costs, cutoff grades, and drawpoint shut-off NSR values for underground caving mining.

The Measured and Indicated material in the updated Mineral Resource models (December 2019) was analyzed using mining shapes and mine designs in the 2016 PFS, and the resulting tonnes, diluted grades, and metal content were compared to the 2016 PFS Mineral Reserves. The results are shown in Table 12.1.

Table 12.1 Reserve Assessment Variance Check from 2016 PFS to 2020 Information

Area |  | Diluted Grades | Contained Metal
Ore | Au | Cu | Au | Cu
(Mt) | (g/t) | (%) | (Moz) | (Mlb)
Mitchell 2019 vs. PFS | 0.0% | 0.6% | 1.0% | 0.6% | 0.9%
Kerr 2019 vs. PFS | -3.2% | 1.3% | 2.2% | -1.9% | -1.0%
Sulphurets 2019 vs. PFS | 14.3% | 1.5% | 1.0% | 16.0% | 15.5%
Mitchell Underground | -1.5% | -0.2% | 3.3% | -1.7% | 1.7%
Iron Cap Underground | 0.0% | -10.4% | 10.7% | -10.4% | 10.8%
Total Proven & Probable | 1.3% | 0.0% | 2.4% | 1.3% | 3.6%

Comments on the Reserve variance data checks follows:

| ● | the gold and copper content of the 2016 Mineral Reserves make up 96% of the revenue in the 2016 PFS, so they are the only relevant revenue metals to compare in the analysis

| ● | Mitchell open pit makes up the majority of the mill feed for the first 35 years of the mine life, and the percentage differences in the grade and contained metal are less than 2%, which is well within measurement error, and would not materially affect the outcome of the 2016 PFS

| ● | the differences in the Iron Cap underground Au and Cu grade and metal content offset each other to some extent

| ● | the overall differences indicated in Table 12.1 are relatively small, are well within expected measuring errors, and would not be considered to have a material impact on the outcome of the 2016 PFS

| ● | the comparisons are based on the most recent Mineral Resource models (December 2019) and are constrained within the 2016 PFS mining shapes. Although the mine designs are not updated to reflect potential optimizations, they are technically achievable, and any reserves generated from them and used in the resultant economic analysis will reflect an achievable economic result which in turn allows a corresponding Mineral Reserves to be relied upon

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| ● | because the changes to the 2016 PFS Mineral Reserves are not material, no change was made to the production schedule, process plant, and supporting infrastructure for the economic analysis verification checks.

In conclusion, the Mineral Reserves in the 2016 PFS would not be expected to materially change if the Mineral Reserves were updated using the December 2019 Mineral Resource models. Therefore the 2016 Mineral Reserves are considered current and suitable to be used in the 2016 PFS that is included in the 2020 Technical Report on the KSM Property.

| 12.3.2 | Verification Checks on Financial Analysis and Cost Estimates

The verification process for the 2020 Technical Report included comparisons of costs and financial inputs to the 2016 PFS to the most recent information available to determine if there would be a material difference to the outcomes of the 2016 PFS if the information was updated. The inputs that were verified include the following:

| ● | capital costs

| ● | operating costs

| ● | metal pricing, and

| ● | exchange rate and tax changes since 2016.

The results of updating the financial model with the updated cost and financial inputs is shown in comparison to the 2016 PFS in Table 12.2.

Table 12.2 Summary of the 2016 PFS Verification Checks

Item | Unit | 2016 PFS (Base Case) | 2020 Verification Check | Variance
NPV5 | US$M | 3,263 | 3,596 | 333
IRR | % | 10.4 | 10.6 | 0.2
Payback | years | 6 | 5.8 | -0.2

Comments on the cost and financial input data verification checks follows:

| ● | checks to the capital cost estimates showed the initial and sustaining capital costs in the 2016 PFS, if re-estimated today, would likely increase by 3.8% and 3.2% respectively, on a pre-tax basis. This expected difference is well within the precision expected from a prefeasibility mine capital cost estimate and if updated, the capital cost figures would not represent a material change to the 2016 PFS. The 2016 PFS mine capital cost estimates are considered to be current and suitable to be used unchanged in the 2016 PFS that will be summarized in the 2020 Technical Report on the KSM property

| ● | the checks on operating cost estimates show the operating costs, if re-estimated today, would be within 2.12% of the 2016 PFS estimates. These changes are well within the precision expected from a prefeasibility operating cost estimate and if updated, would not represent a material change to the 2016 PFS. Therefore, the 2016 PFS operating cost estimates are considered to be current and suitable to continue to be used unchanged in the 2016 PFS that will be summarized in the 2020 Technical Report on the KSM property

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| ● | checks on the commodity prices and exchange rate show that any updates to these parameters would not cause a material change to the economic analysis in the 2016 PFS. The commodity prices and exchange rate used in the 2016 PFS are considered to be current and suitable to continue to be used in the 2020 Technical Report on the KSM property

| ● | the checks on changes to the tax rates show that any update to these parameters would not cause a material change to the economic analysis in the 2016 PFS. The net tax applied to the financial model in the 2016 PFS are considered to be current and suitable to continue to be used unchanged in the 2020 Technical Report on the KSM property.

In conclusion, the economic analysis of the 2016 PFS would not be expected to materially change if the capital and operating costs, commodity prices and exchange rate were updated. Therefore the 2016 capital and operating costs, commodity prices and exchange rates are considered current and suitable to be used unchanged in the 2016 PFS that is to be included in the 2020 Technical Report on the KSM Property.

| 12.4 | QP’s Opinion

| 12.4.1 | Drill Hole Data Verification

Based on the QPs’ review of the various vintages of data that were used to estimate Mineral Resources for the KSM deposits, the drill hole assay, survey, and geologic data were found to be professionally collected and are thought to be adequate for estimating Mineral Resources.

| 12.4.2 | 2016 Prefeasibility Study Data Verification

Based on the QPs’ review of the new information generated on the KSM property since the completion of the 2016 PFS, the new information does not materially change the outcome of the 2016 PFS and therefore the 2016 PFS is considered to be current and suitable to be included unchanged in this Report.

| 12.4.3 | 2020 PEA vs 2016 PFS

The 2020 PEA is a standalone, alternative development option that does not invalidate the 2016 PFS which remains as a viable development option.

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| 13.0 | Mineral Processing and Metallurgical Testing

| 13.1 | Introduction

The Property includes four major mineralized zones, identified as the Mitchell, Kerr (upper Kerr and Deep Kerr), Sulphurets, and Iron Cap (upper Iron Cap and lower Iron Cap) deposits. Mineralization in all of the zones includes gold, copper, silver, and molybdenum.

Principal sulphides in all the KSM zones are pyrite and chalcopyrite. Lower portions of the Kerr and Iron Cap deposits contain minor secondary copper minerals.

Extensive metallurgical test programs have been carried out to determine the metallurgical characterization of the deposits (Table 13.1) culminating in the 2016 KSM PFS and 2016 PEA (Tetra Tech, 2016)

Most of the materials in the mineralization zones have generally higher than typical pyrite-to-chalcopyrite ratios. Pyrite is a significant gold carrier, and copper-to-gold ratios vary significantly by deposit.

Table 13.1 Typical Mineralogical Characteristics and Average Copper-to-Gold Grade Ratios in Planned Mill Feed

Deposit | Average Pyrite/Chalcopyrite Ratio* | Average Copper Grade, %** | Average Gold Grade, g/t** | Copper/Gold Grade Ratio**
Mitchell | 12 | 0.20 | 0.77 | 0.26
Sulphurets | n/a | 0.31 | 0.73 | 0.42
Iron Cap | 8 | 0.35 | 0.58 | 0.61
Kerr*** | 5 | 0.49 | 0.31 | 1.59

* Approximate figures

** 2020 PEA average mill feed grades

*** Deep Kerr only

In general, the average Bond ball mill work index of the tested KSM samples range from 13.9 to 15.7 kWh/t that are of moderate hardness to ball mill grinding, excluding the Sulphurets mineralization, which is much harder, with an average Bond ball mill work index of 18.5 kWh/t. Table 13.2 shows the average Bond ball mill work index and abrasion index for the four mineralization deposits.

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Table 13.2 Average Ball Mill Grindability and Abrasion Index

Deposit | Average BMWi, kWh/tonne | Average Bond Abrasion Index, g
Mitchell | 14.4 | 0.29
Sulphurets | 18.5 | 0.23
Iron Cap | 15.7 | 0.10
Deep Kerr | 14.3 | 0.06

The mineral samples from the Mitchell Zone were tested in the earliest test programs starting in 2007 to develop the process flowsheet for the KSM process design, including flotation for copper, gold, silver and molybdenum recovery and cyanide leach on the gold-bearing pyrite tailings rejected from copper and gold bulk flotation for additional gold and silver recovery to produce doré. The flowsheet had been further optimized and then tested on the mineral samples from the other zones to investigate amenability of these samples to the flowsheet, in particular the copper and gold recovery by flotation. With the verification testing of the other mineral zone sample amenability to the flowsheet developed, the developed flowsheet formed the metal recovery basis of the KSM processing plant.

Since 2016 KSM PFS and PEA, additional metallurgical test programs have been conducted on the expanded underground components of the Kerr and Iron Cap deposits. The mineral samples tested showed similar metallurgical responses to the samples tested from the other deposits and mineralization zones. The metallurgical performance projections for the Deep Kerr and Iron Cap mineralization have been updated according to the updated test results produced.

Table 13.3 summarizes the flotation locked cycle test results from various samples and test programs.

Table 13.3 Locked Cycle Flotation Test Result Summary

| Mitchell | Upper Kerr (Before 2012) | Iron Cap (Before 2015)
Min | Max | Average | Min | Max | Average | Min | Max | Average
Head Grade
- Copper, % Cu | 0.12 | 0.24 | 0.20 | 0.59 | 0.69 | 0.62 | 0.14 | 0.38 | 0.25
- Gold, g/t Au | 0.55 | 0.92 | 0.73 | 0.22 | 0.25 | 0.24 | 0.24 | 1.28 | 0.57
Cu-Au Concentrate
- Grade, % Cu | 20.2 | 30.1 | 24.9 | 22.3 | 30.7 | 27.8 | 22.6 | 26.7 | 24.6
- Recovery, % Cu | 71.5 | 89.3 | 84.5 | 80.6 | 86.3 | 83.0 | 81.4 | 88.1 | 85.1
- Recovery, % Au | 43.2 | 73.5 | 60.8 | 37.7 | 49.7 | 41.8 | 45.0 | 71.4 | 60.4
table continues…

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| Sulphurets | Deep Kerr 2013–2017 | Iron Cap Lower (2017–2020)
Min | Max | Average | Min | Max | Average | Min | Max | Average
Head Grade
- Copper, % Cu | 0.16 | 0.46 | 0.26 | 0.26 | 1.83 | 0.64 | 0.20 | 1.03 | 0.56
- Gold, g/t Au | 0.50 | 0.70 | 0.59 | 0.21 | 0.94 | 0.46 | 0.03 | 0.59 | 0.35
Cu-Au Concentrate
- Grade, % Cu | 22.7 | 29.3 | 26.9 | 21.0 | 28.7 | 25.2 | 19.2 | 28.5 | 24.8
- Recovery, % Cu | 60.6 | 85.7 | 78.0 | 81.5 | 96.6 | 89.7 | 82.0 | 93.6 | 90.1
- Recovery, % Au | 40.1 | 58.6 | 52.9 | 50.5 | 77.1 | 62.3 | 46.8 | 74.8 | 62.5

The gold and silver extractions by cyanide leach from the gold-bearing pyrite tailings (first cleaner scavenger tailings and pyrite concentrate) rejected from copper and gold bulk flotation varied significantly. On average, the additional gold recovered into the doré is expected to be 11 % to 17%, while the additional silver recovery to the doré is estimated to be 13% to 16%. The test results suggest that the gold and silver recovery of the gold-bearing sulphide products from the Deep Kerr and lower Iron Cap deposits did not seem to respond well to the gold recovery by the established cyanide leaching treatment, compared to the samples from Mitchell and Sulphurets deposits. Further test work is required to optimize the flowsheet for additional gold and silver recovery.

| 13.2 | Summary of Metallurgical Test Programs

Since 1989, extensive metallurgical test programs have been carried out to determine the metallurgical characterization of the deposits (Table 13.4). Test work from 1989 to 2016 inclusive of all the deposits established the baseline metallurgical recovery method presented in 2016 PFS. The most recent tests between 2016 and 2020 have been conducted to assess the established metallurgical recovery method with major samples from Deep Kerr and Iron Cap deposits, including additional cyanidation tests on Mitchell samples.

The following sections summarize the test work prior to 2016 with a more detailed review on the most recent test programs.

Table 13.4 Metallurgical Test Work Programs

Report Year | Program ID | Lab | Mineralogy | Flotation/ Cyanide Leach | Grindability | Others
2020 | KM6004 | ALS | ✓ | ✓ | ✓ | ✓
2019 | KM5806 | ALS | ✓ | ✓ | ✓ | ✓
2018 | KM5501 | ALS |  |  |  | ✓
2018 | KM5501 | ALS | ✓ | ✓ | ✓ | ✓
2018 | KM5455 | ALS |  | ✓ |  |
2017 | KM5367 | ALS |  | ✓ |  | ✓
|  |  |  |  | table continues…

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Report Year | Program ID | Lab | Mineralogy | Flotation/ Cyanide Leach | Grindability | Others
2017 | KM5266 | ALS | ✓ | ✓ |  | ✓
2017 | KM5248 | ALS | ✓ | ✓ |  | ✓
2017 | KM5204 | ALS |  | ✓ |  |
2017 | KM5063 | ALS | ✓ | ✓ | ✓ | ✓
2016 | KM5087 | ALS |  | ✓ |  |
2016 | SSW30216SD | SSW | ✓ |  |  |
2015 | KM4514 | ALS | ✓ | ✓ | ✓ | ✓
2015 | KM4672 | ALS | ✓ | ✓ | ✓ | ✓
2015 | - | Pocock |  |  |  | ✓
2014 | KM4029 | ALS | ✓ | ✓ | ✓ | ✓
2013 | KM3735 | ALS | ✓ | ✓ |  |
2013 | 12628-002 | SGS |  |  |  | ✓
2012 | KM3174 | G&T |  | ✓ | ✓ |
2012 | KM3080 | G&T |  | ✓ |  | ✓
2011 | KM3081 | G&T |  | ✓ |  |
2011 | KM 2897 | G&T |  | ✓ |  |
2011 | SSW47110 | SSW | ✓ |  |  |
2010/2011 | KM 2748 | G&T | ✓ | ✓ | ✓ | ✓
2010 | KM 2755 | G&T | ✓ | ✓ |  | ✓
2010 | KM 2670 | G&T | ✓ | ✓ |  |
2009/2010 | KM 2535 | G&T |  | ✓ | ✓ |
2009/2010 | 12157-001 | SGS |  | ✓ | ✓ |
2009/2010 | 12248-001 | SGS |  | ✓ | ✓ |
2009/2010 | #20001153 | Metso |  |  | ✓ |
2009/2010 | KMA913snR | Köeppern |  |  | ✓ |
2009 | KM 2344 | G&T | ✓ | ✓ | ✓ | ✓
2009 | - | Pocock |  |  |  | ✓
2008 | KM 2153 | G&T | ✓ | ✓ | ✓ | ✓
2008 | - | Hazen |  |  | ✓ |
2007 | KM 1909 | G&T | ✓ | ✓ | ✓ | ✓
1991 | - | PDRC |  | ✓ | ✓ | ✓
1990 | - | PDRC | ✓ | ✓ | ✓ | ✓
1989 | - | BMML |  | ✓ | ✓ |
1989 | - | CRI |  | ✓ |  |

ALS – ALS Metallurgy Kamloops

BMML – Brenda Mines Ltd. Metallurgical Laboratory

CRI – Coastech Research Inc.

G&T – G&T Metallurgical Services Ltd. (now ALS Metallurgy Pty Ltd.)

Köeppern – Köeppern Machinery Australia Pty Ltd.’s HPGR Pilot Plant at UBC

Metso – Metso Minerals Industries Inc.

PDRC – Placer Dome Research Centre

Pocock – Pocock Industrial Inc.

SGS – SGS Mineral Services

SSW– Surface Science Western

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| 13.3 | Summary of Initial Test Work 1989–1991

The metallurgical responses of Kerr mineral samples were assessed by initial metallurgical test programs by Coastech and Placer Dome between 1989 to 1991. The copper head grade varied from 0.40% to 1.30% Cu. The associated gold and silver concentrations were 0.26 g/t and 1.21 g/t Au and 0.9 g/t to 3 g/t Ag, respectively.

The historical test work involved mineralogy, grindability and metallurgical responses to flotation concentration process; the major conclusions are listed as follows:

| ● | preliminary mineralogy analysis was carried out and identified sericite in 3 of the 4 tested samples

| ● | indicative work index was determined by using a comparative method, which indicates that the tested samples were soft to intermediate hardness materials

| ● | preliminary open circuit flotation tests produced differing metallurgical upgrading responses to the test conditions. The optimum results were obtained by Placer Dome in 1991, which produced salable copper concentrates with a certain level of gold content. However, it was reported that significant gold had been lost to the flotation tailing.

| 13.4 | Summary of Test Work 2007–2016

The 2007 to 2016 metallurgical tests were performed on mineralized samples from Mitchell, Sulphurets, upper and Deep Kerr, and Iron Cap deposits.

| 13.4.1 | Test Programs

The following test programs were performed to establish a metallurgical flowsheet and to optimize process-related parameters:

| ● | mineralogy, flotation, cyanidation, and grindability test work by G&T Metallurgical Services Ltd. (G&T) and SGS Minerals Services (SGS)

| ● | metallurgical variability responses and copper-molybdenum separation techniques were also investigated. Flotation locked cycle tests were performed on the composite samples from all the deposits, particularly on a variety of samples from the Mitchell deposit. Cyanidation tests were conducted to further recover gold and silver from the gold-bearing sulphide streams (scavenger cleaner tailing from the copper-gold bulk flotation and pyrite concentrate)

| ● | semi-autogenous grinding (SAG) mill comminution (SMC) grindability tests to determine the grinding resistance of the mineralization to SAG/ball milling by Hazen Research Inc. (Hazen) and G&T

| ● | crushing resistance parameters to high-pressure grinding rolls (HPGR) crushing of the Mitchell and Sulphurets ore samples by SGS, and pilot plant scale HPGR testing on the Mitchell ore sample by Köeppern Machinery Australia Pty Ltd.’s (Köeppern) HPGR pilot plant at UBC

| ● | dewatering tests by Pocock Industrial Inc. (Pocock) on various product pulp samples.

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| 13.4.2 | Baseline Test Process Flowsheet and Conditions

The test results indicate that the mineral samples from the 4 separate mineralized deposits are all amenable to the flotation-cyanidation combined process. The process flowsheet consists of the following stages:

| ● | copper-gold-molybdenum bulk rougher flotation followed by gold-bearing pyrite flotation

| ● | regrinding the bulk rougher concentrate followed by 3 stages of cleaner flotation to produce a copper-gold-molybdenum bulk cleaner flotation concentrate

| ● | molybdenum separation of the bulk cleaner flotation concentrates to produce a molybdenum concentrate and a copper/gold concentrate containing associated silver if molybdenum content in the bulk concentrate is economical for the separation

| ● | cyanide leaching of the gold-bearing pyrite flotation concentrate and the scavenger cleaner tailing to further recover gold and silver values as doré bullion.

The flotation reagents used in the testing were 3418A (dithiophosphinates), A208 (dithiophosphate), and fuel oil for copper-gold-molybdenum bulk flotation, and A208 and potassium amyl xanthate (PAX) for gold-bearing pyrite flotation. The primary grind size used was 80% passing approximately 125 µm to 150 µm, and concentrate regrind size was 80% passing approximately 20 µm.

| 13.4.3 | Mitchell Zone Major Metallurgical Test Results

Mitchell Zone samples were tested by G&T, SGS, Meso, and Hazen from 2007 to 2012 in various metallurgical programs. All samples were originally collected/prepared from diamond drill core samples in respective drilling campaigns.

Mitchell Samples Characteristics

Mitchell samples were examined for chemical compositions, mineralogy, and crushability/grindability associated with using HPGR, SAG mill, ball mill, tower mill, and IsaMill technologies. Comminution circuit simulations were performed to develop a baseline crushing and grinding circuit to reduce the particle size of the Mitchell mineralized materials to the target grind size.

Chemical Composition and Mineralogy

The chemical compositions of the Mitchell Zone samples are summarized in Table 13.5. The copper grade is between 0.07% and 0.71% Cu, with an average level of 0.21% Cu; the gold content varies from 0.01 g/t to 1.49 g/t Au, with an average value of 0.79 g/t Au; the silver content varies from 1 g/t to 18 g/t Ag, with an average value of 4 g/t Ag; and the molybdenum content averages approximately 0.007% Mo.

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Table 13.5 Test Samples – Mitchell (2007–2012)

Year/Test Program | Sample Quantity | Cu (%) | Au (g/t) | Ag (g/t) | Mo (%)
Min | Max | Avg. | Min | Max | Avg. | Min | Max | Avg. | Min | Max | Avg.
2007/KM1909 | 3 | 0.19 | 0.22 | 0.20 | 0.86 | 0.90 | 0.87 | 3 | 4 | 4 | n/a | n/a | n/a
2008/KM2153 | 34 | 0.07 | 0.52 | 0.23 | 0.22 | 1.49 | 0.79 | 1 | 18 | 4 | 0.001 | 0.025 | 0.005
2009/KM2344 | 10 | 0.10 | 0.27 | 0.22 | 0.60 | 1.08 | 0.93 | 3 | 8 | 4 | 0.001 | 0.006 | 0.007
2009/KM2535 | 11 | 0.10 | 0.71 | 0.24 | 0.35 | 1.02 | 0.69 | 2 | 9 | 4 | 0.002 | 0.015 | 0.008
2010/KM2670 | 3 | 0.12 | 0.20 | 0.16 | 0.01 | 0.79 | 0.65 | 2 | 3 | 3 | 0.006 | 0.013 | 0.010
2012/KM3174 | 25 | 0.12 | 0.30 | 0.20 | 0.56 | 1.10 | 0.76 | 2 | 5 | 3 | 0.002 | 0.015 | 0.005
Overall | 0.07 | 0.71 | 0.21 | 0.01 | 1.49 | 0.79 | 1 | 18 | 4 | 0.001 | 0.025 | 0.007

The dominate copper mineral identified in the Mitchell samples is chalcopyrite; the main sulphide mineral is pyrite, which was present as approximately 6% to 8% of the sample weight. The degree of chalcopyrite liberation ranged from 46% to 56% across the samples tested at a primary grind size of 80% passing 116 µm to 136 µm.

Crushability/Grindability – Bond Ball Mill Work Index

Standard Bond ball mill work index tests were determined in several programs on the Mitchell mineralization. The results indicate that the Bond work indices range from 12.5 kWh/t to 15.5 kWh/t, averaging 14.4 kWh/t. This suggests that the Mitchell samples are of moderate hardness. The Bond abrasion index (Ai) of Composite PP1 was measured at 0.293 g.

Crushability/Grindability – SAG Mill Comminution (SMC) Index

SMC grindability tests indicate that the Mitchell samples are moderately resistant to SAG mill grinding with A x b values from 30.0 to 59.6. Based on these results, JK SimMet simulations were conducted for a comminution circuit of 120,000 t/d, 92% availability, and a feed particle size of 80% passing 150 mm. Three scenarios were simulated with different BWi and product size levels as represented in Table 13.6.

| ● | simulation 1: Bond ball mill work index 14.8 kWh/t, a product particle size of 80% passing 150 μm

| ● | simulation 2: Bond ball mill work index 16.0 kWh/t, a product particle size of 80% passing 150 μm

| ● | simulation 3: Bond ball mill work index 15.0 kWh/t, a product particle size of 80% passing 120 μm.

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Table 13.6 JK SimMet Simulation Results (60,000 t/d SABC Circuit, 2008)

Simulation | 1a | 1b | 2a | 2b | 3a | 3b
SAG Mill | Size, D x L (EGL) (ft x ft) | 40 x 24 | 37.7 x 21 | 40 x 24 | 37.7 x 21 | 40 x 24 | 37.7 x 21
Circulation Load (% of Feed) | 19.5 | 18.4 | 19.5 | 18.4 | 19.5 | 18.4
Gross Power Draw (kW) | 18,843 | 15,570 | 18,843 | 15,570 | 18,843 | 15,570
Transfer Particle Size, µm | 2,500 | 3,035 | 2,500 | 3,035 | 2,500 | 3,035
Ball Mills | Size, D x L (EGL) (ft x ft) | 22 x 36 | 22 x 36 | 22 x 36 | 22 x 36 | 22 x 36 | 24 x 38
Mill Number | 2 | 2 | 2 | 2 | 2 | 2
Gross Power Draw (kW) | 15,644 | 17,293 | 16,912 | 18,695 | 19,283 | 21,017
Product Size P80, µm | 150 | 150 | 150 | 150 | 120 | 120
Total Power Draw (kW) | 34,487 | 32,863 | 35,755 | 34,265 | 38,126 | 36,587
Cyclone Diameter (inches) | 26 | 26 | 26 | 26 | 26 | 26

Crushability/Grindability – HPGR

Bench scale HPGR test work was conducted by SGS on both the Michell and Sulphurets samples, followed by a pilot plant scale test at Köeppern’s HPGR pilot plant at UBC. The test results from both programs indicate the following:

| ● | the bench scale LABWAL tests by SGS showed that both Mitchell and Sulphurets materials are amenable to HPGR crushing. On average, the net specific energy requirement is 2.33 kWh/t for the Mitchell sample and 3.08 kWh/t for the Sulphurets sample. The Sulphurets mineralization is more resistant to HPGR crushing than the Mitchell mineralization

| ● | a lower net specific energy consumption (approximately 1.94 kWh/t) was recorded from the closed circuit pilot plant tests, in comparison with 1.99 kWh/t obtained from the single pass tests. A specific pressing force of 4 N/mm2 was considered to be optimum on the basis of both size reduction and throughput performance. An increase in feed moisture is expected to result in a reduction in throughput and an increase in energy consumption.

Crushability/Grindability – Tower Mills/Isa Mills

Metso Minerals Industries Inc. (Metso) investigated the specific energy consumption for secondary grinding, using tower mills in the jar mill tests. The mill feed particle size was 80% passing 173 µm, and the product particle size was 125 µm. From the test results, Metso projected the specific energy requirement by a stirred tower mill would be approximately 0.88 kWh/t.

SGS used the IsaMill™ procedure to determine the specific energy requirement for regrinding the gold-bearing pyrite rougher concentrate produced from the Mitchell samples. The specific energy requirement to regrind the concentrate from 80% passing 66 µm to 80% passing 16 µm was determined to be about 24.2 kWh/t. The grinding media consumption was estimated to be 6 g/kWh.

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Mitchell Cu-Mo Bulk Concentrate Flotation Results

Open Cycle Tests

In the 2008 testing program, 32 drill core interval composite samples were used for variability tests, excluding two samples (Met 35 and Met 36) from the Sulphurets Zone. Primary grind sizes ranged from 80% passing 115 µm to 171 µm, averaging 149 µm. The rougher concentrates from the copper circuit were reground to approximately 80% passing 18 µm prior to cleaner flotation. The variation in the copper and gold flotation performance of various Mitchell mineral samples is shown in Figure 13.1. The results indicate some variability in copper and gold metallurgical responses, generally related to head grade changes.

Figure 13.1 Copper and Gold Open Cycle Flotation Variability Test Results (KM2153)

Source: KM 2153, G & T Metallurgical Services Ltd., Kamloops, BC, Canada, September 16, 2008

As shown in Figure 13.2, G&T established the relationship between copper recovery and copper feed grade at a fixed cleaner concentrates grade of 25% copper. This figure indicates that feed grade is the main factor causing copper performance variation.

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Figure 13.2 Copper Open Cycle Flotation Performance vs Copper Head Grade at a Concentrate Grade of 25% Copper (KM2153)

Source: KM 2153, G & T Metallurgical Services Ltd., Kamloops, BC, Canada, September 16, 2008

Besides the gold reporting into the final copper concentrate, as shown in Figure 13.1, additional gold recoveries to the gold-bearing pyrite concentrate from the pyrite flotation circuit averaged approximately 16%. Combined gold recoveries from both the copper-gold flotation circuit and gold-bearing pyrite flotation circuit ranged from 73% to 96%, averaging approximately 86%.

Additional open batch cleaner flotation tests were conducted on nine composite samples representing the major Mitchell Zone mineralization types projected to be mined during various operating periods of an open pit mine plan. Some of the head samples contain higher than the average planned mill feed grades, especially for gold. The tests were conducted at primary grind sizes, averaging 80% passing 119 µm and regrind sizes of 80% passing 18 µm. The test results show that the average flotation metal recovery was 84.6% for copper and 61.2% for gold from these nine Mitchell metallurgical composites. The test results are shown in Figure 13.3.

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Figure 13.3 Flotation Performance – Open Circuit Flotation Tests, Mitchell (KM2153)

Source: KM 2153, G & T Metallurgical Services Ltd., Kamloops, BC, Canada, September 16, 2008

The seven composites produced concentrates higher than 25% copper except for 16.2% concentrate copper grade from the low grade IARG 0-10 composite and 24.0% concentrate copper grade from the QSP LG 0-10 composite.

The relationship between the adjusted copper recovery and copper feed grade is plotted in Figure 13.4 with a copper recovery adjustment to reflect a concentrate grade of 25% copper. The graph for the batch flotation tests indicates a good correlation of copper recovery with copper head grade.

Figure 13.4 Copper Recovery vs. Copper Feed – Open Cleaner Circuit Tests (KM2153)

Source: KM 2153, G & T Metallurgical Services Ltd., Kamloops, BC, Canada, September 16, 2008

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Although some metallurgical performance variation was observed, the Mitchell mineral samples responded well to the conventional flotation.

Locked Cycle Tests

Fifteen locked cycle tests (LCTs) were conducted on the various Mitchell composite samples. The test results are summarized in Table 13.7 for the Mitchell mineralization and in Table 13.8 for Mitchell mineralization samples blended with the other mineralization. The LCTs showed that:

| ● | a substantial variation in the concentrate grade, from 20% to 30% copper, was noticed for individual Mitchell samples. On average, the final copper concentrate contained approximately 25% copper. The average recoveries to the concentrate were 84.7% for copper, 61% for gold, 50% for silver, and 56% for molybdenum

| ● | approximately 26% of the gold and 28% of the silver in the feed reported to other gold-bearing products, which can be further extracted by cyanide leaching

| ● | for the blended samples, the metallurgical performance appeared comparable to that produced when treating the Mitchell material on its own.

G&T conducted pilot tests on Mitchell samples in 2009 that produced lower metal recoveries and concentrate grades compared with the LCTs results. According to G&T, this was caused by the issues with the pilot plant control or circuit stability. The pilot test results were considered as reference data only.

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Table 13.7 Locked Cycle Test Results – Mitchell

Test Program* | Comp | Grind Size (P80 µm**) | Feed Grade | Bulk Concentrate Grade | Flotation Recovery
Cu (%) | Au (g/t) | Ag (g/t) | Mo (%) | Cu (%) | Au (g/t) | Ag (g/t) | Mo (%) | Cu (%) | Au (%) | Ag (%) | Mo (%)
G&T 2153/141 | Master | 119/16 | 0.21 | 0.89 | 4.2 | - | 20.2 | 62.8 | 273 | - | 87.8 | 63.0 | 58.5 | -
G&T 2153/142 | Master | 119/17 | 0.21 | 0.92 | 3.7 | - | 22.0 | 64.7 | 242 | - | 87.0 | 58.5 | 52.5 | -
G&T 2344/73 | PP Comp 1 | 103/14 | 0.24 | 0.81 | - | - | 22.3 | 55.7 | - | - | 89.3 | 66.2 | - | -
G&T 2535/18 | PP Comp 1 | 103/16 |  |  |  |  | 28.0 | 77.8 | 260 | - | 87.2 | 67.4 | 47.0 | -
G&T 2535/20 | PP Comp 1 | 137/17 | 0.24 | 0.82 | 3.9 | - | 23.8 | 62.0 | 248 | - | 88.1 | 66.2 | 55.6 | -
G&T 2670/12 | PP Comp 3 | 147/15 | 0.20 | 0.74 | 3.2 | 0.006 | 30.1 | 77.7 | 264 | 0.39 | 84.2 | 58.0 | 52.6 | 35.7
G&T 2670/18 | PP Comp 3 | 147/22 | 0.20 | 0.79 | 3.2 | 0.006 | 27.4 | 70.5 | 272 | 0.46 | 86.1 | 56.5 | 53.0 | 49.7
G&T 2670/22 | PP Hi Mo | 143/21 | 0.16 | 0.60 | 3.3 | 0.014 | 22.4 | 61.7 | 243 | 1.20 | 78.9 | 56.9 | 43.8 | 47.9
G&T 2670/26 | BS Hi Mo | 143/17 | 0.12 | 0.55 | 2.4 | 0.010 | 24.9 | 70.3 | 185 | 1.26 | 71.5 | 43.2 | 26.0 | 42.2
G&T 2897/01 | Comp 46 of KM2344 | 120/16 | 0.15 | 0.65 | 2.3 | 0.012 | 22.6 | 80.5 | 226 | 1.76 | 89.1 | 73.5 | 58.6 | 86.3
G&T 3081/93 | Mitchell 3081-M2# | 137/18 | 0.20 | 0.65 | 4.7 | 0.004 | 27.1 | 58 | 427 | 0.33 | 83.3 | 56.3 | 57.1 | 55.6
G&T 3081/82 | Mitchell 3081-M3# | 123/22 | 0.21 | 0.57 | 3.5 | 0.004 | 23.8 | 44.2 | 223 | 0.24 | 88.2 | 59.9 | 49.5 | 49.2
G&T 3081/103 | Mitchell 3081-M3# | 123/17 | 0.22 | 0.55 | 4.0 | 0.006 | 29.8 | 56 | 299 | 0.27 | 76.7 | 57.8 | 43.0 | 26.3
ALS 4514/30*** | Mitchell 4515-M1# | 133/15/15 | 0.21 | 0.90 | 5 | 0.006 | 26.7 | 98.2 | 431 | 0.72 | 81.6 | 68.6 | 53.5 | 70.0
SGS | PP Comp 1 | 129/28 | 0.21 | 0.72 | - | 0.005 | 23.1 | 53.7 | - | 0.41 | 89.0 | 59.6 | - | 65.0

| Notes: | *Au grade in the Bulk Cleaner Tailings ranged from 1.01 g/t to 2.12 g/t Au, averaging at 1.67 g/t Au; Au grade in the pyrite concentrate ranged from 0.58 g/t to 2.26 g/t Au, averaging at 1.48 g/t Au.

**Primary grind size/regrind size
***Including a copper flotation on the pyrite flotation concentrate

|  | #Composite samples IDs based on the previous mine plans (planned in 2010/2011) have been relabeled: Mitchell M1 for Mitchell Yr 0-5;

Mitchell M2 for Mitchell Yr 0-10; Mitchell M3 for Mitchell Yr 0-20;

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Table 13.8 Locked Cycle Test Results – Blended Samples (Mitchell and Other Deposits)

Test Program* | Comp | Grind Size (P80 µm**) | Head Grade | Bulk Concentrate Grade | Flotation Recovery
Cu (%) | Au (g/t) | Ag (g/t) | Mo (%) | Cu (%) | Au (g/t) | Ag (g/t) | Mo (%) | Cu (%) | Au (%) | Ag (%) | Mo (%)
G&T 2535/19 | Mitchell (PP Comp1)/ Kerr (52/53 Blend); 80%:20% | 127/20 | 0.31 | 0.70 | 3.5 | - | 25.3 | 40.0 | 168 | - | 87.4 | 60.4 | 51.4 | -
G&T2670/62 | Mitchell/Sulphurets Blend; 60%:40% | 141/22 | 0.22 | 0.67 | 2.8 | 0.007 | 24.2 | 52.0 | 178 | 0.664 | 85.9 | 59.8 | 50.9 | 72.4
G&T 2748/18 | Mitchell (PP Comp 1)/ Iron Cap C1/Iron Cap C2; 33%:33%:33% | 135/15 | 0.24 | 0.79 | - | 0.004 | 27.6 | 59.6 | - | 0.250 | 87.8 | 58.2 | - | 51.5
ALS 4672/32*** | Mitchell (Mitchell M1#)/ Iron Cap (IC-2014-MC4) | 117/17/17 | 0.24 | 0.67 | 4 | 0.005 | 25.0 | 54.3 | 304 | 0.430 | 82.7 | 65.7 | 58.4 | 70.3
ALS 4514/31*** | Mitchell (Mitchell M1#)/ Kerr(DK-2014-MC3) | 129/17/17 | 0.37 | 0.59 | 3 | 0.006 | 24.5 | 28.3 | 150 | 0.328 | 87.9 | 62.8 | 57.2 | 75.1

| Notes: | *Au grade in the bulk cleaner tailing ranged from 1.07 g/t to 1.89 g/t Au, averaging at 1.45 g/t Au; Au grade in the pyrite concentrate ranged from to 0.41 g/t to 1.85 g/t Au, averaging at 1.15 g/t Au.

**Primary grind size/regrind size
***Including a copper flotation on the pyrite flotation concentrate

#Sample ID relabeled, see note in Table 13.7

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Mitchell Copper-Gold and molybdenum Separation Flotation Tests

Copper-gold and molybdenum separation flotation tests were performed to further recover molybdenum concentrate from copper-gold-molybdenum bulk concentrates in the 2009/2010 testing program. The results indicate that:

| ● | the best results were obtained from the bulk Cu-Mo concentrates produced in the 2010 pilot plant tests in an open circuit flotation test. It produced a 51% molybdenum concentrate with a molybdenum recovery of 72%

| ● | the molybdenum-copper separation LCT recovered 88.5% of the molybdenum from the molybdenum-copper concentrate and produced a 41% Mo concentrate. The test results are provided in Table 13.9.

Table 13.9 Cu-Mo Separation LCT Results, 2010

Product | Weight (%) | Grade (%) | Recovery (%)
Cu | Mo | C | Cu | Mo
Bulk Concentrate | 100.0 | 19.3 | 1.28 | 0.63 | 100.0 | 100.0
Mo Concentrate | 2.8 | 2.66 | 41.2 | 5.76 | 0.4 | 88.5
Cu Concentrate | 97.2 | 19.8 | 0.15 | 0.48 | 99.6 | 11.5

Furthermore, G&T conducted preliminary leaching tests on the molybdenum concentrates to reduce the contained copper and lead concentrations. The tests used both the Brenda-leach procedure and hydrochloric acid leaching methods. The Brenda-leach method can lower the copper and lead contents from 2.06% to 0.26% for copper and from 0.14% to 0.03% for lead. The hydrochloric acid leaching method is less efficient which reduced the copper content from 1.50% to 0.81%.

The assay on the final molybdenum concentrates indicated that the concentrates contained approximately 2,200 g/t rhenium (Re).

Mitchell Cyanide Leach Tests

Cyanide leach tests on the gold-bearing products (first cleaner scavenger tailing and pyrite concentrate) generated from the flotation tests on Mitchell samples were mainly conducted during 2007 and 2016. Additional tests have been conducted after 2017, on the samples generated during the 2010 pilot plant campaigns, to evaluate cyanidation and other potential treatments.

Mitchell Cyanide Leach Tests 2007–2016

Cyanide leach tests were conducted on products from the open circuit flotation tests, locked cycle flotation tests, and pilot plant tests. The tests mainly used Carbon-In-Leach (CIL) procedure while some of the tests used Direct Cyanidation (DCN) procedure. CIL procedure was used for the cyanidation tests on the pilot flotation samples. The combined first cleaner tailing and the gold-pyrite concentrate were reground and subjected to cyanide leaching for additional gold and silver recovery.

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For leaching on the open circuit flotation products, the average gold extraction rate in the 24-hour leach tests in 2008 ranged from 63% to 91%, with an approximate average of 79%. For leaching of the LCT products, the average extraction rate is 66% for gold and 56% for silver (Table 13.10). The lower gold leach extraction rate appears to be related to some difficult leach gold-bearing middling materials rejected into the first cleaner tailings and also a lower gold head grade.

Cyanide leaching tests on the pilot plant products exhibited notable variabilities between G&T and SGS. The gold extraction by G&T was from 70% to 78%, while SGS obtained a lower gold extraction of 50% to 59%. The lower gold extraction rate by SGS is possibly due to the leach testing being focused to generate leach solutions for cyanide destruction testing program.

Some of the leaching tests were conducted separately on the reground first cleaner tailing and reground pyrite concentrate. The results indicate that the first cleaner tailing produced lower gold extraction than the pyrite concentrate.

Table 13.10 Cyanidation Test Results on LCT Products – Mitchell

Testing Program | Sample | Regrind Size (P80 µm) | Feed (Au, g/t) | Extraction (Au, %) | Feed (Ag, g/t) | Extraction (Ag, %)
G&T-2153 | Master | 15 | 1.8 | 67.6 | 9.1 | 62.1
G&T-2153 | Master | 15 | 2.2 | 73.2 | 10.1 | 64.4
G&T-2344 | PP Comp 1 | 12 | 1.6 | 68.0 |  |
G&T-2535 | PP Comp 1 | 15 | 1.7 | 69.0 | 12.6 | 54.4
G&T-2535 | PP Comp 1 | 15 | 1.6 | 81.1 | 10.9 | 54.7
G&T-2670 | PP Comp 3 | 21 | 1.6 | 61.6 |  |
G&T-2670 | PP Comp 3 | 18 | 2.0 | 66.5 | 8.1 | 55.5
G&T-2670 | PP Hi Mo | 19 | 1.9 | 68.0 | 8.6 | 50.6
G&T-2670 | BS Hi Mo | 19 | 1.7 | 68.9 | 7.6 | 48.7
G&T-2897 | Comp 46 |  | 1.1 | 63.5 |  |
G&T-3081 | Mitchell M2* | 24 | 1.5 | 51.2 |  |
G&T-3081 | Mitchell M3* | 21 | 1.2 | 50.4 |  |
SGS | PP Comp 1 | 16 | 1.1 | 69.8 |  |
Average – Mitchell | 18 | 1.6 | 66.1 | 9.6 | 55.8

*Sample ID relabeled, see note in Table 13.7

Mitchell Cyanide Leach Tests 2017–2020

Additional cyanidation test work has been conducted by ALS Metallurgy in 2017 (KM5367) and 2018 (KM5455) on Mitchell samples. Test KM5367 was to investigate gold, silver, and copper extractions with lower sodium cyanide concentration at varied leach time. The leach feed samples were treated in a single-stage copper flotation prior to leaching.

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2017 Samples

The bulk cleaner tailing and pyrite rougher concentrate samples produced in the previous pilot plant testing of the program KM2344 and KM2670 were prepared for the leaching test. PP samples were the composited samples from Test 2670, while P7 samples were from Test 2344. Details for sample composition can be found in the Appendix I of the original test report (KM5367). Table 13.11 shows the chemical compositions of these samples.

Table 13.11 Metal Contents of Leach Test Head Samples – Mitchell, 2017 (ALS KM5367)

Leach Feed Samples | Metal Content
Cu (%) | S (%) | Au (g/t) | Ag (g/t) | Fe (%)
PP Bulk Cleaner Tailing (KM2670) | 0.14 | 25.4 | 1.45 | 5.0 | n/a
P7 Bulk Cleaner Tailing (KM2344) | 0.16 | 39.2 | 2.09 | 7.8 | 33.5
PP Pyrite Rougher Concentrate (KM2670) | 0.14 | 27.7 | 2.44 | 6.1 | n/a
P7 Pyrite Rougher Concentrate (KM2344) | 0.25 | 10.5 | 1.35 | 5.0 | 9.7

2018 Samples

The samples used in the 2018 cyanidation tests were the composited samples of the P7 cleaner tailing from KM2344 and PP bulk cleaner tailing from KM2670. The composite samples were ground to 80% passing 14 µm. The head assay results of the composite sample are presented in Table 13.12.

Table 13.12 Metal Contents of Leach Test Head Samples – Mitchell, 2018 (ALS KM5367)

Leach Feed Samples | Metal Content
Cu (%) | S (%) | Au (g/t) | Ag (g/t)
Reground Mitchell Bulk Cleaner Tailing | 0.14 | 32.0 | 1.75 | 8.0

Cyanidation Optimization Test (KM5367)

DCN and CIL tests were conducted on bulk cleaner tailing and pyrite concentrate samples from previous tests. Sodium concentration level, leaching time, and the effects of residual copper minerals were investigated to assess the extraction of gold, silver, and copper.

The cyanidation test results are presented in Table 13.13. The following major observations are made from the leaching test results:

| ● | reducing the initial cyanide concentration from 1,000 ppm to 500 ppm generally led to lower extractions of gold, silver, and copper, especially for pyrite concentrate samples

| ● | longer leaching time was beneficial for silver extraction. Compared with direct leaching process, silver extraction rates were improved significantly by using CIL test procedure

| ● | similar to the 2007–2016 test results, lime consumption for bulk cleaner tailing was significantly higher than pyrite rougher concentrate samples.

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Table 13.13 Optimization Cyanidation Test Results – Mitchell 2017 (ALS)

Test | Head Samples | Test Type | Leach Time (h) | Initial NaCN Concentration (ppm) | Cyanide Consumption (g/t) | Lime Consumption (g/t) | Feed Grade, Calculated | Extraction
Au (g/t) | Ag (g/t) | Cu (%) | Au (%) | Ag (%) | Cu (%)
KM5367-1 | PP Bulk Cleaner Tailing | DCN | 8 | 1,000 | 2,000 | 3,100 | 1.40 | 7.7 | 0.11 | 49.4 | 55.9 | 45.2
KM5367-5 | 500 | 1,400 | 2,000 | 1.60 | 6.34 | 0.12 | 55.6 | 38.5 | 36.4
KM5367-2 | PP Pyrite Ro. Concentrate | DCN | 8 | 1,000 | 1,500 | 2,900 | 2.34 | 6.99 | 0.13 | 77.0 | 58.5 | 28.7
KM5367-6 | 500 | 1,100 | 1,800 | 2.65 | 6.03 | 0.14 | 66.4 | 48.6 | 22.4
KM5367-3 | P7 Cleaner Tailing | DCN | 8 | 1,000 | 2,200 | 10,900 | 2.22 | 8.62 | 0.14 | 52.6 | 49.0 | 41.1
KM5367-7 | 500 | 2,000 | 6,600 | 2.29 | 7.63 | 0.15 | 51.6 | 34.5 | 41.6
KM5367-4 | P7 Pyrite Concentrate | DCN | 8 | 1,000 | 1,800 | 4,200 | 1.43 | 6.27 | 0.25 | 79.8 | 31.5 | 20.1
KM5367-8 | 500 | 1,400 | 1,800 | 1.71 | 5.16 | 0.27 | 57.8 | 12.8 | 16.4
KM5367-9 | P7 Cleaner Tailing | CIL | 10 | 1,000 | 3,000 | 6,100 | 2.26 | 8.81 | 0.15 | 52.6 | 68.2 | 48.4
KM5367-10 | 500 | 2,300 | 6,100 | 2.23 | 8.59 | 0.15 | 51.2 | 65.1 | 43.9
KM5367-11 | P7 Pyrite Concentrate | CIL | 10 | 1,000 | 2,200 | 1,500 | 1.42 | 5.75 | 0.26 | 73.3 | 66.9 | 17.9
KM5367-12 | 500 | 1,600 | 1,500 | 1.53 | 5.34 | 0.26 | 62.7 | 51.3 | 15.1
KM5367-15 | Combined PP Bulk Cleaner Tails/P7 Cleaner Tailing | CIL | 10 | 1,000 | 2,400 | 3,500 | 1.82 | 7.42 | 0.13 | 47.1 | 58.2 | 41.4
KM5367-16 | 500 | 2,100 | 3,500 | 1.84 | 7.14 | 0.14 | 45.9 | 55.2 | 38.7
KM5367-17 | Combined PP Pyrite Ro. Concentrate/P7 Pyrite Concentrate | CIL | 10 | 1,000 | 2,100 | 1,500 | 1.89 | 6.59 | 0.20 | 68.8 | 60.5 | 19.3
KM5367-18 | Combined PP Pyrite Ro. Concentrate/P7 Pyrite Concentrate | CIL | 10 | 500 | 1,500 | 1,500 | 1.95 | 5.81 | 0.20 | 62.0 | 55.2 | 16.8
KM5367-19 | Test 13 1st Cleaner Tailing/ Test 14 1st Cleaner Tailing* | CIL | 10 | 1,000 | 1,000 | 5,000 | 0.98 | 3.60 | 0.04 | 47.6 | 61.1 | 25.9
KM5367-20 | Test 13 1st Cleaner Tailing/ Test 14 1st Cleaner Tailing* | CIL | 10 | 500 | 500 | 4,400 | 1.32 | 3.38 | 0.05 | 48.5 | 43.8 | 18.3

Note: *Test 13 and Test 14 cleaner tails were produced from the copper flotation of the combined PP bulk rougher concentrate/P7 pyrite concentrate.

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Cyanidation Optimization Test (KM5455)

In this testing program, the leach feed was also subjected to one single-stage copper flotation prior to cyanidation. The test results are shown in Figure 13.5. The following major observations were made:

| ● | similar to the 2017 test results, the copper flotation prior to leaching process was beneficial for reducing cyanide consumption

| ● | finer grind size feed to the cyanide leaching can improve the extractions of gold, silver, and copper. The increased extractions were 68.1% Au, 73.4% Ag, and 66.2% Cu (KM5455-03) as compared with the previous results of 47.1% Au, 58.2% Ag, and 41.4% Cu (KM5367-15). However, this increase may also result from the longer leaching retention time

| ● | an increase in pre-aeration from 2 hours to 16 hours had only a marginal increase of gold extraction by 1%.

Figure 13.5 Leaching Test Results – Mitchell 2017 and 2018 (ALS KM5455)

| Source: | ALS Metallurgy (2018). Metallurgical Testing of a Mitchell Bulk Cleaner Scavenger Tailings Sample (KM 5455). June 19, 2018.

| 13.4.4 | Sulphurets Zone Major Metallurgical Test Results

G&T tested the Sulphurets Zone samples in 2008 and 2011/2012. The 3 composite samples from crushed drill cores were tested to investigate the metallurgical responses of Sulphurets mineralization.

Sulphurets Samples Characteristics

Chemical Composition and Mineralogy

The chemical compositions of the Sulphurets Zone samples are summarized in Table 13.14. The copper grade is between 0.14% and 0.46% Cu with an average of 0.28% Cu. The gold content varies from 0.26 g/t to 0.81 g/t Au, averaging 0.61 g/t Au. The concentrations of silver are relatively constant with an average of 1 g/t Ag. The molybdenum content averages approximately 0.006% Mo.

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Table 13.14 Test Samples – Sulphurets (2009-2012)

Year/Test Program | Sample Quantity | Cu (%) | Au (g/t) | Ag (g/t) | Mo (%)
Min | Max | Avg | Min | Max | Avg | Min | Max | Avg | Min | Max | Avg
2009 (G&T) | 3 | 0.14 | 0.37 | 0.25 | 0.26 | 0.81 | 0.53 | 1 | 2 | 1 | 0.003 | 0.011 | 0.007
2011/2012(G&T) | 2 | 0.17 | 0.46 | 0.32 | 0.65 | 0.76 | 0.70 | 1 | 2 | 1 | 0.004 | 0.008 | 0.006
Overall | 0.14 | 0.46 | 0.28 | 0.26 | 0.81 | 0.61 | 1 | 2 | 1 | 0.003 | 0.011 | 0.006

Crushability/Grindability – Bond Ball Mill Work Index

The Bond ball work test results indicate that the Sulphurets samples are more resistant to ball mill grinding compared to the Mitchell samples. The average Bond ball work index is 18.5 kWh/t; the Bond Ai of the overall Sulphurets composite is 0.233 g.

Crushability/Grindability – SMC Index

The SMC grindability tests indicate that the Sulphurets samples are more resistant to SAG mill grinding than the Mitchell samples. The A x b values of the tested Sulphurets and Mitchell samples were 41.7 and 38.7, respectively.

Crushability/Grindability – HPGR

SGS conducted bench scale HPGR tests on the Sulphurets composite samples. The tests included batch open circuit tests and LCTs. The test results indicate that the Sulphurets mineralization is more resistant to HPGR crushing than the Mitchell mineralization. On average, the net specific energy requirement determined by bench tests is 3.08 kWh/t for the Sulphurets sample, compared to 2.33 kWh/t for the Mitchell sample.

The preliminary HPGR/ball mill circuit simulation results by SGS suggested that the unit power requirement for the HPGR/ball mill circuit would be approximately 14.8 kWh/t for the Sulphurets mineralization, compared to 10.4 kWh/t for the Mitchell mineralization.

Sulphurets Cu-Mo Bulk Concentrate Flotation Results

Open cycle cleaner flotation and locked cycle flotation tests were conducted on Sulphurets samples to test metallurgical responses of the samples to the process flowsheet developed for the Mitchell mineralization. The open cycle flotation tests by both G&T and SGS indicate that the Sulphurets samples show good metallurgical performance and may produce higher grade copper concentrates than the Mitchell samples.

Several LCTs have been conducted on the various Sulphurets composite samples. The test results are summarized in Table 13.15:

| ● | SGS test recovered 85.7% copper to the bulk concentrate graded at 22.7% Cu, while lower copper recovery but higher copper grade bulk concentrates were produced from the G&T testing program. The testing on Composite 9 produced a much lower copper recovery compared to the other tests, which could be the result of the low head grade

| ● | the average gold recovery for both tests was approximately 56%, excluding the lower gold recovery of Composite 9. Silver recovery in the G&T tests averaged 30%. Molybdenum to the bulk concentrate averaged 69%.

The locked cycle flotation test results for the metallurgical performances of the Mitchell-Sulphurets blend sample (60% Mitchell and 40% Sulphurets) are presented in Table 13.8 of the Mitchell mineralization section. Similar performance was observed between the blended samples and Mitchell samples.

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Table 13.15 LCT Results – Sulphurets

Test Program* | Composite | Grind Size (P80 µm**) | Head Grade | Bulk Conc. Grade | Flotation Recovery (%)
Cu (%) | Au (g/t) | Ag (g/t) | Mo (%) | Cu (%) | Au (g/t) | Ag (g/t) | Mo (%) | Cu | Au | Ag | Mo
SGS | Composite | 125/20 | 0.20 | 0.66 | - | 0.007 | 22.7 | 49.1 | - | 0.630 | 85.7 | 56.1 | - | 66.6
G&T 2670/44 | Master Composite (Comp49/50/51) | 154/16 | 0.24 | 0.52 | 1.6 | 0.006 | 28.3 | 41.8 | 82.0 | 0.701 | 80.5 | 53.9 | 34.3 | 72.2
G&T 2897/22 | Master Composite (Comp49/50/51) | 113/- | 0.24 | 0.50 | 1.5 | 0.008 | 28.4 | 41.6 | 71.4 | 0.850 | 79.4 | 55.6 | 31.5 | 68.5
G&T 3174/8 | Composite 8 | 121/19 | 0.46 | 0.70 | 1 | 0.008 | 29.3 | 31.4 | 34 | 0.227 | 83.6 | 58.6 | 31.1 | 37.7
G&T 3174/9 | Composite 9 | 127/21 | 0.16 | 0.59 | 2 | 0.004 | 26.0 | 63.7 | 130 | 0.170 | 60.6 | 40.1 | 21.3 | 14.1

| Notes: | * Au grade in the Bulk Cleaner Tailings ranged from 1.82 g/t to 3.55 g/t Au, averaging at 2.38 g/t Au; Au grade in the pyrite concentrate ranged from 0.67 g/t to 1.41 g/t Au, averaging at 1.16 g/t Au.

** Primary grind size/regrind size

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Sulphurets Cyanidation Leach Results

The first cleaner tailing and the gold-pyrite concentrate from the varied flotation circuit were subjected to cyanide leaching for additional gold and silver recovery. Both G&T and SGS conducted cyanidation tests on the products from the locked cycle flotation tests.

The test results are provided in Table 13.16. In general, the Sulphurets samples produced lower gold and silver extractions, in comparison with the Mitchell samples. The best gold extraction obtained was 70.5% by SGS using the CIL leach procedure. The direct cyanide leach test produced inferior results.

Table 13.16 Cyanidation Test Results – Flotation LCT Products, Sulphurets, 2009–2011

Test Program | Sample | Regrind Size (P80 µm) | Feed (Au g/t) | Extraction (Au %) | Feed (Ag g/t) | Extraction (Ag %)
G&T-2670 | Master Composite | 16 | 1.7 | 40.9 | 3.7 | 52.4
G&T 2897 | Master Composite | - | 1.5 | 34.5 | 3.3 | 47.9
Composite 8 – 2011/2012 | Raewyn CV | 25 | 1.6 | 41.7 |  |
Composite 9 – 2011/2012 | Lower Hazelton | 19 | 2.5 | 68.3 |  |
SGS (DCN) | Composite | - | 1.6 | 51.5 | - | -
SGS (CIL) | Composite | - | 1.3 | 70.5 | - | -

| 13.4.5 | Upper Kerr Zone Metallurgical Test Results

There are two mineralization zones in the Kerr deposit. The Deep Kerr Zone mineralization may be mined by underground block caving, while the upper Kerr Zone material will likely be mined by open pit mining.

Early test work from 2010 to 2012 completed by G&T were focused on the samples from the surface Kerr Zone. Between 2013 and 2017, 5 test programs—KM3735, KM4514, KM4029, KM5063, and part of KM5266—were completed to investigate the metallurgical performance of the mineralization from the Deep Kerr Zone.

Upper Kerr Samples Characteristics

Four composite samples from the upper Kerr were prepared for metallurgical testing from the drill core intervals. It appears that the upper zone materials may have experienced some oxidation and contained more clay minerals.

Chemical Composition

The assay data of the composites are presented in Table 13.17. The mineralization of each sample can be found in the 2016 PFS report.

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Table 13.17 Metal Contents of Composites – Upper Kerr, 2010 (G&T)

Composite | Mineralization Type | Metal Content
Cu (%) | Au (g/t) | Mo (%) | Ag (g/t)
Comp 52 – 2010 | Rubble Zone | 0.59 | 0.22 | 0.004 | 2.0
Comp 53 – 2010 | Quartz Stockwork | 0.61 | 0.17 | 0.001 | 1.5
Composite 10 – 2011/2012 | CL Quartz Crackle | 0.59 | 0.26 | 0.001 | 1.0
Composite 11 – 2011/2012 | QSP Quartz Crackle | 0.68 | 0.29 | 0.001 | 2.0

Crushability/Grindability – Ball Mill

The samples from the upper Kerr Zone are more amenable to ball mill grinding when compared to the Mitchell and Sulphurets mineralization. The average Bond ball mill work index is 13.9 kWh/t.

Crushability/Grindability – SAG Mill

The 2011/2012 testing program determined the grindability of the upper Kerr samples to SAG mill milling. The test results revealed that the grindability of the upper Kerr samples to SAG mill grinding is very similar to the samples from the Mitchell deposit. The A x b values of the 2 tested samples are very close for upper Kerr and Mitchell, 46.1 and 47.0, respectively.

Upper Kerr Cu-Mo Bulk Concentrate Flotation Results

Upper Kerr samples were tested using the same flotation flowsheet developed with Mitchell and Sulphurets samples.

The open circuit batch flotation tests showed that the upper Kerr samples produced better concentrate grades than the Mitchell or Sulphurets samples. Copper recovery produced was slightly lower than the Mitchell or Sulphurets samples at the equivalent copper concentrate tenor. Gold recovery for the upper Kerr samples was lower because the gold head grades were considerably lower than the samples from the Mitchell and Sulphurets deposits.

The LCT results, as presented in Table 13.18, indicate that the metallurgical performance of the upper Kerr samples was not as good as that achieved with the Mitchell and Sulphurets samples, despite their lower copper head grades. The inferior metallurgical performance of the upper zone materials may be caused by oxidation and more clay mineral contents.

On average, the upper Kerr samples produced a 27.8% copper concentrate. The copper and gold reporting to the concentrate were 83% and 41%, respectively. Approximately 51% of the gold reported to the gold-bearing pyrite products (first cleaner tailing and gold-bearing pyrite concentrate). The 2011/2012 test program produced better metallurgical performance from the samples tested than what had been achieved previously.

The flotation LCT results for the metallurgical performance of the Mitchell-Kerr blend sample (80% Mitchell and 20% upper Kerr) are presented in Table 13.8 of the Mitchell mineralization section. Similar performance was observed between the blended samples and Mitchell samples.

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Table 13.18 LCT Results – Upper Kerr (G&T)

Test Program* | Comp | Grind Size (P80 µm**) | Head Grade | Bulk Conc. Grade | Flotation Recovery
Cu (%) | Au (g/t) | Ag (g/t) | Cu (%) | Au (g/t) | Ag (g/t) | Cu (%) | Au (%) | Ag (%)
G&T 2535/16 | Comp 52 | 119/15 | 0.59 | 0.22 | 1.9 | 22.3 | 4.05 | 33.5 | 81.6 | 38.8 | 37.6
G&T 2535/17 | Comp 53 | 122/14 | 0.62 | 0.25 | 1.4 | 29.3 | 5.58 | 31.8 | 80.6 | 37.7 | 37.9
G&T 3174/10 | Composite 10 | 124/18 | 0.59 | 0.24 | 2 | 30.7 | 7.2 | 49 | 86.3 | 49.7 | 39.8
G&T 3174/11 | Composite 11 | 130/19 | 0.69 | 0.24 | 3 | 29.0 | 5.1 | 77 | 83.4 | 41.1 | 47.4

| Note: | *Au grade in the Bulk Cleaner Tailings ranged from 0.51 to 0.97 g/t Au, averaging at 0.70 g/t Au; Au grade in the pyrite concentrate ranged from 0.38 to 0.66 g/t Au, averaging at 0.55 g/t Au.

**Primary grind size/regrind size

Cyanidation Leach Results

The first cleaner tailing and the gold-pyrite concentrate from the varied flotation circuit were subjected to cyanide leaching for additional gold and silver recovery. G&T conducted cyanidation tests on the products produced from the locked cycle flotation tests; CIL procedure was used for the leaching process. Test results are provided in Table 13.19.

On average, gold extraction from both the gold-bearing products was approximately 57%, slightly lower than the results obtained from the Mitchell samples. The average gold feed grade to the cyanide leach circuit was lower in comparison with the cyanide leach feeds of the Mitchell samples. The test results also indicated that the first cleaner tailing produced slightly lower gold and silver recoveries compared to the gold-bearing pyrite concentrate. The average silver extraction was 32%, which was lower than the average extraction of 56% obtained from the Mitchell samples.

Table 13.19 Cyanidation Test Results on LCT Products – Upper Kerr (G&T)

Test Program | Sample | Regrind Size (P80 µm) | Feed (Au g/t) | Extraction (Au %) | Feed (Ag g/t) | Extraction (Ag %)
G&T 2535/G&T 2897 | Comp 52 | 17 | 1.1 | 76.0 | 5.5 | 45.8
G&T 2535/G&T 2897 | Comp 53 | 15 | 0.6 | 59.7 | 3.2 | 18.7
G&T 3174 | Composite 10 | 20 | 0.6 | 47.2 |  |
G&T 3174 | Composite 11 | 20 | 0.6 | 45.6 |  |
Average – Upper Kerr | 18 | 0.7 | 57.1 | 4.4 | 32.3

| 13.4.6 | Deep Kerr Zone Major Metallurgical Test Results

Five test programs were completed by ALS Metallurgy from 2013 to 2017 with drill core samples from the Deep Kerr Zone. KM3735, KM4029-B, and KM4514 tests were completed between 2013 and 2015. KM5063 and KM5266 has been completed in 2017 with some of the test results reported in the 2016 PFS report, which are summarized in Table 13.20. The rest of the work will be discussed in the present test work review of this report.

Deep Kerr Samples Characteristics

Composite samples from the Deep Kerr Zone were prepared for metallurgical testing from the drill core intervals.

Chemical Composition and Mineralogy

The metal assays of the individual and master composites samples are summarized in Table 13.20 and Table 13.21. With individual samples, the copper concentration varied from 0.18% to 1.75% Cu, averaging at 0.56% Cu. The master samples presented less variations of the copper grade: between 0.25% and 0.67% Cu with an average of 0.47% Cu.

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Table 13.20 Individual Test Samples – Deep Kerr (2012–2016)

Year/Test Program | Sample Quantity | Cu (%) | Au (g/t) | Ag (g/t) | Mo (%)
Min | Max | Ave | Min | Max | Ave | Min | Max | Ave | Min | Max | Ave
2012/KM3735 | 1 | 0.92 | 0.32 | 5 | 0.004
2013/KM4029 | 6 | 0.41 | 1.75 | 0.89 | 0.27 | 1.04 | 0.62 | 1 | 4 | 3 | 0.004 | 0.008 | 0.006
2014/KM4514 | 12 | 0.25 | 0.86 | 0.52 | 0.12 | 0.76 | 0.34 | 1 | 4 | 2 | 0.002 | 0.009 | 0.005
2016/KM5063 | 17 | 0.18 | 0.86 | 0.51 | 0.24 | 0.58 | 0.40 | 0.3 | 4 | 2 | 0.001 | 0.005 | 0.003
2016/KM5266 | 6 | 0.22 | 0.67 | 0.40 | 0.25 | 0.50 | 0.42 | 2 | 4 | 3 | 0.001 | 0.003 | 0.002
Overall | 0.18 | 1.75 | 0.56 | 0.12 | 1.04 | 0.41 | 0.3 | 5 | 2 | 0.001 | 0.009 | 0.004

Table 13.21 Master Test Samples – Deep Kerr (2012–2016)

Year/Test Program | Sample Quantity | Cu (%) | Au (g/t) | Ag (g/t) | Mo (%)
Min | Max | Ave | Min | Max | Ave | Min | Max | Ave | Min | Max | Ave
2014/KM4514 | 5 | 0.46 | 0.54 | 0.50 | 0.20 | 0.44 | 0.31 | 1 | 3 | 2 | 0.003 | 0.006 | 0.004
2016/KM5063 | 13 | 0.25 | 0.67 | 0.47 | 0.30 | 0.84 | 0.40 | 0.8 | 3 | 2 | 0.002 | 0.006 | 0.003
2016/KM5266 | 2 | 0.25 | 0.51 | 0.38 | 0.35 | 0.43 | 0.39 | 3 | 4 | 4 | 0.002 | 0.003 | 0.003
Overall | 0.25 | 0.67 | 0.47 | 0.20 | 0.84 | 0.37 | 1 | 4 | 2 | 0.002 | 0.006 | 0.003

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The mineralogical composition of some of the Deep Kerr head samples indicated that the primary copper sulphide mineral was chalcopyrite with a high pyrite content between 5.5% and 10.9%.

Grindability

The Deep Kerr Zone master samples exhibited moderately soft to average hardness to ball mill grinding process. The average Bond ball mill work index is 13.9 kWh/t, ranging from 13.2 to 15.0 kWh/t.

Deep Kerr Cu-Mo Bulk Concentrate Flotation Results

Deep Kerr samples were tested using the same flotation flowsheet developed with the Mitchell, Sulphurets, and upper Kerr samples. In addition to batch open circuit cleaner flotation tests, flotation LCTs were completed, with the results shown in Table 13.22.

The LCT results show that the Deep Kerr mineralizations responded well to the tested flow sheet. The copper recovery ranged from 86% to 97%, and the gold recovery varied from 56% to 77%. The flotation copper concentrate grades were between 22.8% and 28.7%.

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Table 13.22 Flotation LCT Results – Deep Kerr

Test | Sample | Grind Size (P80 µm**) | Head Grade | Bulk Conc. Grade | Recovery
Cu (%) | Au (g/t) | Ag (g/t) | Mo (%) | Cu (%) | Au (g/t) | Ag (g/t) | Mo (%) | Cu (%) | Au (%) | Ag (%) | Mo (%)
KM4029-22 | DK2013-01 | 118/15/15 | 0.59 | 0.36 | 1.9 | 0.007 | 25.3 | 11.5 | 36.5 | 0.243 | 92.7 | 69.2 | 41.0 | 78.1
KM4029-23 | DK2013-02 | 119/15/11 | 0.44 | 0.23 | 2.0 | 0.004 | 24.4 | 9.0 | 48.0 | 0.157 | 90.6 | 62.6 | 40.0 | 67.4
KM4029-18 | DK2013-03 | 121/14/10 | 0.50 | 0.94 | 4.0 | 0.007 | 23.0 | 32.9 | 136 | 0.215 | 89.7 | 68.2 | 65.3 | 59.4
KM4029-17 | DK2013-04 | 116/16/- | 0.86 | 0.50 | 2.0 | 0.003 | 26.3 | 9.5 | 46.0 | 0.068 | 91.2 | 55.8 | 69.9 | 58.6
KM4029-16 | DK2013-04 | 116/15/11 | 0.82 | 0.50 | 2.4 | 0.004 | 26.3 | 10.6 | 48.0 | 0.071 | 93.6 | 61.5 | 58.3 | 54.8
KM4029-26* | DK2013-05 | 127/23/10 | 1.83 | 0.93 | 3.3 | 0.002 | 28.7 | 11.2 | 43.5 | 0.010 | 96.6 | 74.7 | 80.5 | 26.4
KM4029-24* | DK2013-05 | 127/30/11 | 1.81 | 0.93 | 3.1 | 0.002 | 22.8 | 9.4 | 33.8 | 0.006 | 96.6 | 77.1 | 82.6 | 29.6
KM4029-25 | DK2013-06 | 128/15/9 | 1.44 | 0.67 | 3.4 | 0.004 | 28.4 | 9.3 | 50.0 | 0.039 | 92.7 | 65.5 | 69.2 | 50.4
KM4514-23 | DK-2014-MC3 | 124/16/15 | 0.52 | 0.32 | 2 | 0.004 | 23.5 | 10.2 | 55 | 0.164 | 91.4 | 65.9 | 58.5 | 76.4
KM4514-24 | DK-2014-MC1 | 121/16/18 | 0.55 | 0.42 | 2 | 0.003 | 24.7 | 14.8 | 36 | 0.091 | 91.3 | 72.6 | 48.7 | 62.8
KM4514-25 | DK-2014-MC2 | 126/16/16 | 0.51 | 0.21 | 3 | 0.005 | 26.4 | 7.3 | 74 | 0.210 | 86.1 | 56.9 | 46.3 | 67.1
KM4514-26 | DK-2014-MC4 | 115/16/16 | 0.57 | 0.47 | 1 | 0.003 | 27.4 | 18.6 | 34 | 0.104 | 89.5 | 73.8 | 48.4 | 67.1
KM4514-27 | DK-2014-MC5 | 128/16/16 | 0.50 | 0.24 | 3 | 0.006 | 26.9 | 8.8 | 99 | 0.266 | 88.5 | 59.1 | 63.9 | 74.7
ALS 4514/31*** | Mitchell (Mitchell M1#)/Kerr (DK-2014-MC3) | 129/17/17 | 0.37 | 0.59 | 3 | 0.006 | 24.5 | 28.3 | 150 | 0.328 | 87.9 | 62.8 | 57.2 | 75.1

| Notes: | *Au grade in the Bulk Cleaner Tailings ranged from 0.29 to 1.07 g/t Au, averaging at 0.64 g/t Au; Au grade in the pyrite concentrate ranged from 0.11 to 0.66 g/t Au, averaging at 0.27 g/t Au.

**Primary grind size/regrind size

***Repeat tests, the cleaner flotation for Test 26 was conducted on more diluted slurry (using a larger flotation cell)

#Sample ID relabeled, see note in Table 13.7

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Cyanidation Leach Results

ALS Metallurgy conducted cyanidation tests on the first cleaner tailing and the gold-bearing pyrite concentrate produced from the flotation LCTs. Test results are presented in Table 13.23. The gold extractions by cyanide leaching (CIL procedure) fluctuated from 57% to 73% for the bulk cleaner scavenger tailing and from 64% to 78% for the gold-bearing pyrite concentrates.

Table 13.23 Preliminary Cyanidation Test Results – Deep Kerr

Test | Sample | Extraction*
Gold (%) | Silver (%)
KM4029-29 | Test 22,23 Bulk Cleaner Scavenger Tailing | 72.6 | 86.2
KM4029-30 | Test 22,23 Gold-Bearing Pyrite Concentrate | 77.2 | 90.8
KM4029-31 | Test 16,18 Bulk Cleaner Scavenger Tailing | 59.1 | 76.5
KM4029-32 | Test 16,18 Gold-Bearing Pyrite Concentrate | 73.9 | 91.2
KM4029-33 | Test 24,25 Bulk Cleaner Scavenger Tailing | 57.5 | 73.1
KM4029-34 | Test 24,25 Gold-Bearing Pyrite Concentrate | 72.2 | n/a
KM4514-33 | Bulk Cleaner Scavenger Tailing | 64.2 | 82.6
KM4514-34 | Gold-Bearing Pyrite Concentrate | 77.8 | 70.8
KM4514-35 | Bulk Cleaner Scavenger Tailing | 68.1 | 56.9
KM4514-36 | Gold-Bearing Pyrite Concentrate | 72.9 | 78.9
KM4514-37 | Bulk Cleaner Scavenger Tailing | 66.3 | 78.2
KM4514-38 | Gold-Bearing Pyrite Concentrate | 68.4 | 54.3
KM4514-39 | Bulk Cleaner Scavenger Tailing | 58.2 | 75.7
KM4514-40 | Gold-Bearing Pyrite Concentrate | 63.7 | 72.1
KM4514-41 | Bulk Cleaner Scavenger Tailing** | 56.6 | 71.9
KM4514-42 | Gold-Bearing Pyrite Concentrate** | 67.1 | 71.8

| Notes: | * Cyanide concentration: 1,000 ppm; pH: 11; carbon addition: 28 g/L ** Mitchell (Mitchell Year 0-5)/ Kerr (DK-2014-MC3)

The mineralogical study by Surface Science Western (SSW) on the leaching residues from Test KM4514-41 found that the residual gold is present in colloidal type sub-microscopic gold, mainly in pyrite, which occurs in coarse and porous types. SSW pointed out that the pre-treatment by pressure or bio-oxidation would be required to release this locked gold.

| 13.4.7 | Iron Cap Zone Major Metallurgical Test Results

Previous test programs on Iron Cap were conducted in 2010 and between 2014 to 2015. The 2010 test work was based on two composite samples, while further tests in 2014 and 2015 were performed using samples from lower Iron Cap zone.

Iron Cap Samples Characteristics

Chemical Composition and Mineralogy

The assays of the head samples from the Iron Cap zone are summarized in Table 13.24 and Table 13.25. The average copper and gold grades of the individual Iron Cap samples were 0.29% Cu and 0.54 g/t Au, which is similar with the master sample results.

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Table 13.24 Individual Test Samples – Iron Cap (2010–2015)

Year/Test Program | Sample Quantity | Cu (%) | Au (g/t) | Ag (g/t) | Mo (%)
Min | Max | Avg | Min | Max | Avg | Min | Max | Avg | Min | Max | Avg
2013/2014 | 3 | 0.22 | 0.25 | 0.24 | 0.28 | 0.59 | 0.45 | 3 | 4 | 4 | 0.003 | 0.008 | 0.006
2014/2015 | 15 | 0.16 | 0.47 | 0.30 | 0.11 | 1.72 | 0.56 | 1 | 10 | 4 | 0.001 | 0.007 | 0.004
Overall | 0.16 | 0.47 | 0.29 | 0.11 | 1.72 | 0.54 | 1 | 10 | 4 | 0.001 | 0.008 | 0.004

Table 13.25 Master Test Samples – Iron Cap (2010–2015)

Year/Test Program | Sample Quantity | Cu (%) | Au (g/t) | Ag (g/t) | Mo (%)
Min | Max | Avg | Min | Max | Avg | Min | Max | Avg | Min | Max | Avg
2010 | 3 | 0.14 | 0.36 | 0.25 | 0.32 | 1.06 | 0.71 | 5 | 6 | 6 | 0.002 | 0.003 | 0.003
2014/2015 | 4 | 0.16 | 0.34 | 0.27 | 0.35 | 0.63 | 0.49 | 2 | 5 | 4 | 0.002 | 0.004 | 0.003
Overall | 0.14 | 0.36 | 0.26 | 0.32 | 1.06 | 0.58 | 2 | 6 | 4 | 0.002 | 0.004 | 0.003

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Preliminary mineralogy work was conducted in 2010 and 2015. The results from the two programs indicated that chalcopyrite is the major copper-bearing mineral with a portion of chalcocite, covellite, and tennantite/tetrahedrite. Pyrite is the dominant sulphide mineral. Feldspars, micas, and quartz are the main gangue minerals.

Grindability

The 2010 grindability determination tests on the two Iron Cap composite samples showed that the mineralization is of moderate hardness to ball mill grinding. The Bond ball mill work indices of both the samples are 14.9 kWh/t.

The IC-2014-MC4 master composite tested by the 2015 test work shows a slightly higher Bond ball mill work index of 16.5 kWh/t. The Ai was measured to be 0.099 g.

The program also tested the grindability of the sample to SAG mill grinding using the SMC procedure. The results show that the SMC parameters are A = 68.7, b = 0.54, and A x b = 37.1.

Iron Cap Cu-Mo Bulk Concentrate Flotation Results

Using the same metallurgical flowsheets used for the other deposits, Iron Cap samples were tested with batch open circuit flotation test procedure, including cleaner flotation tests, and locked cycle flotation test procedure.

In the 2014 testing program, a total of 15 drill core interval composite samples (assay coarse rejects), including four composite samples constructed from the 15 assay reject samples, were used for variability tests. The target primary grind sizing was 80% passing 125 μm. The pH was adjusted to 10.5 in the rougher flotation and 11.5 in the cleaner flotation with lime.

As shown in Figure 13.6 and Figure 13.7, copper recoveries to the bulk concentrates (open circuit batch tests) for the 15 Iron Cap variability samples ranged from 74% to 85% and averaged approximately 81%. The bulk concentrate grades ranged from approximately 25% to 30% copper, averaging around 27% copper. Considerable variation in gold recovery into the flotation concentrate was observed. On average, approximately 57% of the gold was recovered into the final cleaner concentrates from open circuit batch flotation.

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Figure 13.6 Variability Test Results – Copper – Iron Cap, 2015 (ALS) – KM4672

| Source: | ALS Canada, Ltd., ALS Metallurgy Kamloops (2015). Preliminary Metallurgical Testing – Iron Cap – KSM Project – (KM 4672). August 04, 2015.

Figure 13.7 Variability Test Results – Gold – Iron Cap, 2015 (ALS) – KM4672

| Source: | ALS Canada, Ltd., ALS Metallurgy Kamloops (2015). Preliminary Metallurgical Testing – Iron Cap – KSM Project – (KM 4672). August 04, 2015.

The rougher flotation tests for Composite IC-2014-MC4 conducted in 2015 showed that copper and gold recoveries were improved, after the primary grind size had been reduced from 80% passing 171 µm to 80% passing 89 µm. The test results are depicted in Figure 13.8 and Figure 13.9.

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Figure 13.8 Copper Recovery vs. Rougher Mass Recovery and Grind Size – Iron Cap, 2015 (ALS) – KM4672

| Source: | ALS Canada, Ltd., ALS Metallurgy Kamloops (2015). Preliminary Metallurgical Testing – Iron Cap – KSM Project – (KM 4672). August 04, 2015.

Figure 13.9 Gold Recovery vs. Rougher Mass Recovery and Grind Size – Iron Cap, 2015 (ALS) – KM4672

| Source: | ALS Canada, Ltd., ALS Metallurgy Kamloops (2015). Preliminary Metallurgical Testing – Iron Cap – KSM Project – (KM 4672). August 04, 2015.

The flotation LCT results are presented in Table 13.26. The results indicate that the copper recoveries from the Iron Cap samples were comparable to the Mitchell mineralization. The gold recoveries to the concentrates from the 2012 samples were lower than those achieved with the Mitchell mineralization; however, the 2014/2015 test work produced better gold recoveries to the flotation concentrates and the results are in line with the Mitchell mineralization. The averaged silver recovery to the flotation concentrate was slightly higher than the recovery achieved from the Mitchell mineralization. On average, approximately 61% of the molybdenum from the Iron Cap samples reported to the final bulk concentrate.

As shown in Table 13.8, the Mitchell and Iron Cap blended samples did not show detrimental effects of the blending on the metallurgical responses.

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Table 13.26 Locked Cycle Test Results – Iron Cap

Test Program* | Composite | Grind Size (P80 µm**) | Head Grade | Bulk Conc. Grade | Flotation Recovery
Cu (%) | Au (g/t) | Ag (g/t) | Mo (%) | Cu (%) | Au (g/t) | Ag (g/t) | Mo (%) | Cu (%) | Au (%) | Ag (%) | Mo (%)
G&T 2748/11 | Iron Cap 2010 Composite1 | 150/15 | 0.14 | 1.28 | 6 | 0.002 | 25.4 | 147 | 774 | 0.180 | 81.6 | 55.2 | 61.0 | 37.9
G&T 2748/12 | Iron Cap 2010 Composite2 | 147/22 | 0.38 | 0.31 | 5 | 0.003 | 24.9 | 10 | 255 | 0.115 | 88.1 | 45.0 | 62.0 | 55.2
G&T 2748/17 | 50%Comp 1: 50%Comp 2 | 108/19 | 0.26 | 0.82 | - | 0.003 | 26.7 | 51.9 | - | 0.144 | 85.2 | 53.3 | - | 41.5
ALS 4029/19 | IC-2013-01 | 117/16/7 | 0.25 | 0.56 | 3 | 0.006 | 26.7 | 50.4 | 179 | 0.481 | 83.4 | 71.4 | 48.4 | 69.0
ALS 4029/20 | IC-2013-02 | 119/17/11 | 0.26 | 0.51 | 4 | 0.004 | 22.9 | 36.4 | 273 | 0.238 | 85.4 | 68.4 | 62.8 | 64.4
ALS 4029/21 | IC-2013-03 | 130/15/11 | 0.23 | 0.24 | 4 | 0.003 | 23.4 | 14.8 | 258 | 0.205 | 87.5 | 53.4 | 56.3 | 56.4
ALS 4514/30*** | IC-2014-MC1 | 125/14/14 | 0.33 | 0.69 | 4 | 0.004 | 25.4 | 41.5 | 230 | 0.238 | 87.1 | 67.8 | 60.1 | 70.3
ALS 4514/29*** | IC-2014-MC2 | 127/14/16 | 0.25 | 0.45 | 4 | 0.004 | 23.4 | 29.3 | 275 | 0.311 | 85.7 | 59.1 | 56.0 | 76.2
ALS 4514/31*** | IC-2014-MC3 | 124/16/18 | 0.16 | 0.32 | 2 | 0.002 | 22.6 | 37.6 | 139 | 0.187 | 81.4 | 67.9 | 47.4 | 62.9
ALS 4514/25*** | IC-2014-MC4 | 124/14/15 | 0.28 | 0.56 | 4 | 0.003 | 24.9 | 36.2 | 250 | 0.257 | 85.7 | 62.9 | 65.1 | 73.6

| Notes: | *Au grade in the Bulk Cleaner Tailings ranged from 0.61 to 2.17 g/t Au, averaging at 1.27 g/t Au; Au grade in the pyrite concentrate ranged from 0.08 to 1.88 g/t Au, averaging at 0.52 g/t Au.

**Primary grind size/regrind size

***Including a copper flotation on the pyrite flotation concentrate

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Cyanidation Leach Results

G&T conducted cyanidation tests on the first cleaner tailing and the gold-bearing pyrite concentrate produced from the flotation LCTs. Test results are provided in Table 13.27.

Table 13.27 Cyanidation Test Results on LCT Products – Iron Cap

Testing Program | Sample | Regrind Size (P80 µm) | Feed (Au g/t) | Extraction (Au %) | Feed (Ag g/t) | Extraction (Ag %)
G&T-2748 | Iron Cap Comp1 | 14 | 1.9 | 49.7 | 9.4 | 62.8
G&T-2748 | Iron Cap Comp2 | 15 | 1.1 | 40.4 | 6.9 | 56.8
G&T-2748 | 50% Comp1/50% Comp2 | 16 | 1.5 | 48.6 | - | -
ALS-4029 | IC-2013-01/02/03 Cl.Sc.Tls | 16 | 0.8 | 45.8 | 4.4 | 70.7
ALS-4029 | IC-2013-01/02/03 Py Conc | 10 | 0.2 | 54.7 | 1.6 | 87.4
ALS-4672 | IC-2014-MC1 Cl.Sc.Tls | 14 | 1.1 | 46.7 | 6 | 68.0
ALS-4672 | IC-2014-MC2 Cl.Sc.Tls | 14 | 1.2 | 29.2 | 7 | 69.7
ALS-4672 | IC-2014-MC4 Cl.Sc.Tls | 14 | 1.1 | 40.1 | 6 | 60.5
ALS-4672 | IC-2014-MC1 Py Conc | 14 | 0.4 | 50.6 | 3 | 57.2
ALS-4672 | IC-2014-MC2 Py Conc | 16 | 0.3 | 36.3 | 2 | 72.0
ALS-4672 | IC-2014-MC4 Py Conc | 15 | 0.1 | 74.9 | 5 | 73.2
Average – Iron Cap | 15 | 1.0 | 46.8 | 5.6 | 66.5

On average, the gold extraction from both the gold-bearing products was approximately 47%. As compared with the Mitchell samples, Iron Cap samples produced lower gold recoveries, especially for leaching of the first cleaner tailing. The average silver extraction of 67% is higher than the average extraction of 56% obtained the Mitchell samples.

| 13.5 | Recent Test Work 2017–2020

Since 2017, further testing programs have been carried out on the samples from the Deep Kerr deposit and the Iron Cap deposit to investigate their metallurgical response to the previously developed process flowsheet. The respective mineralogical characteristics and ore grindability have also been determined.

| 13.5.1 | Deep Kerr Zone Metallurgical Test Work (2016/2017)

In 2017, two additional metallurgical testing programs, KM5063 and KM5266, have been completed by ALS Metallurgy with the Deep Kerr deposit samples.

Deep Kerr Samples

Variability and master samples were prepared for both the test programs. The samples used in test program KM5036 included 1) the coarse assay reject materials at a size minus 6 mesh (3.4 mm) for approximately 783 kg, and 2) the fresh half core samples of approximately 262 kg. The samples used in test program KM5266 were assay reject materials less than 10 mesh (2 mm), with a total weight of approximately 662 kg.

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Table 13.28 displays the major chemical composition of the head samples in the test program KM5063. All the individual samples and master samples including Master Composites 215E, 215W, 280, 775, and Overall were prepared with coarse reject materials; the remaining master samples were composited using half drill cores. In addition, composite samples, DK-2014-MC3 and Mitchell M1 from other programs (KM4514 and KM3080), were also tested in KM5063.

Table 13.28 Metal Contents of Composites – Deep Kerr, 2016/2017 (ALS KM5063)

Master Samples* | Metal Content | Individual Samples | Metal Content
Cu (%) | Au (g/t) | Mo (%) | Ag (g/t) |  | Cu (%) | Au (g/t) | Mo (%) | Ag (g/t)
DK-2015-Master-215E | 0.46 | 0.34 | 0.005 | 1.1 | DK-2015-215E-01 | 0.34 | 0.25 | 0.005 | 0.3
DK-2015-Master-215W | 0.51 | 0.43 | 0.002 | 2.6 | DK-2015-215E-02 | 0.52 | 0.37 | 0.005 | 0.7
DK-2015-Master-280 | 0.48 | 0.36 | 0.003 | 2.1 | DK-2015-215W-01 | 0.51 | 0.34 | 0.002 | 2.6
DK-2015-Master-775 | 0.53 | 0.31 | 0.003 | 1.8 | DK-2015-215W-02 | 0.49 | 0.49 | 0.002 | 3.2
DK-2015-Overall Master | 0.50 | 0.36 | 0.003 | 1.9 | DK-2015-215W-03 | 0.51 | 0.50 | 0.002 | 2.4
DK-2015-215E Comp 2** | 0.48 | 0.31 | 0.003 | 0.8 | DK-2015-215W-04 | 0.49 | 0.45 | 0.002 | 2.4
DK-2015-215W Comp 2** | 0.30 | 0.39 | 0.002 | 1.5 | DK-2015-215W-05 | 0.50 | 0.58 | 0.002 | 2.8
DK-2015-280 Comp 2** | 0.42 | 0.37 | 0.002 | 1.4 | DK-2015-215W-06 | 0.44 | 0.32 | 0.001 | 1.8
DK-2015-775 Comp 2** | 0.49 | 0.30 | 0.004 | 1.7 | DK-2015-215W-07 | 0.54 | 0.38 | 0.005 | 2.7
DK-2015 Comp P2A** | 0.59 | 0.39 | 0.002 | 1.0 | DK-2015-215W-08 | 0.50 | 0.52 | 0.004 | 1.8
DK-2015 Comp P2B** | 0.67 | 0.36 | 0.006 | 1.1 | DK-2015-215W-09 | 0.86 | 0.51 | 0.004 | 3.4
DK-2015 Comp SED’s WR** | 0.25 | 0.84 | 0.003 | 1.9 | DK-2015-215W-10 | 0.86 | 0.48 | 0.002 | 4.1
|  |  |  |  | DK-2015-215W-11 | 0.18 | 0.24 | 0.001 | 1.3
|  |  |  |  | DK-2015-280-1 | 0.42 | 0.36 | 0.003 | 2.2
|  |  |  |  | DK-2015-280-2 | 0.50 | 0.43 | 0.002 | 1.5
|  |  |  |  | DK-2015-775-01 | 0.52 | 0.27 | 0.003 | 0.8
|  |  |  |  | DK-2015-775-02 | 0.49 | 0.38 | 0.003 | 2.1

| Notes | *Master composite samples were prepared to represent different proposed “block caving” zones 215 W, 215E, 775, and 280

**Constructed from half drill core samples

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Table 13.29 shows the major chemical compositions of the head samples used in the KM5266 test program involving six individual samples and two master samples.

Table 13.29 Metal Contents of Composites – Deep Kerr, 2017 (ALS KM5266)

Master Samples | Metal Content | Individual Samples | Metal Content
Cu (%) | Au (g/t) | Mo (%) | Ag (g/t) | Cu (%) | Au (g/t) | Mo (%) | Ag (g/t)
Mineral Intrusive Blend | 0.51 | 0.35 | 0.002 | 3 | DK-2016-01 | 0.22 | 0.43 | 0.002 | 3
Wall Rock Sedimentary | 0.25 | 0.43 | 0.003 | 4 | DK-2016-02 | 0.49 | 0.39 | 0.002 | 2
|  |  |  |  | DK-2016-03 | 0.67 | 0.50 | 0.001 | 2
|  |  |  |  | DK-2016-04 | 0.27 | 0.44 | 0.002 | 2
|  |  |  |  | DK-2016-05 | 0.50 | 0.25 | 0.003 | 2
|  |  |  |  | DK-2016-06 | 0.26 | 0.50 | 0.003 | 4

Grindability

A Bond ball mill work index testing has been conducted on the Deep Kerr master samples at a closing screen size of 106 µm. The measured BWi index is between 13.2 kWh/t and 15.0 kWh/t, which indicates the tested samples were moderately soft to average hardness. Table 13.30 displays the results.

Table 13.30 Bond Ball Mill Work Index Test Results – Deep Kerr, 2017 (ALS)

Samples | BWi (kWh/t)
KM5063 |
DK-2015 Master 215E | 13.2
DK-2015 Master 215W | 15.0
DK-2015 Master 280 | 14.6
DK-2015 Master 775 | 14.6
Average | 14.4

Mineralogy

The mineralogical composition of the Deep Kerr head samples has been determined in both the test programs and is summarized in Table 13.31. As with Mitchell Zone samples, the primary copper sulphide mineral has been identified as chalcopyrite. Pyrite content is high, between 5.1% and 10.9%. The resulting pyrite to chalcopyrite ratio is between 2.8:1 and 8.5:1.

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Table 13.31 Mineral Composition Data – Deep Kerr 2016/2017 (ALS)

Sample | Mineral Composition (%)
Chalcopyrite | Pyrite | Gangue*
KM5063 |  |  |
DK-2015 Master 215E | 1.7 | 8.9 | 89.4
DK-2015 Master 215W | 1.4 | 5.5 | 93.1
DK-2015 Master 280 | 1.4 | 10.9 | 87.7
DK-2015 Master 775 | 1.7 | 10.0 | 88.3
KM5266 |  |  |
Mineral Intrusive Blend | 1.8 | 5.1 | 93.1
Wall Rock Sedimentary | 0.8 | 6.8 | 92.4

| Note: | *Gangue minerals include iron oxides, quartz, feldspars, muscovite, chlorite, carbonates, biotite, apatite, calcium sulphate, titanium minerals, kaolinite, etc.

As presented in Figure 13.10 and Figure 13.11, liberation rates of the copper sulphide minerals has also been estimated with a range from 56% to 66% (KM5063) and 53% to 61% (KM5266), at a primary grind size of 80% passing approximately 130 µm. Most unliberated copper sulphide particles were associated in binary form with non-sulphide gangue minerals.

Figure 13.10 KM5063 Copper Sulphides Liberation – Mitchell and Deep Kerr 2017 (ALS)

Source: ALS Metallurgy (2017). Preliminary Metallurgical Testing – Deep Kerr Zones – KSM Project (KM 5063). January 18, 2017.

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Figure 13.11 KM5266 Copper Sulphides Liberation – Deep Kerr 2017 (ALS)

Source: ALS Metallurgy (2017), Preliminary Metallurgical Testing – Deep Kerr Zones – KSM Project (KM 5266), April 13, 2017.

Metallurgical and Processing Tests

The recent metallurgical testing on the Deep Kerr samples has been conducted to establish the metallurgical performance of the samples using the previous metallurgical flowsheet. The test flowsheet included conventional flotation to produce a copper-gold-molybdenum bulk concentrate, with cyanide leach testing to further recover gold and silver from the cleaner flotation tailing and the gold-bearing pyrite concentrate. A copper scavenging flotation has been added to float copper minerals from the gold-bearing pyrite concentrate prior to cyanidation.

Variability Flotation Tests

Batch open circuit cleaner flotation testing has been conducted on variability samples in both the KM5063 and KM5266 test programs. The metallurgical flowsheet is shown in Figure 13.12. In an effort to improve copper and gold recoveries, an optional flowsheet with two stages of regrinding on the bulk copper and gold rougher concentrate was tested in KM5063. The testing produced a low grade bulk copper and gold concentrate and the test results from this flowsheet are not included in the following discussions.

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Figure 13.12 KM5063 and KM5266 Open Circuit Flotation Flowsheet – Deep Kerr 2016/2017 (ALS)

Source: ALS Metallurgy (2017), Preliminary Metallurgical Testing – Deep Kerr Zones – KSM Project (KM 5063) January 18, 2017 and ALS Metallurgy (2017), Preliminary Metallurgical Testing – Deep Kerr Zones – KSM Project (KM 5266) April 13, 2017.

The following observations have been made from both cleaner open cycle test programs:

| ● | Cleaner flotation tests in KM5063 – The coarse reject samples produced a relatively low copper recovery to the bulk concentrate, which ranged from 61% to 80%, at copper grades from 21% to 30% Cu. The test results were improved when using fresh drill core samples that produced a higher copper recovery from 81% to 90%, at copper grades from 18% to 31% Cu. The average gold recovery was similar, which was about 51% at a grade of 14 g/t Au for the coarse reject samples, and 48% and 17 g/t Au for the fresh core samples.

| ● | Cleaner flotation tests in KM5266 – The fine reject samples produced copper recoveries ranged from 72% to 86% at grades from 18% to 28% Cu. The two master samples produced similar copper grades of the bulk concentrate between 25% and 26% Cu; the copper recovery was 81% for the Mineral Intrusive Blend composite master sample and 77% for the Wall Rock Sedimentary composite master sample. A similar gold recovery for the two master samples occurred between 50% and 51%; however, a higher gold grade of 29 g/t Au was produced from the Wall Rock Sedimentary sample, while the other was graded at 12 g/t Au.

Locked Cycle Flotation Tests

The locked cycle flotation tests in the testing program KM5063 and KM5266 have been completed using the flowsheet from the previous tests (Figure 13.13) on the master samples.

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Figure 13.13 KM5063 and KM5266 LCT Flotation Flowsheet – Deep Kerr 2016/2017 (ALS)

Source: ALS Metallurgy (2017), Preliminary Metallurgical Testing – Deep Kerr Zones – KSM Project (KM 5063) January 18, 2017 and ALS Metallurgy (2017), Preliminary Metallurgical Testing – Deep Kerr Zones – KSM Project (KM 5266) April 13, 2017.

The test results and the grind size information are presented in Table 13.32. The following major observations were made from these two sets of LCTs:

| ● | Locked cycle flotation tests in KM5063 – The coarse reject samples responded well to the applied flowsheet. Copper recovery to the bulk concentrates ranged from 81% to 88% and gold recovery varied from 55% to 77%. The bulk concentrate grades were 22% to 26% Cu and 11 g/t to 16 g/t Au. However, molybdenum grades in the bulk concentrates were low and between 0.09% and 0.24% Mo.

As shown Table 13.32, the LCTs on the fresh drill core samples produced higher copper recoveries to the bulk concentrate ranged from 89% to 93%, grading at 25% and 27% Cu. Gold recoveries to the bulk concentrates were between 51% and 60%, with grades from 7.7 g/t to 42 g/t Au. This suggests that the assay reject samples might have undergone some degree surface oxidation.

| ● | Locked cycle flotation tests in KM5266 – For the Mineral Intrusive Blend master sample, approximately 90% copper was recovered to the bulk concentrate graded at 25% Cu; gold recovery was 61% at a grade of 12 g/t Au. For the Wall Rock Sedimentary master sample, copper recovery was about 85% at a grade of 21% Cu; the gold recovery was 62% graded at 28 g/t Au. Similarly, molybdenum grades of the bulk concentrate were low and between 0.04% and 0.11% Mo.

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Table 13.32 Flotation Locked Cycle Test Results – Deep Kerr 2016/2017

Test* | Sample | Grind Size (P80 µm**) | Feed Grade | Bulk Conc. Grade | Recovery
Cu (%) | Au (g/t) | Ag (g/t) | Mo (%) | Cu (%) | Au (g/t) | Ag (g/t) | Mo (%) | Cu (%) | Au (% | Ag (%) | Mo (%)
KM5063-10 | DK2015-Overall Master | 121/15/14 | 0.50 | 0.36 | 3 | 0.003 | 25.7 | 12.8 | 81 | 0.15 | 84.1 | 57.4 | 49.1 | 70.9
KM5063-11 | DK2015-Master 215E | 115/16/15 | 0.47 | 0.33 | 2 | 0.005 | 23.2 | 12.7 | 51 | 0.24 | 87.1 | 67.6 | 51.2 | 81.4
KM5063-12 | DK2015-Master 215W | 120/14/14 | 0.51 | 0.45 | 3 | 0.003 | 25.8 | 15.6 | 103 | 0.091 | 87.5 | 59.6 | 53.3 | 61.3
KM5063-13 | DK2015-Master 280 | 134/16/15 | 0.48 | 0.38 | 2 | 0.003 | 21.6 | 11.3 | 63 | 0.093 | 84.0 | 56.6 | 53.0 | 53.6
KM5063-14 | DK2015-Master 775 | 133/16/15 | 0.54 | 0.34 | 2 | 0.003 | 24.5 | 10.6 | 68 | 0.13 | 81.5 | 56.7 | 52.6 | 71.3
KM5063-64 | DK2015-Fresh 775 | 125/17/17 | 0.51 | 0.26 | 2 | 0.004 | 25.1 | 7.7 | 56 | 0.11 | 91.5 | 55.6 | 52.6 | 48.6
KM5063-65 | DK2015-Fresh 280 | 130/16/19 | 0.43 | 0.34 | 2 | 0.002 | 26.3 | 12.2 | 79 | 0.068 | 90.4 | 53.1 | 51.5 | 44.1
KM5063-66 | DK2015-Fresh 215W | 122/14/17 | 0.33 | 0.38 | 3 | 0.002 | 27.4 | 18.1 | 166 | 0.043 | 90.1 | 50.5 | 56.8 | 23.0
KM5063-67 | DK2015-Fresh 215E | 120/18/19 | 0.49 | 0.28 | 1 | 0.003 | 25.9 | 9.4 | 38 | 0.078 | 93.1 | 60.3 | 47.0 | 42.6
KM5063-78 | DK2015-Fresh SEDS WR | 128/15/20 | 0.26 | 0.74 | 3 | 0.003 | 24.7 | 42.0 | 144 | 0.14 | 88.8 | 53.4 | 54.1 | 39.5
KM5266-09 | Mineral Intrusive Blend Master | 127/12/14 | 0.56 | 0.39 | 3 | 0.002 | 24.8 | 11.7 | 73 | 0.038 | 89.8 | 60.9 | 54.8 | 38.6
KM5266-10 | Wall Rock Sedimentary Master | 134/14/15 | 0.27 | 0.48 | 3 | 0.002 | 21.0 | 27.7 | 192 | 0.105 | 85.0 | 61.6 | 61.9 | 46.0

| Notes: | *Au grade in the Bulk Cleaner Tailings ranged from 0.47 to 1.50 g/t Au, averaging at 0.76 g/t Au; Au grade in the pyrite concentrate ranged from 0.07 to 0.62 g/t Au, averaging at 0.24 g/t Au.

**Grind Size includes primary grind size/bulk concentrate regrind size/pyrite concentrate regrind size.

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Leach Tests

DCN leach tests have been conducted by ALS Metallurgy in 2016/2017 on the bulk cleaner scavenger tailing and copper first cleaner tailing (pyrite concentrate) produced from the locked cycle testing programs KM5063 and KM5266. The direct leaching procedure for both programs are identical, including a 16-hour pre-aeration and a 24-hour cyanide leaching. Additional CIL tests were conducted in the test program KM5063.

All the leach test results are presented in Table 13.33. The average gold extraction rates from both the tailing samples are higher than previous results.

Table 13.33 Preliminary Cyanidation Test Results – Deep Kerr

Test | Head Samples | Test Procedure* | Extraction (%)
Au | Ag
Bulk Cleaner Scavenger Tailing Cyanidation
KM5063-15 | DK-2015-Overall Master | CIL | 62.2 | 83.6
KM5063-16 | DK-2015-Master 215E | CIL | 65.7 | 71.7
KM5063-17 | DK-2015-Master 215W | CIL | 61.0 | 55.5
KM5063-18 | DK-2015-Master 280 | CIL | 52.9 | 77.5
KM5063-19 | DK-2015-Master 775 | CIL | 54.6 | 82.2
KM5063-69 | DK-2015-Seds-WR | CIL | 58.1 | 74.8
KM5063-70 | DK-2015-775 Comp 2 | CIL | 61.5 | 69.6
KM5063-71 | DK-2015-280 Comp 2 | CIL | 48.2 | 72.7
KM5063-72 | DK-2015-215W Comp 2 | CIL | 57.1 | 79.8
KM5063-73 | DK-2015-215E Comp 2 | CIL | 60.8 | 73.5
KM5063-25 | DK-2015-Overall Master | DCN | 70.7 | 55.5
KM5063-40** | DK-2015-Overall Master | DCN | 64.2 | 42.6
KM5063-41** | DK-2015-Overall Master | DCN | 65.8 | 43.3
KM5063-74 | DK-2015-775 Comp 2 | DCN | 68.1 | 52.8
KM5063-75 | DK-2015-280 Comp 2 | DCN | 52.3 | 67.8
KM5063-76 | DK-2015-215W Comp 2 | DCN | 53.2 | 68.4
KM5063-77 | DK-2015-215E Comp 2 | DCN | 63.9 | 46.5
KM5063-79 | DK-2015-Seds-WR | DCN | 67.4 | 82.7
KM5266-11 | Mineral Intrusive Blend Master | DCN | 70.3 | 93.4
KM5266-12 | Wall Rock Sedimentary Master | DCN | 72.9 | 78.9
Copper First Cleaner Tailing Cyanidation (Au-Pyrite Concentrate)
KM5063-20 | DK-2015-Overall Master | CIL | 83.5 | 71.8
KM5063-21 | DK-2015-Master 215E | CIL | 91.5 | 58.3
KM5063-22 | DK-2015-Master 215W | CIL | 70.2 | 73.1
KM5063-23 | DK-2015-Master 280 | CIL | 71.9 | 58.2
KM5063-24 | DK-2015-Master 775 | CIL | 82.3 | 66.6
KM5266-13 | Mineral Intrusive Blend Master | DCN | 87.9 | 94.3
KM5266-14 | Wall Rock Sedimentary Master | DCN | 71.2 | 96.5

| Notes: | *Cyanide concentration: 1,000 ppm; pH: 11; carbon addition: 28 g/L

**Cyanidation was performed on a finer feed particle size

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For the bulk cleaner scavenger tailing containing approximately 0.6 g/t to 1.6 g/t gold and 1.8 g/t to 6.3 g/t silver, the extractions fluctuated widely from 48% to 73% for gold, averaging 61.5%, and from 42% to 93% for silver, averaging 68.6%.

For the gold-bearing pyrite concentrates grading at 0.17 g/t to 0.49 g/t gold and 0.9 g/t to 2.0 g/t silver, the extractions also varied significantly, ranging from 70% to 92% for gold, averaging 79.8%, and from 58% to 97% for silver, averaging 74.1%. The gold and silver extractions varies substantially and did not exhibit correlations with the leach feed grades including gold and copper, as well as the regrind size which was maintained at 80% passing 14-15 µm.

The CIL test results also show that a significant amount of copper was leached out together with the precious metals. It appears that the copper dissolution had a significant impact on the gold adsorption onto the activated carbon but no obvious effect on the silver adsorption. On average, only approximately 66% of the extracted gold was captured onto the carbon.

In 2016, ALS Metallurgy also completed a metallurgical test program (KM5087) to generate materials for gold and silver extraction testing. The samples generated were the first cleaner scavenger tailings and pyrite concentrates from a blended master composite made up of both the Mitchell and Deep Kerr zone samples. ALS Metallurgy conducted preliminary cyanide leach tests after the samples had been treated by various pre-aeration methods. On average, approximately 64% of the gold was extracted. The average silver extraction was approximately 90%, while the copper extraction averaged at approximately 29%.

| 13.5.2 | Iron Cap Zone Metallurgical Test Work (2017–2020)

Four additional metallurgical test programs, KM5248, KM5501, KM5806, and KM6004, have been performed by ALS Metallurgy on samples from Iron Cap zone through 2017 to 2020.

Iron Cap Samples

Coarse assay rejects and drill core samples have been used to construct the metallurgical composite samples in the four testing programs. Table 13.34 shows the major chemical analysis results of these samples. The copper content is between 0.01% and 0.86% Cu, while gold and silver grades are 0.17 g/t to 2.91 g/t Au and 1 g/t to 6 g/t Ag. The IC-2017-08 composite sample is from the area close to the surface (out of the potential upper panel caves), contains high gold and low copper (approximately 0.01% Cu), and has significantly different mineralogical characteristics. The sample contained elevated arsenic.

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Table 13.34 Metal Contents of Composites – Iron Cap, 2017/2018 (ALS KM5248/5501)

Test Program | Head Samples | Cu (%) | Au (g/t) | Ag (g/t) | Mo (%) | Fe (%) | S (%) | Hg (g/t) | As (%)
KM5248 | IC-2016-01 | 0.86 | 0.97 | 5 | 0.003 | 3.8 | 3.26 | 0.60 | 0.01
IC-2016-02 | 0.37 | 0.42 | 6 | 0.007 | 3.1 | 3.18 | 0.28 | <0.01
IC-2016-03 | 0.24 | 1.30 | 5 | 0.002 | 2.7 | 1.92 | 0.09 | <0.01
IC-2016-04 | 0.18 | 0.78 | 4 | 0.002 | 3.6 | 1.56 | 0.05 | <0.01
IC-2016-01/02 | 0.60 | 0.76 | 4 | 0.004 | 3.3 | 3.24 | n/a | n/a
IC-2016-003/04 | 0.21 | 1.00 | 3 | 0.001 | 3.1 | 1.66 | n/a | n/a
KM5501 | IC-2017-01 | 0.40 | 0.47 | 2 | 0.003 | 2.66 | 1.93 | 0.43 | <0.01
IC-2017-02 | 0.72 | 0.98 | 4 | 0.005 | 3.48 | 2.97 | 0.51 | <0.01
IC-2017-03 | 0.25 | 0.50 | 3 | 0.003 | 0.24 | 2.47 | 0.21 | <0.01
IC-2017-04 | 0.36 | 0.62 | 2 | 0.003 | 2.82 | 1.50 | 0.10 | <0.01
IC-2017-05 | 0.39 | 0.63 | 4 | 0.004 | 3.17 | 3.27 | 0.14 | <0.01
IC-2017-06 | 0.55 | 0.17 | 5 | 0.016 | 3.40 | 4.20 | 1.61 | 0.03
IC-2017-07 | 0.25 | 0.69 | 1 | 0.002 | 3.43 | 1.80 | 0.07 | <0.01
IC-2017-08 | 0.01 | 2.91 | <1 | <0.001 | 5.40 | 2.65 | 0.08 | 0.31
IC-MC-A | 0.31 | 0.62 | 2 | 0.003 | 3.29 | 2.36 | 0.10 | <0.01
IC-MC-B | 0.58 | 0.53 | 1 | 0.008 | 3.27 | 3.21 | 0.88 | <0.01
IC-MC-C | 0.43 | 0.57 | 1 | 0.005 | 3.24 | 2.85 | 0.45 | <0.01
KM5806 | IC-2018-01 | 0.18 | 0.19 | 2 | 0.006 | 3.3 | 2.08 | 0.27 | <0.01
IC-2018-02 | 0.36 | 0.37 | 2 | 0.003 | 3.9 | 1.99 | 0.42 | <0.01
IC-2018-03 | 0.42 | 0.74 | 2 | 0.002 | 2.5 | 1.98 | 0.13 | <0.01
IC-2018-04 | 0.28 | 0.35 | 2 | 0.005 | 3.6 | 1.44 | 0.10 | <0.01
IC-2018-05 | 0.49 | 0.74 | 3 | 0.002 | 3.0 | 1.91 | 0.22 | <0.01
IC-2018-06 | 0.34 | 0.57 | 4 | 0.007 | 5.6 | 6.40 | 3.04 | 0.11
KM6004 | IC-2018-07 | 0.32 | 0.64 | 4 | 0.004 | 3.8 | 3.54 | 0.50 | 0.01
IC-2018-08 | 0.33 | 0.46 | 1 | 0.002 | 3.9 | 1.82 | 0.32 | <0.01
IC-2018-09 | 0.38 | 0.36 | 2 | 0.001 | 2.4 | 1.72 | 0.20 | <0.01
IC-2018-10 | 0.55 | 1.03 | 4 | 0.004 | 3.9 | 2.95 | 0.28 | <0.01
IC-2018-11 | 0.38 | 0.53 | 2 | 0.002 | 3.1 | 2.46 | 0.51 | <0.01
IC-2018-12 | 0.20 | 0.28 | 1 | 0.002 | 2.9 | 1.75 | 0.12 | <0.01
IC-2018-13 | 0.47 | 0.39 | 4 | 0.005 | 3.3 | 3.81 | 0.85 | 0.02
IC-2018-14 | 0.41 | 0.61 | 2 | 0.003 | 3.1 | 2.34 | 0.14 | <0.01
IC-2018-15 | 0.32 | 0.51 | 2 | 0.003 | 3.3 | 1.91 | 0.14 | <0.01
IC-2018-16 | 0.28 | 0.88 | 2 | 0.002 | 2.5 | 1.40 | 0.18 | <0.01
IC-2018-17 | 0.37 | 0.47 | 2 | 0.002 | 2.6 | 2.39 | 0.20 | <0.01
IC-2018-18 | 0.36 | 0.48 | 2 | 0.003 | 3.1 | 2.42 | 0.34 | <0.01

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Grindability

Bond ball mill work index tests were conducted as part of metallurgical programs KM5501, KM5806, and KM6004. The overall results are presented in Table 13.35, which shows that the work index ranges from 13.4 kWh/t to 17.4 kWh/t. The measured index indicates a moderate hardness to ball mill grinding for the tested Iron Cap samples. The measured abrasion index fluctuates between 0.094 g and 0.200 g, with an average level of 0.155 g.

In addition, SMC testing on the composite sample in the program KM5806 produced an A x b value of approximately 41 with a SAG Circuit Specific Energy value of 9.8 kWh/t.

Table 13.35 Bond Ball Mill Work Index Test Results – Deep Kerr, 2018 (ALS)

Samples | Wi (kWh/t) | Samples | Wi (kWh/t) | Samples | Wi (kWh/t) | Abrasion Index (g)
KM5501 | KM5806 | KM6004
IC-2017-01 | 15.9 | IC-2018-01 | 17.4 | IC-2018-13 | 15.2 | 0.094
IC-2017-02 | 16.1 | IC-2018-02 | 15.8 | IC-2018-14 | 15.8 | 0.130
IC-2017-05 | 15.6 | IC-2018-03 | 15.6 | IC-2018-15 | 15.9 | 0.143
IC-2017-07 | 16.6 | IC-2018-04 | 15.9 | IC-2018-16 | 15.6 | 0.186
IC-2014-MC1 | 16.3 | IC-2018-05 | 15.6 | IC-2018-17 | 15.1 | 0.200
IC-2014-MC2 | 15.9 | IC-2018-06 | 13.4 | IC-2018-18 | 15.7 | 0.175
IC-2014-MC3 | 16.7 |  |  |  |  |
IC-2016-01/02 | 17.4 |  |  |  |  |
IC-2016-03/04 | 16.0 |  |  |  |  |
IC-MC-A | 16.0 |  |  |  |  |
IC-MC-B | 15.9 |  |  |  |  |
KM5501 Average | 16.2 | KM5806 Average | 15.6 | KM6604 Average | 15.6 | 0.155
Overall Wi Average, kWh/t | 15.9

Mineralogy

The mineralogical composition and liberations of the composite samples have been included in all the testing programs. Table 13.36 shows the major results. As with the Mitchell and Deep Kerr zone samples, the primary copper sulphide mineral identified is chalcopyrite. However, tennantite and enargite have been identified as secondary dominate copper minerals in the sample IC-2018-06.

The pyrite content is between 2.1 and 9.2. The pyrite-to-chalcopyrite ratio fluctuates between 1.8:1 and 23:1.

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Table 13.36 Mineral Composition Data – Iron Cap 2017-2020 (ALS)

Sample | Mineral Composition (%) | Sizing (µm) | Liberation (%)
Chalcopyrite | Tennantite/ Enargite | Other Cu Minerals* | Pyrite | Gangue | P80 | Copper Sulphide
KM5248 |  |  |  |  |  |  |
IC-2016-01 | 2.4 | 0.1 | <.01 | 4.4 | 93.1 | 116 | 65
IC-2016-02 | 1.0 | <0.1 | <.01 | 5.5 | 93.5 | 127 | 53
IC-2016-03 | 0.6 | <0.1 | <0.1 | 2.6 | 96.8 | 121 | 54
KM5501 |  |  |  |  |  |  |
IC-2017-08 | <0.1 | <0.1 | <0.1 | 4.6 | 95.4 | 80 | 56
IC-MC-A | 0.7 | <0.1 | <0.1 | 3.8 | 95.5 | 123 | 46
IC-MC-B | 1.5 | 0.1 | <0.1 | 4.6 | 93.8 | 111 | 52
IC-2014-MC1 | 0.9 | <0.1 | <0.1 | 4.6 | 94.5 | n/a | n/a
IC-2014-MC2 | 0.6 | <0.1 | 0.0 | 4.0 | 95.4 | n/a | n/a
IC-2014-MC3 | 0.4 | - | <0.1 | 3.0 | 96.6 | n/a | n/a
IC-2016-04 | 0.5 | - | <0.1 | 3.0 | 96.5 | n/a | n/a
KM5806 |  |  |  |  |  |  |
IC-2018-01 | 0.4 | 0.0 | 0.0 | 3.8 | 95.8 | 82 | 41
IC-2018-02 | 0.8 | <0.1 | <0.1 | 3.3 | 95.9 | 95 | 40
IC-2018-03 | 1.2 | <0.1 | 0.0 | 3.6 | 95.2 | 86 | 46
IC-2018-04 | 0.8 | - | 0.0 | 2.2 | 97.0 | 85 | 52
IC-2018-05 | 1.2 | <0.1 | 0.0 | 2.1 | 96.7 | 85 | 44
IC-2018-06 | 0.4 | 0.5 | <0.1 | 9.2 | 89.9 | 80 | 51
KM6604 |  |  |  |  |  |  |
IC-2018-13 | 1.3 | 0.2 | <0.1 | 6.2 | 92.3 | 114 | 51
IC-2018-14 | 1.3 | <0.1 | <0.1 | 3.9 | 94.8 | 128 | 47
IC-2018-15 | 1.1 | <0.1 | <0.1 | 3.2 | 95.7 | 124 | 50
IC-2018-16 | 0.9 | <0.1 | <0.1 | 2.1 | 97.0 | 124 | 50
IC-2018-17 | 1.0 | <0.1 | <0.1 | 4.1 | 94.9 | 128 | 45
IC-2018-18 | 1.0 | <0.1 | <0.1 | 4.6 | 94.4 | 122 | 45

| Note: | *Other copper minerals refer to bornite and chalcocite/covellite.

Liberation of the copper sulphide minerals at the specified grind size were reported in a range of 53% to 65% (KM5248), 46% to 56% (KM5501), 40% to 52% (KM5806), and 45% to 51% (KM6004). Liberation of less than 50%, based on ALS Metallurgy, is not sufficient to efficiently recover copper minerals at the rougher flotation stage.

Metallurgical and Processing Tests

The Iron Cap deposit is porphyry copper mineralization associated with significant gold and silver contents, excluding a gold-only mineralization (Composite IC-2017-08) which was tested in the KM5501 test program. The metallurgical and processing test work are summarized in the following sections.

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Flotation Tests on Copper Sulphide Mineralization Samples

The test flowsheet used for the recent metallurgical tests on the copper sulphide mineralization samples is similar to the previous metallurgical tests. The flotation test work focused on the effect of the primary grind size on the copper and gold metallurgical performance. The initial open circuit batch flotation tests conducted in 2010 indicated that the upper Iron Cap mineralization was not sensitive to the primary grind sizes ranging from 80% passing 120 µm to 170 µm. However, the later test results show that a finer primary grind size would improve copper and gold recoveries. In addition, the recent KM6004 Iron Cap test work (2019) indicates that a lower rougher flotation pulp density (approximately 22% w/w) would improve the overall copper and gold metallurgical performance on some samples, compared to a “standard” slurry pulp density of approximately 35% to 40% w/w.

KM6004 in 2019 further tested the effect of the primary grind size on copper and gold flotation. The effect of slurry density on the target metal recovery was also investigated. The test results for copper and gold metallurgical performance are summarized in Figure 13.14 and Figure 13.15 respectively.

The test results indicate that a finer primary grind size would produce better copper and gold recoveries. The rougher flotation copper and gold recovery metallurgical performance was better in a diluted flotation slurry density (% solids) of approximately 22% w/w, compared to a standard rougher flotation slurry density of approximately 35%. This may be due to high amounts of mica and alteration products of potassium feldspar, such as sericite in some of the Iron Cap samples. These mineral components impact flotation pulp viscosities and modify froth cleaning abilities, thereby producing poor flotation separations and lower concentrate grades. The change to lower rougher flotation slurry densities resulted in a 30% to 40% reduction in rougher flotation weight.

Figure 13.14 Effect of Slurry Solid Density and Primary Grind Size – Rougher Flotation – Copper – KM6004

Source: ALS Metallurgy (2020). Preliminary Metallurgical Testing – Iron Cap Zones – Composites IC-2018-13 to IC-2018-18 – KSM Project (KM6004). January 8, 2020.

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Figure 13.15 Effect of Slurry Solid Density and Primary Grind Size – Rougher Flotation – Gold – KM6004

Source: ALS Metallurgy (2020), Preliminary Metallurgical Testing – Iron Cap Zones – Composites IC-2018-13 to IC-2018-18 – KSM Project (KM6004), January 8, 2020.

It should be noted that Composite IC-2018-06 from the Mylonite zone, which was tested in 2018 (KM5806), shows significantly inferior metallurgical performance compared to the other samples. As indicated by Seabridge geological team, the sample represents a very small portion of the deposit (a geologically distinct unit located in the upper portion of the deposit as a veneer near the intersection of the Sulphurets and North Iron Cap faults). Therefore, no detailed testing, including process condition optimization testing, has been conducted on this sample.

Following the parameters established from the batch open circuit flotation metallurgical testing, more LCTs were carried out in all the additional testing programs (KM5248, KM5501, KM5806, and KM6604). Figure 13.16 shows the flowsheet used by these test programs.

Figure 13.16 LCT Flotation Flowsheet – Iron Cap (2017 – 2020 ALS)

Source: ALS, 2017-2020

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The locked cycle flotation test results are summarized in Table 13.37. Two tests were repeated in test program KM5806 in order to improve the copper grade in the bulk concentrates. In the test program KM6004, the initial LCTs (KM6004-21 to KM6004-23) were conducted at a normal solid content of 35% w/w; while the remaining LCTs were performed approximately 22% by weight to improve the metallurgical performance. Some of the Iron Cap samples contained high amounts of mica and alteration products of potassium feldspar, such as sericite. These clay-related mineral components caused an increase in flotation pulp viscosity and a reduction in froth cleaning ability. In addition, a high solid density resulted in poor flotation separation efficiencies and lower concentrate grades. The change to lower rougher flotation slurry densities resulted in a 30% to 40% reduction in rougher flotation weight. The mass pull reduction is expected to reduce the amount of energy required for fine regrinding of rougher concentrates prior to the flotation cleaning stages.

Major observations of the results from the 21 LCTs are as follows:

| ● | the tested head samples contain varied contents of copper, gold, and silver, which range between 0.18% and 0.59% Cu, 0.20 g/t and 1.03 g/t Au, and 1 g/t and 6 g/t Ag, respectively

| ● | on average, approximately 90% copper has been recovered to the bulk concentrate product grading at 24.7% Cu. Approximately, 62% gold and 55% silver have been recovered into the flotation concentrate

| ● | the bulk cleaner scavenger tailing still contain significant gold and silver contents which, on average, represent 23% gold at a grade of 1.06 g/t Au and 21% silver at a grade of 4 g/t Ag. Additional cyanidation leaching tests have been conducted to further recover the contained gold and silver and are discussed in the following section

| ● | on average, the gold-pyrite concentrate and copper first cleaner tailing contain approximately 4.2% gold at 0.43 g/t Au and 3.7% silver at 2 g/t Ag, respectively. Similarly, the metallurgical response of the product to cyanidation leaching is discussed in the following section.

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Table 13.37 Flotation Locked Cycle Test Results – Iron Cap 2017 -2020 (KM5248/KM5501/KM5806/KM6004)

Test* | Sample | Grind Size (P80 µm**) | Head Grade | Bulk Conc. Grade | Recovery
Cu (%) | Au (g/t) | Ag (g/t) | Mo (%) | Cu (%) | Au (g/t) | Ag (g/t) | Mo (%) | Cu (%) | Au (%) | Ag (%) | Mo (%)
KM5248-10 | IC2016-02 | 87/13/13 | 0.37 | 0.45 | 6 | 0.006 | 24.0 | 20.9 | 245 | 0.300 | 86.9 | 63.1 | 59.8 | 66.6
KM5248-11 | IC2016-03/04 | 88/13/12 | 0.20 | 1.03 | 3 | 0.002 | 22.2 | 85.5 | 264 | 0.113 | 82.0 | 61.6 | 61.1 | 54.6
KM5501-16 | IC-MC-A | 123/19/18 | 0.32 | 0.68 | 3 | 0.002 | 22.9 | 34.4 | 151 | 0.094 | 88.3 | 62.3 | 67.3 | 56.5
KM5501-17 | IC-MC-B | 111/17/16 | 0.59 | 0.58 | 4 | 0.007 | 26.8 | 18.9 | 137 | 0.193 | 92.2 | 66.8 | 77.1 | 54.3
KM5501-19 | IC-MC-C | 116/16/17 | 0.43 | 0.64 | 3 | 0.004 | 27.0 | 28.5 | 156 | 0.132 | 90.0 | 64.2 | 68.2 | 43.6
KM5806-18 | IC-2018-01 | 100/22/19 | 0.18 | 0.20 | 1 | 0.006 | 21.6 | 13.9 | 86 | 0.590 | 88.4 | 51.8 | 42.4 | 74.2
KM5806-19 | IC-2018-02 | 95/16/14 | 0.35 | 0.33 | 2 | 0.003 | 23.1 | 14.0 | 60 | 0.125 | 90.8 | 59.1 | 52.4 | 60.5
KM5806-20 | IC-2018-03 | 108/20/17 | 0.42 | 0.69 | 2 | 0.003 | 23.9 | 28.3 | 67 | 0.105 | 92.0 | 67.1 | 59.7 | 63.5
KM5806-21 | IC-2018-04 | 100/15/14 | 0.27 | 0.35 | 1 | 0.005 | 25.5 | 26.5 | 67 | 0.386 | 92.1 | 74.8 | 50.1 | 83.2
KM5806-22 | IC-2018-05 | 108/19/21 | 0.50 | 0.75 | 2 | 0.002 | 25.2 | 27.0 | 64 | 0.074 | 92.4 | 65.7 | 48.7 | 55.5
KM5806-23 | IC-2018-02 | 95/17/16 | 0.35 | 0.31 | 2 | 0.003 | 23.9 | 13.2 | 66 | 0.135 | 90.3 | 56.3 | 51.9 | 60.0
KM5806-24 | IC-2018-03 | 108/19/22 | 0.43 | 0.69 | 2 | 0.002 | 25.1 | 28.8 | 68 | 0.107 | 90.7 | 64.7 | 61.5 | 76.2
KM6004-21 | IC-2018-13 | 90/16/16 | 0.46 | 0.46 | 5 | 0.005 | 22.7 | 11.4 | 150 | 0.19 | 92.7 | 46.8 | 57.1 | 70.2
KM6004-22 | IC-2018-14 | 93/19/16 | 0.40 | 0.63 | 3 | 0.003 | 19.2 | 21.7 | 72 | 0.12 | 91.0 | 65.5 | 47.2 | 69.8
KM6004-23 | IC-2018-15 | 94/17/16 | 0.32 | 0.51 | 2 | 0.003 | 24.4 | 26.4 | 75 | 0.16 | 93.6 | 64.0 | 58.2 | 63.6
KM6004-34 | IC-2018-14 | 93/18/20 | 0.40 | 0.64 | 2 | 0.004 | 24.3 | 27.6 | 84 | 0.14 | 90.3 | 63.4 | 52.6 | 57.8
KM6004-35 | IC-2018-16 | 89/15/16 | 0.027 | 0.70 | 2 | 0.002 | 26.2 | 52.5 | 116 | 0.03 | 90.5 | 70.6 | 50.9 | 20.0
KM6004-44 | IC-2018-17 | 86/18/20 | 0.36 | 0.45 | 2 | 0.002 | 28.4 | 23.0 | 90 | 0.070 | 91.9 | 59.2 | 52.4 | 37.8
KM6004-45 | IC-2018-18 | 89/16/17 | 0.35 | 0.53 | 2 | 0.003 | 27.6 | 26.1 | 109 | 0.127 | 91.3 | 57.4 | 56.5 | 59.0
KM6004-58 | IC-2018-14 | 152/16/16 | 0.40 | 0.60 | 3 | 0.004 | 27.6 | 29.7 | 102 | 0.16 | 86.7 | 63.0 | 50.8 | 50.4
KM6004-59 | IC-2018-15 | 152/16/16 | 0.32 | 0.46 | 2 | 0.003 | 28.5 | 30.8 | 85 | 0.12 | 88.0 | 65.9 | 47.5 | 37.4

| Notes: | *Au grade in the Bulk Cleaner Tailings ranged from 0.53 to 1.80 g/t Au, averaging at 1.05 g/t Au; Au grade in the pyrite concentrate ranged from 0.05 to 0.80 g/t Au, averaging at 0.42 g/t Au.

**Primary grind size/regrind size

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Gravity and Flotation Tests on Mylonite Samples

A series of gravity concentration and flotation tests has been carried out on the Mylonite sample (Composite IC-2017-08), whose sample was from the area close to the surface (out of the potential upper panel caves). It contained high gold and low copper (approximately 0.01% Cu) contents. The sample had significantly different mineralogical characteristics. The gold recovery flowsheet is shown in Figure 13.17, including gravity concentration by centrifugal concentration followed by panning. The tests have been carried out at two primary grind sizes: 80% passing 80 µm and 149 µm. The gold recoveries to the panned concentrates were 18.4% and 24.5%. The gold grades of the gravity concentrates were low, only at 87.4 g/t and 64.7 g/t Au. This may imply that most of the gold in the mineralization might have not occurred in coarse free gold forms. The subsequent flotation tests on the gravity tailing have been performed on the gravity tailing produced from the head sample with a grind size of 80% passing 149 µm. The test results show a poor metallurgical response to the flotation concentration.

Figure 13.17 Gold Recovery Flowsheet

Source: ALS, 2017

Leach Tests

Leaching Tests on Main Copper Sulphide Mineralization Samples

Cyanidation leach tests including DCN leaching and CIL tests have been conducted by ALS Metallurgy during 2017 and 2020. The bulk cleaner scavenger tailing and copper first cleaner tailings produced from the locked cycle testing programs of KM5248, KM5501, KM5806, and KM6004 have been tested.

Similar to the leaching conditions of the Deep Kerr leach test, the direct leaching procedure involves 16-hour pre-aeration and 24-hour cyanide leaching stages. Sodium cyanide was added to maintain an initial concentration of 1,000 ppm for test programs KM5248, KM5501, and KM5806, which was increased to 2,000 ppm in the latest test program KM6004. With the CIL tests, activated carbon was added after the first 6-hour leaching for the three earlier programs; in the test program KM6004, activated carbon was added following the pre-aeration stage.

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All the cyanide leach test results are provided in Table 13.38. The following conclusions could be made based on the leach test results:

| ● | gold extractions from the bulk cleaner scavenger tailing were 35% to 70% with an average of 51%. On average, about 1.9 kg/t of sodium cyanide and 3.4 kg/t of lime were consumed in the leaching process

| ● | gold extractions to the leach feed from the copper first cleaner tailing were 40% to 73% with an average of 53%. On average, about 1.6 kg/t of sodium cyanide and 2.6 kg/t of lime were consumed

| ● | the average silver extraction rates were approximately 65% for the bulk cleaner scavenger tailing and 66% for the copper first cleaner tailing

| ● | for the composite samples constructed from both the tailing products, gold extraction rates were slightly lower, ranging between 40 and 44%

| ● | the CIL tests show that, on average, less than 80% of the gold extracted was adsorbed onto the activated carbon, ranging from 42% to 98%. The low gold adsorption onto the activated carbon might be due to significant copper dissolution in the leach treatment, which might have been caused the inferior gold adsorption onto the activated carbon.

Table 13.38 Preliminary Cyanidation Test Results – Iron Cap (ALS 2017)

Test | Test Type | Cyanide Consumption (g/t) | Lime Consumption (g/t) | Feed Grade, Calculated | Extraction
Au (g/t) | Ag (g/t) | Cu (%) | Au (%) | Ag (%) | Cu (%)
Bulk Cleaner Scavenger Tailing Cyanidation
KM5248-12 | DCN | 2,800 | 3,900 | 0.91 | 11.45 | 0.17 | 70.3 | 57.2 | 55.8
KM5248-13 | DCN | 2,600 | 3,700 | 1.99 | 8.50 | 0.18 | 50.8 | 49.4 | 54.6
KM5501-21 | CIL | 1,900 | 3,000 | 1.43 | 4.78 | 0.07 | 51.9 | 60.3 | 21.9
KM5501-22 | CIL | 2,700 | 4,000 | 1.30 | 4.06 | 0.06 | 51.7 | 58.1 | 40.5
KM5501-23 | CIL | 2,300 | 4,100 | 1.45 | 4.22 | 0.08 | 57.9 | 66.8 | 37.3
KM5806-18 | DCN | 1,100 | 5,400 | 1.2 | 4.0 | 0.05 | 60.8 | 64.7 | 29.8
KM5806-19 | DCN | 900 | 4,900 | 1.1 | 3.5 | 0.06 | 48.6 | 62.9 | 25.2
KM5806-20 | DCN | 800 | 2,800 | 1.2 | 3.3 | 0.06 | 43.2 | 63.8 | 26.5
KM5806-21 | DCN | 1,100 | 3,100 | 0.9 | 2.3 | 0.06 | 63.3 | 65.7 | 34.3
KM5806-22 | DCN | 900 | 3,700 | 1.4 | 3.7 | 0.07 | 43.5 | 70.1 | 29.0
KM6004-36 | CIL | 2,000 | 3,600 | 1.17 | 5 | 0.07 | 34.6 | 57.0 | 40.1
KM6004-37 | CIL | 1,300 | 2,400 | 1.07 | 3 | 0.08 | 40.0 | 68.5 | 20.1
KM6004-39 | CIL | 2,500 | 2,700 | 1.29 | 4 | 0.11 | 45.1 | 73.2 | 39.5
KM6004-38 | CIL | 1,500 | 2,400 | 0.76 | 2 | 0.05 | 43.4 | 74.3 | 32.1
KM6004-52 | CIL | 2,500 | 2,300 | 1.21 | 3 | 0.08 | 54.5 | 65.7 | 43.2
KM6004-53 | CIL | 2,600 | 3,100 | 1.09 | 5 | 0.09 | 58.2 | 76.8 | 43.4
KM6004-54 | CIL | 3,100 | 3,200 | 1.03 | 4 | 0.09 | 42.0 | 70.6 | 49.1

table continues…

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Test | Test Type | Cyanide Consumption (g/t) | Lime Consumption (g/t) | Feed Grade, Calculated | Extraction
Au (g/t) | Ag (g/t) | Cu (%) | Au (%) | Ag (%) | Cu (%)
Copper First Cleaner Tailing Cyanidation (Au-Pyrite Concentrate)
KM5248-14 | DCN | 1,800 | 2,000 | 0.29 | 2.60 | 0.10 | 72.5 | 73.1 | 64.6
KM5248-15 | DCN | 1,700 | 2,200 | 0.73 | 3.28 | 0.08 | 53.2 | 60.3 | 68.3
KM5501-24 | CIL | 1,600 | 1,200 | 0.46 | 1.43 | 0.03 | 56.9 | 65.1 | 22.4
KM5501-25 | CIL | 2,300 | 1,300 | 0.48 | 1.18 | 0.05 | 57.0 | 74.6 | 45.3
KM5501-26 | CIL | 2,000 | 1,300 | 0.52 | 1.64 | 0.05 | 54.2 | 69.4 | 43.8
KM5806-18 | DCN | 1,200 | 4,800 | 0.87 | 3.0 | 0.04 | 63.8 | 63.6 | 33.6
KM5806-19 | DCN | 1,000 | 4,400 | 0.77 | 2.0 | 0.04 | 51.9 | 60.0 | 36.7
KM5806-20 | DCN | 1,000 | 4,600 | 0.84 | 2.1 | 0.04 | 49.9 | 62.5 | 39.9
KM5806-21 | DCN | 800 | 4,400 | 0.42 | 1.0 | 0.03 | 60.5 | 71.2 | 43.5
KM5806-22 | DCN | 1,100 | 3,700 | 0.82 | 2.3 | 0.04 | 40.0 | 60.7 | 37.6
KM6004-40 | CIL | 1,000 | 2,200 | 0.26 | 2 | 0.03 | 65.8 | 70.0 | 37.5
KM6004-41 | CIL | 900 | 2,200 | 0.32 | 2 | 0.02 | 51.3 | 44.4 | 38.5
KM6004-43 | CIL | 1,700 | 2,600 | 0.50 | 2 | 0.03 | 56.9 | 68.0 | 69.5
KM6004-42 | CIL | 1,100 | 2,200 | 0.38 | 2 | 0.04 | 45.7 | 75.3 | 30.6
KM6004-55 | CIL | 2,400 | 1,800 | 0.46 | 1 | 0.04 | 54.7 | 71.5 | 53.2
KM6004-56 | CIL | 2,100 | 2,100 | 0.54 | 2 | 0.05 | 54.4 | 66.4 | 40.9
KM6004-57 | CIL | 2,200 | 2,500 | 0.55 | 2 | 0.04 | 42.0 | 78.2 | 52.0
Combined Tailing Cyanidation (Selected Bulk Cleaner Scavenger Tailing+ Copper First Cleaner Tailing)
KM5501-32 | DCN | 1,200 | 2,300 | 0.98 | 4.0 | 0.06 | 40.0 | 62.7 | 37.4
KM6004-67 | CIL | 2,600 | 2,000 | 0.74 | 3 | 0.07 | 43.6 | 58.4 | 42.0
KM6004-68 | CIL | 1,900 | 2,000 | 0.72 | 3 | 0.07 | 40.4 | 56.4 | 39.8

Leaching Tests on Mylonite Samples

To recover the gold from the Mylonite samples (Composite IC-2017-08), both whole ore cyanidation and cyanidation on flotation concentrate treatments have been tested.

The whole ore cyanidation leach tests show that with a 24-hour cyanide leaching, less than 5% of the gold was extracted at the two cyanide concentrations of 1,000 ppm and 2,000 ppm. Roasting was then conducted on the leaching residues, followed by additional cyanide leaching. The overall gold extraction was improved to 75%. This indicates that a significant portion of gold in the composite sample was refractory.

The cyanidation leach tests on the rougher concentrates were conducted for 48 hours, after ultra fine regrinding treatment to a grind size of 80% passing 5 µm. Only 26% of the gold in the concentrate was extracted with very high cyanide and lime consumptions. The silver and copper extractions were higher: 57% Ag and 63% Cu. The test results further indicate the refectory gold cyanidation properties of the tested samples.

A diagnostic leach of the scavenger rougher tailing from the flotation test (KM5501-18) shows that only approximately 3% of the gold in the tailing samples was cyanide leachable and approximately 83% of the gold was associated with the hydrochloric and nitric acid soluble minerals.

| 13.5.3 | Flotation Concentrate Assay (2007–2020)

The multi-element assay data are presented in Table 13.39 and Table 13.40 for the concentrates from various deposits. On average, the impurities in the copper-gold concentrates produced from the Mitchell, Sulphurets, Iron Cap, and Kerr deposits should not incur smelting penalties as set out by most smelters.

However, arsenic, antimony, and mercury contents in some of the concentrates from the Iron Cap deposit and the Kerr samples may incur smelting penalties. Also, the lead content of the concentrate from the Iron Cap Composite 1 may be higher than the penalty thresholds. Fluorine levels in some of the concentrates may be also higher than the penalty thresholds. It is anticipated that the mill feeds will be supplied from different deposits and mineralization zones. Impurity contents in the copper concentrates produced from these blended mill feeds should be lower than the penalty thresholds set by most of the smelters. Further review with respect to smelting penalties should be conducted.

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Table 13.39 Flotation Concentrate Assay from Different Deposit Samples

Element | Unit | Mitchell | Sulphurets | Kerr | Iron Cap
Min | Max | Avg | Min | Max | Avg | Min | Max | Avg | Min | Max | Avg
Copper (Cu) | % | 22.0 | 28.0 | 25.0 | 26.0 | 29.3 | 27.9 | 21.0 | 30.7 | 25.6 | 21.7 | 27.0 | 24.4
Gold (Au) | g/t | 44.2 | 98.2 | 65.1 | 31.4 | 63.7 | 45.6 | 4.1 | 32.9 | 12.1 | 10.9 | 146.8 | 35.9
Silver (Ag) | g/t | 223 | 431 | 303 | 34 | 130 | 82 | 32 | 192 | 69 | 59 | 774 | 180
Molybdenum (Mo) | % | 0.12 | 0.72 | 0.35 | 0.17 | 0.70 | 0.37 | 0.01 | 0.27 | 0.11 | 0.07 | 0.59 | 0.20
Iron (Fe) | % | 26.8 | 32.7 | 30.0 | 27.2 | 29.6 | 28.7 | 23.7 | 32.3 | 29.3 | 25.4 | 30.0 | 28.0
Zinc (Zn) | % | 0.23 | 0.43 | 0.33 | 0.18 | 0.54 | 0.34 | 0.01 | 0.94 | 0.30 | 0.02 | 2.67 | 0.78
Arsenic (As) | ppm | 690 | 2,080 | 1,181 | 205 | 1,768 | 732 | 143 | 3,276 | 1,575 | 91 | 6,140 | 1,716
Selenium (Se) | ppm | 59 | 102 | 75 | 118 | 118 | 118 | 109 | 247 | 205 | 108 | 400 | 255
Antimony (Sb) | ppm | 210 | 1,182 | 647 | 370 | 2,100 | 972 | 24 | 2,710 | 801.8 | 22 | 4,500 | 1,653
Mercury (Hg) | ppm | 0.6 | 6.8 | 2.8 | 2.0 | 2.0 | 2.0 | 3.4 | 88.0 | 25.5 | <1 | 25.0 | 6.2
Lead (Pb) | % | 0.12 | 0.92 | 0.30 | 0.19 | 0.72 | 0.39 | 0.02 | 1.33 | 0.12 | 0.06 | 1.31 | 0.43
Bismuth (Bi) | ppm | <10 | 150.0 | 76.0 | <10 | <10 | <10 | 3.1 | 105.0 | 21.5 | 5.9 | 205.0 | 47.4
Fluoride (F) | ppm | 69 | 346 | 168 | 155 | 155 | 155 | 20 | 264 | 179 | 162 | 494 | 347
Sulphur (S) | % | 31.1 | 38.1 | 34.4 | 32.4 | 34.4 | 33.5 | 27.1 | 38.4 | 35.4 | 32.3 | 36.0 | 34.2
Chloride (Cl) | % | <0.01 | <0.01 | <0.01 | <0.01 | <0.01 | <0.01 | <0.01 | <0.01 | <0.01 | <0.01 | <0.01 | <0.01

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Table 13.40 Flotation Concentrate Assay from Blended Head Samples

Element | Unit | Mitchell/ Sulphurets | Mitchell/ Upper Kerr | Mitchell/ Deep Kerr | Mitchell/Iron Cap
Min | Max | Avg
Copper (Cu) | % | 24.2 | 25.3 | 24.5 | 25.0 | 25.3 | 25.2
Gold (Au) | g/t | 52.0 | 40.0 | 28.3 | 40.0 | 54.3 | 47.2
Silver (Ag) | g/t | 178 | 168 | 150 | 168 | 332 | 250
Molybdenum (Mo) | % | 0.66 | 0.06 | 0.33 | 0.06 | 0.45 | 0.26
Iron (Fe) | % | 30.0 | 29.3 | 29.6 | 29.3 | 30.8 | 30.1
Zinc (Zn) | % | 0.92 | 0.42 | 0.24 | 0.4 | 0.7 | 0.6
Arsenic (As) | ppm | 969 | 1,369 | 1,404 | 1,369 | 1,970 | 1,670
Selenium (Se) | ppm | 89 | 76 | - | 76 | 76 | 76
Antimony (Sb) | ppm | 500 | 492 | 814 | 492 | 2,116 | 1,304
Mercury (Hg) | ppm | 1.0 | 2.4 | 23.0 | 2.4 | 7.0 | 4.7
Lead (Pb) | % | 0.26 | 0.15 | 0.12 | 0.2 | 0.5 | 0.3
Bismuth (Bi) | ppm | <10 | 121.0 | 12.2 | 39.7 | 121.0 | 80.4
Fluoride (F) | ppm | 174 | 116 | 264 | 116 | 265 | 191
Sulphur (S) | % | 34.9 | 35.0 | 34.6 | 35.0 | 35.5 | 35.3
Chloride (Cl) | % | <0.01 | - | <0.01 | - | - | -

ALS Environmental has conducted additional selenium testing on Iron Cap samples for selenium deportment and distributions on the flotation products, as well as for the effects of cyanidation of selenium extraction. The selenium analysis results are summarized as follows:

| ● | about 21% selenium was present in the bulk concentrate at a grade of 114 g/t Se; very little selenium (<1 ppm) was dissolved into the process water of the rougher and cleaner tailings

| ● | a low percentage (about 3%) of selenium in the leach feed was extracted during the cyanidation leach. During the destruction of the cyanide, the residual selenium concentration was 0.186 mg/L as compared with 0.221 g/t Se in the non-treated cyanide leaching residue. The selenium content can be effectively removed by using reverse osmosis (RO) methods.

| 13.5.4 | Ancillary Tests

During test programs, various environment-related tests have been conducted to determine engineering-related parameters. The key tests are as follows:

| ● | leach residue cyanide destruction, including sulphur dioxide/air, Caro’s acid (H2SO5), and hydrogen peroxide (H2O2)

| ● | cyanide recovery from barren solutions, including acidification, volatilization of hydrogen cyanide gas, and re-neutralization (AVR); and sulphidization, acidification, recycling, and thickening of precipitate (SART)

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| ● | static and dynamic thickening tests for conventional thickener sizing and for high rate thickener sizing for primary grinding product, first cleaner tailing together with gold-bearing pyrite concentrate, cyanidation residues, and rougher/scavenger flotation tailing

| ● | filtration testing, including vacuum filtration and pressure filtration for bulk flotation concentrate.

Cyanide Recovery Tests & Cyanide Destruction Tests – 2009/2010 (SGS)

A large-scale, agitated bulk cyanide leach test was conducted by SGS on a 20 kg combined sample of first cleaner tailing and pyrite rougher concentrate. The sample was sourced from material generated from the flotation pilot plant testing at G&T.

The key chemical analysis of the solution for cyanide recovery and the washed leach pulp for the cyanide destruction are shown in Table 13.41.

Table 13.41 Chemical Analysis of Cyanide Recovery Test Solution and Cyanide Destruction Pulp

Sample | CNT (mg/L) | CNWAD (mg/L) | CNF (mg/L) | Cu (mg/L) | Fe (mg/L) | CNS (mg/L)
Leach Solution | 853 | 850 | 280 | 562 | 1.6 | 700
Washed Pulp | 94 | 90 | - | 90.4 | 1.08 | 220

Note: CNT—total cyanide; CNWAD—weak acid dissociable cyanide; CNF—free cyanide; CNS—thiocyanate

Exploratory AVR tests were conducted to investigate the effect of pH on the recovery of cyanide from the barren leach solution. The scrubbing retention time was 4 hours. The collected cyanide, acid consumption, and lime consumption are summarized in Table 13.42.

Table 13.42 Cyanide Recovery Test Results – AVR

pH | Recovered CNWAD (%) | Sulphuric Acid Addition (g/L) | Hydrated Lime Addition (g/L)
2 | 77 | 3.18 | 0.78
3 | 72 | 2.01 | 0.24
4 | 35 | 1.14 | 0.16

Exploratory SART tests were also conducted on the barren leach solution to investigate the effects of pH and sodium hydrosulphide (NaHS) dosage on recovering cyanide and copper from CNWAD and copper cyanide complexes. The test results are as follows:

| ● | at a sodium hydrosulphide dosage of 100% stoichiometric requirement, 83% to 94% of the copper was precipitated when reducing the pH level from 5 to 3

| ● | at pH 3, an increase of sodium hydrosulphide dosage to 120% of the stoichiometric requirement resulted in near complete removal of copper from the solution and regeneration of all the weak acid dissociable cyanide as free cyanide

| ● | the sulphuric acid addition was approximately 1.9 g/L of feed solution, and the hydrated lime requirement for re-neutralization of the SART-treated solution was 1.3 g/L of feed solution.

Further optimization of the SART conditions could improve these results, if SART was considered for recovery of cyanide into low-concentration cyanide solutions. These SART-generated cyanide solutions might also be considered for feed to further AVR processing to generate higher grade cyanide solutions for recycle to the leaching circuits.

Cyanide Destruction Tests – 2009/2010 (SGS)

Three different cyanide destruction methods — sulphur dioxide/air, Caro’s acid, and hydrogen peroxide — were tested for oxidation of cyanide and detoxification of the washed pulp. The objective of the test work was to produce treated effluent containing less than 2 mg/L CNWAD. The results of the cyanide destruction test results are summarized in Table 13.43.

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Table 13.43 Cyanide Destruction Test Results – 2009/2010 (SGS)

Test | Method | Oxidant Dosage Stoich (%) | Cumulative Retention Time (~h) | Composition (Solution Phase) | Cumulative Reagent Addition* (g/g CNWAD)
pH | CNT mg/L | CNWAD mg/L | SO2 Equivalent | Lime | Cu | H2SO5 100% | H2O2 100% | Cu mg/L Solution
Cyanidation Washed Pulp | 10.7 | 94 | 90 | - | - | - | - | - | -
CND 6&7 | SO2/Air | 160-200 | 1 | 9.6 | 2-4 | <1 | 4-5 | - | 0.14 | - | - | 12
C-1 | Caro’s Acid | 500 | 1.5** | 9.0 | 2.8 | 1.7 | - | 37 | - | 21.9 | - | -
H-1 | H2O2 | 500 | 1.5** | 10.1 | 12 | 11 | - | - | - | - | 6.5 | -
SO2/Air Partially Treated Pulp | 10.0 | 10 | 10 | - |  | - | - | - | -
C-2 | Caro’s Acid | 500 | 1.5** | 9.0 | 2.8 | 1.7 | - | - | - | 21.6 | - | -
H-7 | H2O2 | 1,000 | 0.5 | 10.0 | 2.3 | 0.3 | - | - | 1.5 | - | 13 | 15
SO2/Air Partially Treated Solution | 10.0 | 10 | 10 | - | - | - | - | - | -
H-4 | H2O2 | 500 | 1 | 8.7 | 1.6 | 0.4 | - | - | - | - | 6.5 | -

| Notes: | *Copper was added as CuSO4 5H2O; SO2 was added as Na2S2O5

**Reagent was added in three 30-min stages

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The results indicated that the residual CNWAD in the washed pulp was reduced to less than 1 mg/L after the pulp had been treated with 4 g to 5 g equivalent sulphur dioxide and 0.14 g copper (added as copper sulphate) per gram of CNWAD in the pulp. The reaction time for this process was one hour at the natural pH. The sulphur dioxide/air-treated pulp contained small amounts of CNT in the form of ferrocyanide complex.

An exploratory test indicated that the residual CNWAD in the solution phase of the washed pulp was reduced to less than 2 mg/L level by using Caro’s acid treatment. The reagent consumption was 0.74 g/L H2SO5 (250% of the stoichiometric amount) and 0.6 g/L hydrated lime of the feed to the cyanide destruction.

The tests also indicated that the hydrogen peroxide process is not very efficient for cyanide destruction. The residue CNWAD was only reduced from 90 mg/L to 11 mg/L after adding 500% of the stoichiometrically required hydrogen peroxide.

Two-stage cyanide destruction involving sulphur dioxide/air treatment followed by a polishing treatment with Caro’s acid or hydrogen peroxide was investigated on the pulp and also on a tailing filtrate solution. The sulphur dioxide/air-treated pulp was adjusted with sodium cyanide to 10 mg/L CNWAD for the polishing tests. The results are as follows:

| ● | the polishing test using Caro’s acid was unsuccessful. The final product still contained 3.2 mg/CNWAD after adding 500% of the stoichiometric Caro’s acid

| ● | the hydrogen peroxide polishing treatment produced less than 2 mg/L residual CNWAD. The hydrogen peroxide dosage was 10 times of the stoichiometric requirement and the copper addition was 0.011 g/L pulp

| ● | the solution phase (filtrate) of the sulphur dioxide/air partially treated pulp responded well to the hydrogen peroxide polishing treatment. The solution contained less than 1 mg/L residual CNWAD after being treated with five times the stoichiometric hydrogen peroxide requirement (0.065 g/L solution). Copper sulphate was not used in the treatment of this solution.

Cyanide Recovery Tests & Cyanide Destruction Tests – 2017 (BQE Water)

Six leaching residue samples from the test program KM5367 on the Mitchell zone have been tested for cyanide destruction and copper removal using sulphur dioxide/air method. The copper removal was completed after the cyanide destruction to achieve the effluent copper concentration target of 0.5 mg/L.

The results confirmed that the cyanide destruction was successful by reducing the WAD cyanide concentration in the effluent to a level between 0.005 mg/L and 0.075 mg/L, as compared with the target limit of 0.5 ppm CNWAD.

Settling Tests

Preliminary settling tests were conducted on pyrite rougher flotation tailing in the 2008 testing program. As reported by G&T, the tests on the tailing slurry failed to generate normal settling curves. The tests were subsequently carried out on the re-pulped sample from dried tailing.

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The test data reveal that the settling area required for pyrite rougher flotation tailing was 0.73 m2/t/d without adding flocculant and 0.30 m2/t/d with the addition of 10 g/t of flocculant.

In 2009, Pocock conducted solids liquid separation (SLS) tests on five flotation products generated by G&T from the bench scale tests and pilot plant tests. The materials tested included flotation feed, copper concentrate, first cleaner tails and gold-bearing pyrite concentrate, cyanidation residues, and rougher/scavenger flotation tailing. The dewatering tests included:

| ● | flocculant screening tests

| ● | static and dynamic thickening tests for conventional thickener sizing and for high rate thickener sizing

| ● | viscosity (rheological properties) tests for rake mechanism and underflow pipeline sizing

| ● | vacuum filtration tests

| ● | pressure filtration tests.

Hychem AF 303 (a medium to high molecular weight, 7% charge density, anionic polyacrylamide) was selected for thickening tests from preliminary screening of a series of flocculants.

The key test results are summarized in Table 13.44 and Table 13.45.

Table 13.44 Recommended Conventional Thickener Operating Parameters – 2009 (Pocock)

Tested Material | Feed (% Solids) | Flocculant (g/t) | Underflow (% Solids) | Unit Area (m2/t/d)
Flotation Feed Composite | 20-25 | 10-15 | 60-65 | 0.125
Coarse Grind Flotation Feed | 25-30 | 10-15 | 70-74 | 0.125
Final Copper Concentrate | 25-30 | 5-10 | 70-72 | 0.125
Rougher Tailing | 15-20 | 10-15 | 60-62 | 0.125
Au-Pyrite Concentrate and Cu Cleaner Tailing | 15-20 | 20-25 | 55-58 | 0.275-0.307
Cyanide Leach Reside | 10-15 | 20-25 | 50-53 | 0.284-0.312

| Notes: | –All tests were performed at 20°C and as received pH. –Hydraulic loading or rise rate (m3/m2/h) includes a 0.5 scale-up factor. –Unit area includes a 1.25 scale-up factor; the range of unit areas provided corresponds to the range of underflow densities. –Coarse grind flotation feed: at a particle size of P80 170 µm; simulating stage one primary grind size.

Table 13.45 Recommended High Rate Thickener Operating Parameters – 2009 (Pocock)

Tested Material | Feed (% Solids) | Flocculant (g/t) | Underflow (% Solids) | Net Feed Loading (m3/m2h)
Flotation Feed Composite | 15-20 | 15-20 | 60-65 | 4.8-6.1
Coarse Grind Flotation Feed | 20-25 | 10-15 | 70-74 | 4.8-6.1
Rougher Flotation Tailing | 15-20 | ~20 | 57-62 | 3.7-4.8

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Filtration Tests

The 2009 Pocock testing program also determined the filtration rates of the copper concentrates produced from G&T pilot plant tests. Both vacuum filtration and pressure filtration methods were tested. The test results are summarized in Table 13.46.

Table 13.46 Filtration Test Results – 2009 (Pocock)

Filtration Method | Bulk Cake Density (dry kg/m3) | Cake Thickness (mm) | Cake Moisture (%) | Filtration Rate (dry kg/m2h) | Dry Cake Weight (dry kg/m2)
Vacuum | 1,785 | 15 | 19 | 265* | -
Pressure | 2,511 | 51 | 8 | - | 117.8**

Notes: *Includes scale-up factors at vacuum of 67.7 kPa.
**Feed pressure 552 kPa at 51 mm thickness.

Magnetic Separation Tests

In the 2008 test program, Davis Tube magnetic separation was used in an effort to recover the metal contents lost in the coarser than 74 µm fraction of the pyrite flotation tailing from Tests 10, 11, and 25. Test results indicate that less than 3% of the coarse tailing weight was recovered into a magnetic fraction assaying approximately 23% iron. No copper or gold assay data was reported.

| 13.6 | Conclusions

Extensive metallurgical testing programs were conducted on the samples representing the four deposits of the Property in 1989-1991 and 2007-2020.

The substantial test results indicate that the mineral samples from the four separate deposits are amenable to conventional flotation. In general, the Mitchell, Sulphurets, and parts of the Kerr and Iron Cap mineralization responds well to additional gold and silver recovery from the gold-bearing sulphide products by cyanidation. However, the test results show that the gold-bearing sulphide products (first cleaner scavenger tailing and pyrite concentrate) from the Deep Kerr and lower Iron Cap zones did not seem to respond well to gold recovery by using the tested cyanide leaching treatment. Further test work is required to optimize the process conditions, or alternative treatment methods should be used for the gold and silver recovery. The flotation and cyanidation combined process consists of:

| ● | copper-gold-molybdenum bulk rougher flotation followed by gold-bearing pyrite flotation

| ● | regrinding of the resulting bulk rougher concentrate followed by three stages of cleaner flotation to produce a copper-gold-molybdenum bulk cleaner flotation concentrate

| ● | molybdenum separation of the bulk cleaner flotation concentrate to produce a molybdenum concentrate and a copper/gold concentrate containing associated silver, if the molybdenum grade is sufficiently high

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| ● | cyanide leaching of the gold-bearing pyrite flotation concentrate and the scavenger cleaner tailing to further recover gold and silver values as doré, mainly for Mitchell and Sulphurets mineralization and parts of Iron Cap and Kerr mineralization.

The testing programs from 2017 to 2020 show the following:

| ● | a finer primary grind size can improve the copper and gold metallurgical performance, especially for copper

| ● | rougher flotation at a low slurry solid density can improve copper and gold metallurgical performance, especially for the mineralization with more clay-type minerals

| ● | the test results suggest that the gold-bearing sulphide products (first cleaner scavenger tailing and pyrite concentrate) from the Deep Kerr and lower Iron Cap zones did not seem to respond well to the gold recovery by the established cyanide leaching treatment.

On average, the impurities in the copper-gold concentrates produced from the Mitchell, Sulphurets, and Kerr deposits should not incur smelting penalties as set out by most smelters.

However, arsenic, antimony, and mercury contents in some of the concentrates from the Iron Cap deposit and the Kerr samples may incur smelting penalties. It is anticipated that the mill will be supplied with blended feeds from different deposits. Impurity contents in the copper concentrates produced from these blended mill feeds should be lower than the penalty thresholds set by most of the smelters. Further review with respect to smelting penalties should be conducted.

The samples from all the deposits are moderately hard for ball mill and SAG mill grinding, excluding the samples from the Sulphurets deposit, which show much resistance to both ball mill and SAG mill grinding. Also the samples tested are amenable to particle size reduction by HPGR procedure.

| 13.7 | Metallurgical Performance Projection

The metallurgical test results obtained from the various test programs were used to predict the plant metallurgical performance parameters for copper, gold, silver, and molybdenum. Gold and silver recoveries are based on the combined process of flotation to produce a salable concentrate, followed by cyanidation of combined cleaner tailings and pyrite flotation concentrate. The flotation process will produce a copper concentrate containing approximately 25% copper with variable precious metal contents and a molybdenum concentrate with 50% molybdenum. The gold cyanidation process applied to gold-bearing pyrite products will produce gold-silver doré. This process further recovers gold and silver from the mineralization, in addition to the gold and silver recovered by the flotation process.

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The LCT results from the various test programs were used to project flotation metallurgical performance. All of the metal recoveries to the concentrates from the LCTs were adjusted to the target copper grade of the final concentrate. The gold and silver recoveries to the final doré were estimated based on the cyanidation test results including metal losses, which were expected to occur during downstream acid washing, desorption and smelting processing in industrial operation. Due to some variations in metallurgical responses among the mineralized deposits, the projections for their metallurgical performance were on deposit basis.

| 13.7.1 | Metallurgical Performance Projection – Mitchell, Sulphurets, Upper Kerr, and Upper Iron Cap

The results show that the Mitchell mineralization produced better metallurgical performance, compared to the Sulphurets, Kerr, and Iron Cap mineralization. The metallurgical performance projections of the different types of KSM mineralization are summarized in Table 13.47 to Table 13.50. The estimates are based on a primary grind size of 80% passing approximately 125 µm to 150 µm and a regrind size of 80% passing approximately 20 µm.

Table 13.47 Cu-Au Flotation Concentrate Grade Versus Cu Head Grade

Cu Head Grade (%) | Cu Concentrate Grade (%)
>0.80 | 28
0.40-0.80 | 26
0.15-0.40 | 25
0.10-0.15 | 23
0.05-0.10 | 17
<0.05 | 5

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Table 13.48 Cu-Au Flotation Concentrate – Metal Recovery Projections

Deposit | Description | Head Grade | Recovery
Mitchell | Copper Recovery | >1.0% Cu | = 95%
0.8 – 1.0% Cu | = 92%
0.234 – 0.8% Cu | = 90.86 x (Cu Head, %) 0.027
0.05 – 0.234% Cu | = 18.02 x ln(Cu Head, %) + 113.5
0.02 – 0.05% Cu | = 20%
<0.02% | = 3%
Gold Recovery | n/a | = 0.0967 x (Cu Recovery, %) 1.4465
Silver Recovery | n/a | = 1.427 x (Cu Recovery, %) – 70.11
Sulphurets | Copper Recovery | >1.0% Cu | = 93%
0.8 – 1.0% Cu | = 90%
0.234 – 0.8% Cu | = 90.86 x (Cu Head, %) 0.027 – 3.5
0.05 – 0.234% Cu | = 18.02 x ln(Cu Head, %) + 110
0.02 – 0.05% Cu | = 20%
<0.02% | = 3%
Gold Recovery | n/a | = 52.07 x ln(Cu Recovery, %) – 174.1
Silver Recovery | n/a | = 1.065 x (Cu Recovery, %) – 44.80; if copper recovery < 50%, use 5%
Kerr | Copper Recovery | >1.0% Cu | = 88%
0.8 – 1.0% Cu | = 85%
0.234 – 0.8% Cu | = 90.86 x (Cu Head, %) 0.027 – 7
0.05 – 0.234% Cu | = 18.02 x ln(Cu Head, %) + 106.5
0.02 – 0.05% Cu | = 20%
<0.02% Cu | = 3%
Gold Recovery | n/a | = 171.8 x ln(Cu Recovery, %) – 718; if copper recovery < 70%, use 5%
Silver Recovery | n/a | = 132.48 x ln(Cu Recovery, %) - 542.9; if copper recovery < 70%, use 5%
Iron Cap | Copper Recovery | >1.0% Cu | = 95%
0.8 - 1.0% Cu | = 92%
0.49 - 0.8% Cu | = 90%
0.10 - 0.49% Cu | = 90.786 x (Cu Head, %) 0.089
0.05 - 0.10% Cu | = 30%
<0.05% Cu | = 3%
Gold Recovery | >2.0 g/t Au | = 80%
0.75 – 2.0 g/t Au | = 72.5%
0.05 – 0.75 g/t Au | = 78.128 x (Au Head, g/t) 0.3012
<0.05 g/t Au | = 20
Silver Recovery | >20 g/t Ag | = 83%
10 – 20 g/t Ag | = 78%
0.5 – 10 g/t Ag | = 39.945 x (Ag Head, g/t) 0.2602
<0.5 g/t Ag | = 5%

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Table 13.49 Au-Ag Doré – Cyanide Leach Metal Recovery Projections

Deposit | Head Grade | Recovery
Mitchell | Gold
>10 g/t Au | = (98 – ( 0.096 x (Cu Recovery, %) 1.446)) x 80% x 98%
5 – 10 g/t Au | = (95 – ( 0.096 x (Cu Recovery, %) 1.446)) x 75% x 98%
0.1 – 5 g/t Au | = (87.491 x (Au Head, g/t)0.051 – ( 0.096 x (Cu Recovery, %) 1.446)) x 66% x 98%
<0.1 g/t Au | = 0%
Silver
>15 g/t Ag | = 88 – (1.427 x (Cu Recovery, %) – 70.11)
8- 15 g/t Ag | = 86 – (1.427 x (Cu Recovery, %) - 70.11)
1 - 8 g/t Ag | = (42.74 x (Ag Head, g/t) 0.336 ) - ( 1.427 x (Cu Recovery, %) - 70.11) ; if <0, use 0%
<1 g/t Ag | = 0%
Sulphurets | Gold
>10 g/t Au | = (98 – (52.07 x ln(Cu Recovery, %) – 174.1)) x 70% x 98%
5 - 10 g/t Au | = (95 – ( 52.07 x ln(Cu Recovery, %) – 174.1)) x 60% x 98%
0.1 - 5 g/t Au | = ((87.491 x (Au Head, g/t)0.051 +3)- (52.07 x ln(Cu Recovery, %) - 174.1)) x 49% x 98%
<0.1 g/t Au | = 0%
Silver
>15 g/t Ag | = 52.7%
8- 15 g/t Ag | = 50.7%
1 – 8 g/t Ag | = (42.74 x (Ag Head, g/t) 0.336) - (1.065 x (Cu Recovery, %) - 44.80); if <0, use 0%
<1 g/t Ag | = 0%
Kerr | Gold
>10 g/t Au | = (98 – (171.8 x ln(Cu Recovery, %) – 718)) x 75% x 98%
5 - 10 g/t Au | = (95 – (171.8 x ln(Cu Recovery, %) – 718)) x 65% x 98%
0.1 - 5 g/t Au | = ((87.491 x (Au Head, g/t)0.051 + 8)- (171.8 x ln(Cu Recovery, %) - 718))) x 57% x 98%
<0.1 g/t Au | = 0%
Silver
>15 g/t Ag | = (88 – (132.48 x ln(Cu Recovery, %) – 542.9))/100; Cap at 88%
8- 15 g/t Ag | = (86 – (132.48 x ln(Cu Recovery, %) – 542.9))/100; Cap at 86%
1 - 8 g/t Ag | = (21.59 x ln(Ag Head, g/t) + 40.14) - (132.48 x ln(Cu Recovery, %) - 542.9) ; if <0, use 0%
<1 g/t Ag | = 0%
Iron Cap | Gold
>2 g/t Au | = 8%
0.05 - 2 g/t Au | = 10%
<0.05 g/t Au | = 5%
Silver |
>20 g/t Ag | = 8%
10 – 20 g/t Ag | = 11%
0.5 – 10 g/t Ag | = 16%
<0.5 g/t Ag | = 5%

Note: the doré recoveries are in addition to flotation gold and silver recoveries.

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Table 13.50 Mo Flotation Concentrate Metal Recovery and Grade (Molybdenum Concentrate Grade = 50%)

Mo Head Grade (%) | Mo Recovery (%)
>0.010 | 47
0.005-0.010 | 35
0.0025-0.005 | 25
<0.0025 | 0

| 13.7.2 | Metallurgical Performance Projection – Deep Kerr and Lower Iron Cap

After 2016, further metallurgical test work has been conducted, mainly on the mineral samples generated from the Deep Kerr and Iron Cap zones. According to the data obtained from the test programs, the copper, gold, and silver recovery performance projections for the lower zones of the Kerr and Iron Cap deposits have been updated and are summarized in Table 13.51, Table 13.52 and Table 13.53.

The similar copper grade for the copper-gold concentrate is assumed for the lower zone mineralization, compared to the other mineralization shown in Table 13.47.

The average primary grind size used for the metallurgical performances for both the lower Iron Cap and the Deep Kerr lower mineralization is 80% passing approximately 125 µm. For the lower Iron Cap zone mineralization, a diluted slurry solid density of approximately 25% is also considered in the metallurgical performance projections.

Table 13.51 Cu-Au Flotation Concentrate –Metal Recovery Projections

Deposit | Description | Head Grade | Recovery
Deep Kerr | Copper Recovery | > 1.5% Cu | = 96.5
0.455 – 1.5% Cu | = 5.5138 x ln (Cu Head, %) + 93.548
0.20 – 0.455% Cu | = 95.434 x (Cu Head, %) 0.0856
0.10 – 0.20% Cu | = 78
< 0.10% Cu | = 30
Gold Recovery | n/a | = 0.9812 x (Copper Recovery, %) – 26.327
Silver Recovery | n/a | = 0.9773 x (Copper Recovery, %) – 33.149; If silver recovery < 0, use 0
Lower Iron Cap | Copper Recovery | > 1.0% Cu | = 95
0.5 – 1.0% Cu | = 93
0.2 – 0.5% Cu | = 97.915 x (Cu Head, %) 0.0849
0.1 - 0.2% Cu | = 106.39 x (Cu Head, %) 0.1371
0.05 - 0.1% Cu | = 30
< 0.05% Cu | = 3
Gold Recovery | > 2.0 g/t Au | = 73
0.8 – 2.0 g/t Au | = 70
0.15 – 0.8 g/t Au | = 64.543 x (Au Head, g/t) 0.021
< 0.15 g/t Au | = 40
Silver Recovery | >20 g/t Ag | = 80
11 – 20 g/t Ag | = 78
0.5 – 11 g/t Ag | = 46.094 x (Ag Head, %)0.206
<0.5 g/t Ag | = 20

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Table 13.52 Au-Ag Doré – Cyanide Leach Metal Recovery Projections

Deposit | Head Grade | Recovery
Deep Kerr | Gold
> 1.0 | = 12
0.1 – 1.0 | = 14
< 0.1 | = 0
Silver
> 4.5 | = 11
1 – 4.5 | = 16
< 1.0 | = 5
Lower Iron Cap | Gold
>0.8 g/t Au | = 10%
0.15 – 0.8 g/t Au | = 11%
<0.15 g/t Au | = 7%
Silver
>11 g/t Ag | = 8%
0.5 – 11 g/t Ag | = 14%
<0.5 g/t Ag | = 5%

| Notes: | 1) The doré recoveries are in addition to flotation gold and silver recoveries.

2) Gold and silver recoveries to doré from the flotation products of the first cleaner flotation tailings and the pyrite concentrate are projected based on test work in which the cyanide leaching conditions were not optimized. Whether it is economic to run the cyanide leach circuit on the gold-bearing flotation tailings to recover the additional gold and silver depends on leach circuit feed grades, metal prices, and operating costs. Further test work is recommended to improve metallurgical performances and reduce reagent consumption.

The molybdenum metallurgical performance projection is same as shown in Table 13.53 for the lower Iron Cap mineralization; the molybdenum metallurgical performance for the Deep Kerr mineralization is summarized in Table 13.53.

Table 13.53 Mo Flotation Concentrate Metal Recovery and Grade (Molybdenum Concentrate Grade = 50%)

Mo Head (%) | Mo Recovery (%)
> 0.010; Mo/Cu > 0.006 | = 45
0.0050-0.010; Mo/Cu > 0.006 | = 35
0.0025-0.0050; Mo/Cu > 0.006 | = 25
<0.0025 | = 0

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| 14.0 | Mineral Resource Estimates

Mineral Resources were estimated for the KSM Property by Mr. Michael J. Lechner, President of Resource Modeling Inc. (RMI). Mr. Lechner is a P.Geo. (Province of British Columbia), a Registered Professional Geologist in the State of Arizona, a Certified Professional Geologist with the American Institute of Professional Geologists (AIPG), and a registered member of the Society for Mining, Metallurgy, and Exploration (SME). These professional registrations together with Mr. Lechner’s education, professional background and work experience allow him to be the QP for the sections of this report that he is responsible for preparing, as per the requirements set out by NI 43-101. Neither Mr. Lechner nor RMI have any vested interest in Seabridge securities or the Property that is the subject of this Technical Report. Mr. Lechner and RMI have worked as an independent consultant for Seabridge since 2001.

The Qualified Person responsible for Mineral Resources has prepared a number of previous filed NI 43-101 Technical Reports regarding Mineral Resource estimates of the various deposits within the KSM property (Lechner, 2007, Lechner, 2008a, Lechner, 2008b, Lechner, 2009, Lechner, 2010, Lechner, 2011, and Lechner, 2014). In addition, Mineral Resource estimates were prepared by the Qualified Person responsible for resources in the 2012 and 2016 PFS Technical Reports (Tetra Tech, 2012, and Tetra Tech, 2016).

Previously, the four mineralized zones were modeled within a single block model using 25m x 25m x 15m blocks. As more understanding was gained after each annual drilling campaign, individual block models were created for each area. Grade interpolation parameters have also evolved over time, reflecting changes required for modeling deeper mineralization intersected below the Kerr and Iron Cap deposits. Table 14.1 compares and contrasts some of the basic modeling parameters associated with the four KSM resource areas.

Table 14.1 Summary of KSM Block Model Parameters by Deposit

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A variety of basic descriptive statistics and spatial analyses were completed for each area upon the completion of annual drilling campaigns. These investigations include the generation of grade distribution tables, grade histograms, cumulative probability plots, grade box plots, grade contact plots, down-hole variograms, and directional variograms. In addition, new drill hole results were typically compared against the previous grade model to assess model performance.

The following sections summarize the key assumptions, parameters and methods that were used to estimate resources for each of the four mineralized areas for which Mineral Resources have been established.

| 14.1 | Kerr Deposit

The Kerr resource estimate was last updated using drilling data that had been collected through 2016. The block model was informed by 223 diamond core holes totaling about 84,770 m of assayed drilling data. Approximately 75% of that drill data was collected by Seabridge. Placer Dome drilled over 80 relatively shallow holes in 1992 while exploring for an open pit copper-gold deposit.

| 14.1.1 | Grade Distribution – Kerr Deposit

The distribution of gold and copper grades within the Kerr deposit were compared with logged lithology and alteration to determine if those attributes could be used for grade estimation purposes. Grade was often seen to cross-cut various lithologic and/or alteration boundaries. For that reason, grade wireframes were designed to constrain the estimate of block grades. Five gold grade wireframes and six copper wireframes were designed by Seabridge’s geologic staff.

The distribution of gold based on uncapped, uncomposited assay data is summarized at four different cutoff grades by five gold grade wireframes. The data in Table 14.2 shows an average gold grade of 0.21 g/t for all data and a coefficient of variance of 3.1. Table 14.3 summarizes uncapped copper assay statistics for six copper grade wireframes. The average copper grade for all data is seen to be 0.24% with a coefficient of variance of 1.4.

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Table 14.2 Distribution of Gold by AUZON – Kerr Deposit

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Table 14.3 Distribution of Copper by CUZON – Kerr Deposit

| 14.1.2 | Assay Grade Capping – Kerr Deposit

High-grade outlier values were identified using cumulative probability plots for gold, copper, silver, and molybdenum assays. This analysis was completed on the raw assay data prior to compositing for each metal by grade domains. Figure 14.1 shows an example of gold (upper) and copper (lower) cumulative probability plots. Assays from higher grade gold (+ 0.20 g/t) and copper (+0.2%) were used to generate the two probability plots. The circles indicate interpreted capping levels for each metal.

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Figure 14.1 Gold-Copper Probability Plots – Kerr Deposit

Source: (RMI, 2019)

Based on the interpretation of the cumulative probability plots, assays were “cut” or capped at specified thresholds for each metal by grade wireframe domain. 53 gold assays were capped using 1.75 g/t for the lower grade wireframes and 10.0 g/t for the higher grade wireframes. 23 copper assays were capped at 2% for the lower grade domains and 4% for higher grade domains.

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| 14.1.3 | Drill Hole Compositing – Kerr Deposit

The capped drill hole assays were composited into 15 m long composites starting from the drill hole collar keying off of the grade wireframes to start/stop the compositing at wireframe contacts. Two sets of composites were generated, one for precious metals and another for base metals. The precious metal composites keyed off of the gold grade wireframes and the base metal composites keyed off of the copper grade wireframes. Most of the original assay data were in the range of 1.5 to 3.0 m long, with the majority being 2 m long. Based on the scale of the deposit, 15 m long composites were deemed to be an appropriate length for estimating Mineral Resources.

The assays were composited using MineSight® software. Various geologic data were assigned to the 15 m long composites using the majority rule method.

| 14.1.4 | Geologic Constraints - Kerr Deposit

Lithologic, alteration, structural domains, and metal grade envelopes were constructed for the Kerr deposit by Seabridge and the Qualified Person responsible for this section of this Technical Report. Initially these various wireframes were interpreted onto cross sections, which were then reconciled in bench plan prior to building the final wireframe. Leapfrog software has been used in recent years as a method for generating wireframes for various attributes. The Leapfrog wireframes were typically modified to account for structural and lithologic criteria.

| 14.1.5 | Variography – Kerr Deposit

Spatial continuity was investigated by generating both grade and indicator correlograms using Sage2001 software. Down-hole correlograms were used to establish nugget effects using 2 m long drill hole composites. Directional correlograms were generated using 15 m long drill hole composites in 30-degree vectors in plan and section. The autofit function in Sage2001 was used to model the 37 directional correlograms.

Figure 14.2 shows a gold (upper) and copper (lower) down-hole correlogram. Figure 14.3 shows three pertinent directional correlograms for gold (left panel) and copper (right panel) based on correlograms generated from 15 m composites.

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Figure 14.2 Down-hole Grade Correlograms – Kerr Deposit

Source: (RMI, 2019)

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Figure 14.3 Directional Grade Correlograms – Kerr Deposit

Source: (RMI, 2019)

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| 14.1.6 | Grade Estimation Parameters – Kerr Deposit

The Kerr deposit was modeled using MineSight®, a widely recognized commercial mine engineering software package. The non-rotated model uses NAD83 UTM coordinates with 15 m x 15 m x 15m blocks. Block gold, silver, copper, and molybdenum grades were estimated by two methods: inverse distance weighting (IDW), and nearest neighbour (NN). Trend planes were used to select eligible drill hole composites for the inverse distance grade models in lieu of traditional search ellipses. Four structural domain wireframes were developed to reflect different mineralized strike/dip regimes which allowed for specific trends to be applied tor each area. Two-pass strategy was used for estimating block gold, silver, copper, and molybdenum grades. A minimum of three composites, a maximum of six composites with no more than two composites per drill hole were used in the grade estimate process. The first estimation pass used a very narrow search in the direction perpendicular to the plunge of mineralization (400m x 400m x 20m). All blocks estimated by the first pass were flagged and not overwritten by the second estimation pass which used a 400m x 400m x 80m search relative to the specified trend plane. Grade envelopes were used as the primary constraint for selecting eligible drill hole composites. Gold wireframes were used in the estimate of gold and silver. Copper wireframes were used in the estimate of copper and molybdenum. The gold grade wireframes were not used as hard boundaries in the estimation process. Drill hole composites from an adjacent lower grade domain were potentially allowed to inform model blocks, provided the composites were located within the trend plane search parameters. For example, a model block within the 0.20 g/t domain could use composites from either the 0.20 g/t domain or the adjacent 0.10 g/t. The highest-grade copper domain (+ 1%) was treated as a hard contact in the estimation process based on visual observations of drill core.

| 14.1.7 | Grade Model Verification - Kerr Deposit

Estimated block grades were verified by visual and statistical methods. Block grades (gold, silver, copper, and molybdenum) were visually compared with drill hole composite grades in cross section and level plan views. Figure 14.4 and Figure 14.5 are east-west cross sections through the Kerr block model showing gold and copper grades, respectively. For reference, the location of the Kerr cross sections is illustrated in Figure 10.2, a drill hole plan map for the Kerr deposit. Figure 14.6 and Figure 14.7 are block model level maps drawn at the 850 m elevation through the Kerr block model showing estimated block/composite gold and copper grades, respectively.

NN models were prepared for gold, copper, silver, and molybdenum in order to check for potential global biases in the estimated block grades. Swath plots were generated by block column (easting), block row (northing), and block level (elevation) comparing the NN grade with the IDW grade. Figure 14.8 shows two grade swath plots for gold (upper) and copper (lower) that compare the IDW grades in red versus the NN grades in blue by elevation. The results show that the inverse models compare very well with the NN grades on a local basis.

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Figure 14.4 Kerr Cross Section 6,259,650 N. – Gold

Source: (RMI, 2019)

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Figure 14.5 Kerr Cross Section 6,259,650 N. – Copper

Source: (RMI, 2019)

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Figure 14.6 Kerr 850 m Level Plan – Gold

Source: (RMI, 2019)

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Figure 14.7 Kerr 850 m Level Plan - Copper

Source: (RMI, 2019)

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Figure 14.8 Kerr Gold-Copper Swath Plots by Elevation

Source: (RMI, 2019)

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| 14.1.8 | Resource Classification – Kerr Deposit

Mineral Resources were defined by drill hole spacing and mineralized continuity. Both Indicated and Inferred wireframes were constructed in cross section and level plan keying on distribution of drilling, mineralized trends, and post mineral structures.

Indicated Mineral Resources were restricted to the upper portion of the deposit where the density of drilling is much higher than the deeper mineralization. The Inferred Mineral Resource shape was generated to envelope the Indicated shape and to follow the two steep west dipping mineralized limbs of the deposit. In general, the drill hole spacing for Indicated Resources is about 50 m while the spacing for Inferred Resources varies between 125 m and 150 m, with wider spacing with depth.

| 14.2 | Sulphurets Deposit

The Sulphurets geologic and grade models were updated in mid 2019 using data through the 2018 drilling campaign. Eleven new drill holes were completed within the Sulphurets mineralized zone since the last model update (EOY2011). After the Sulphurets model was updated, additional drill holes were completed in late 2019 to the west and east of the currently recognized mineralized system. The 2019 drill holes were compared against the EOY2018 model and were found to confirm the geologic interpretations and were consistent with the grade model in the area of the 2019 drilling.

| 14.2.1 | Grade Distribution – Sulphurets Deposit

The distribution of gold and copper grades within the Sulphurets deposit were compared with logged lithology and alteration to determine if those attributes could be used for grade estimation purposes. Grade was often seen to cross-cut various lithologic and/or alteration boundaries or cut off by structures that form discrete blocks or panels of mineralization. Four gold and copper grade wireframes were designed by Seabridge’s geologic staff by structural block. In addition to the grade wireframes, three post-mineral dikes and a distinct monzonite intrusive unit were combined with the grade domains as an aid in constraining the estimate of block grades.

The distribution of gold based on uncapped, uncomposited assay data is summarized at four different cutoff grades by gold grade and lithologic wireframes. The data in Table 14.4 shows an average uncapped gold grade of 0.44 g/t for all data. The high coefficient of variation shown for all data and for the 0.125 g/t gold envelope were highly skewed by a single gold assay grade of 1,580 g/t from drill hole S-18-81. This highly anomalous sample was capped at 2.0 g/t which lowered the coefficient of variation from 23.4 to 1.6. Table 14.5 summarizes uncapped copper assay statistics by copper grade domain. The average copper grade for all data is seen to be 0.12% with a coefficient of variation of 1.49. The two higher grade domains (0.20% and 0.50%) show coefficients of variation that are less than 1.0, reflecting the role of grade domaining in reducing variability.

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Table 14.4 Distribution of Gold by AUZON – Sulphurets Deposit

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Table 14.5 Distribution of Copper by CUZON – Sulphurets Deposit

| 14.2.2 | Assay Grade Capping – Sulphurets Deposit

High-grade outlier values were identified using cumulative probability plots for gold, copper, silver, and molybdenum assays. This analysis was completed on the raw assay data prior to compositing for each metal by grade domains. Figure 14.9 shows examples of gold (upper) and copper (lower) cumulative probability plots for all Sulphurets assays.

Based on the interpretation of the cumulative probability plots, assays were “cut” or capped at specified thresholds for each metal by grade wireframe domain. A total of 112 gold assay samples were capped. The lower grade gold domains, post mineral dikes, and monzonite were capped at 1 g/t. The upper grade gold domains (0.25 to 0.50 and + 0.50 g/t domains) were capped at 5.0 and 10.0 g/t, respectively. 137 copper assays were capped at 0.60% and 1.00% for the lower grade domains and 2% for high-grade domain (e.g. + 0.5%). The gold and copper limits applied to the end-of-year 2018 Sulphurets model were similar to the previous end-of-year 2011 model.

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Figure 14.9 Gold-Copper Probability Plots – Sulphurets Deposit

Source: (RMI, 2019)

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| 14.2.3 | Drill Hole Compositing – Sulphurets Deposit

The capped drill hole assays were composited into 10 m long composites starting from the drill hole collar keying off of the grade wireframes to start/stop the compositing at wireframe contacts using MineSight® software. Various geologic data were assigned to the drill hole composites using the majority rule method. Four sets of composites were generated, one for each of the estimated metals (i.e. gold, silver, copper, and molybdenum). Most of the original assay data were in the range of 1.5 to 3.0 m long, with the majority being 2 m long. Based on the scale of the deposit and model block dimensions, 10 m long composites were deemed to be an appropriate length for estimating Mineral Resources.

| 14.2.4 | Geologic Constraints – Sulphurets Deposit

Lithologic, alteration, structural domains, and metal grade envelopes were constructed for the Sulphurets deposit by Seabridge and the Qualified Person responsible for this section of this Technical Report. Leapfrog software was used to generate preliminary wireframes that were further modified to account for various structural and lithologic criteria.

After reviewing gold and copper distribution by lithology and alteration it was determined that independently constructed grade wireframes would be more appropriate for estimating block grades. Five structural domain wireframes were also used in the interpolation plan as hard boundaries along post mineral structures that locally juxtapose well mineralized material against poorly mineralized rock.

| 14.2.5 | Variography – Sulphurets Deposit

Spatial continuity was investigated by generating both grade and indicator correlograms using Sage2001 software. Down-hole correlograms were used to establish nugget effects using 2 m long drill hole composites. Directional correlograms were generated using 10 m long drill hole composites in 30-degree vectors in plan and section. The autofit function in Sage2001 was used to model the 37 directional correlograms.

Figure 14.10 shows a gold (upper) and copper (lower) down-hole correlogram. Figure 14.11 shows three pertinent directional correlograms for gold (left panel) and copper (right panel) based on Sage2001 correlograms.

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Figure 14.10 Down-hole Grade Correlograms – Sulphurets Deposit

Source: (RMI, 2019)

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Figure 14.11 Directional Grade Correlograms – Sulphurets Deposit

Source: (RMI, 2019)

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| 14.2.6 | Grade Estimation Parameters – Sulphurets Deposit

The Sulphurets deposit was modeled using MineSight®, a widely recognized commercial mine engineering software package. The non-rotated model uses NAD83 UTM coordinates with 12.5 m x 12.5 m x 15m blocks.

Block gold, copper, silver and molybdenum grades were estimated by three primary methods: IDW, OK, NN. The OK model was selected for resource declaration after comparing that model with the IDW model.

Independently constructed gold, copper, silver, and molybdenum wireframes provided the principal constraint in the grade estimation plan. In addition to the grade envelopes, five structural domain wireframes were developed to reflect different mineralized strike/dip regimes which allowed for specific search ellipses to be incorporated for each structural block. The gradeshells were treated as soft contacts in the grade estimation plan where lower grade samples were allowed to inform the next highest grade shell but the higher grade samples were not allowed to inform adjacent lower grade zones.

A two-pass strategy was used for estimating block gold, silver, copper, and molybdenum grades for both the IDW and OK models. The first estimation pass used a 275 m x 275 m x 55 m search ellipse that required a minimum of five drill hole composites, a maximum of nine composites with no more than three composites per drill hole. That criteria resulted in blocks being estimated at least two drill holes. All blocks estimated by the first pass were flagged and not overwritten by the second estimation pass which used a 125 m x 125 m x 25 m search that required a minimum of two drill hole composites, a maximum of nine composites with no more than three composites per drill hole. That criteria resulted in blocks that could have been estimated by one or more drill holes.

Narrow, post-mineral dike wireframes were developed for the end-of-year 2018 Sulphurets geologic model. The volume percentage of these narrow dikes was stored in the block model. Two grades were estimated for blocks that contained less than 50% dike material. One grade was based on dike-only samples, and the other block grade was estimated with samples that matched the grade shell code of the remaining majority percentage of the block. A final weighted block grade was calculated using the two block proportions and their associated grades. Blocks that contained more than 50% dike were considered to be 100% dike material.

| 14.2.7 | Model Validation – Sulphurets Deposit

Estimated block grades (gold, silver, copper, and molybdenum) were compared to drill hole composite grades in cross section and level plan views. Figure 14.12 and Figure 14.13 are northwest-southeast cross sections drawn through the Sulphurets block model (refer to Figure 10.3 for the location of Section 20). Figure 14.14 and Figure 14.15 are block model level maps drawn at the 1135 m elevation through the Sulphurets block model showing estimated block grades and drill hole composite for gold and copper, respectively.

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Figure 14.12 Sulphurets Cross Section 20 - Gold

Source: (RMI, 2019)

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Figure 14.13 Sulphurets Cross Section 20 – Copper

Source: (RMI, 2019)

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Figure 14.14 Sulphurets 1135m Level Plan - Gold

Source: (RMI, 2019)

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Figure 14.15 Sulpurets 1135 m Level Plan – Copper

Source: (RMI, 2019)

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The grade models were also validated by comparing the OK block grades against NN models that were generated for gold, copper, silver, and molybdenum.

Grade swath plots were generated for rows (east-west), columns (north-south) and levels (elevations) through the block model comparing the OK and NN models at a zero cutoff grade. Figure 14.16 shows swath plots for gold (upper) and copper (lower) by elevation. These swath plots show Indicated and Inferred Resource grades.

Figure 14.16 Sulphurets Gold-Copper Swath Plots by Elevation

Source: (RMI, 2019)

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| 14.2.8 | Resource Classification – Sulphurets Deposit

Mineral Resources were assigned to the estimated blocks by constructing 3D solids for Indicated and Inferred Resources. These shapes were based on mineralized continuity as defined by exploration drill hole results. The Indicated Mineral Resource shape was locally extended 25 m to 100 m deeper than the shape used for the 2016 PFS Mineral Resource based on drilling results obtained in 2013 through 2018 with an average drill hole spacing of approximately 60 m to 75 m. The average drill hole spacing for Indicated and Inferred blocks is approximately 50 m to 75 m and 75 m to 125 m, respectively.

| 14.3 | Mitchell Deposit

The basis for Mitchell Mineral Resources is the block model that was used in the 2016 PFS (Tetra Tech, 2016). That model was constructed using drilling data collected through 2011. Since that model was completed, 23 holes were drilled within the Mitchell resource area, although only 8 holes intersected estimated Mineral Resources based on the end-of-year 2011 block model. Grade comparisons between the 8 new holes and the Mitchell block model were presented in Section 12.2.

| 14.3.1 | Metal Distribution – Mitchell Deposit

Basic descriptive statistics were generated by various logged attributes like lithology and alteration. The interpretation of those results did not identify any single or combined attributes that adequately defined the principal controls of mineralization. Independently generated gold, copper, and molybdenum wireframes were developed by Seabridge’s geologic staff. Table 14.6 summarizes uncapped gold assay distributions at four cutoff grades for seven of the gold wireframe domains. Included with the gold grade shell domains are a structurally controlled bornite copper zone and an associated leach breccia zone. Table 14.7 summarizes uncapped copper assay distributions at four cutoff grades plus the two aforementioned copper zones (bornite and leach).

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Table 14.6 Distribution of Gold by AUZON – Mitchell Deposit

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Table 14.7 Distribution of Copper by CUZON – Mitchell Deposit

| 14.3.2 | Assay Grade Capping – Mitchell Deposit

Cumulative probability plots were used to identify high-grade outliers for gold, copper, silver, and molybdenum based on the original assay samples. The assays were initially examined with respect to logged lithology and alteration types however the final analysis of high-grade outliers was completed using the grade wireframes that were used to interpolate block grades. Figure 14.17 shows gold (upper) and copper (lower) cumulative probability plots that define the distribution of assays.

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Figure 14.17 Gold-Copper Cumulative Probability Plots – Mitchell Deposit

Source: (RMI, 2019)

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| 14.3.3 | Drill Hole Compositing – Mitchell Deposit

Drill hole assay data (both uncapped and capped intervals) were composited into 15 m long composites starting from the drill hole collar. Most of the original assay data were in the range of 1.5 m to 3.0 m long, with the majority being 2 m long. Based on the scale of the deposit, 15 m long composites were deemed to be an appropriate length for estimating Mineral Resources. Prior to creating the drill hole composites, the drill hole intervals were coded with the same grade wireframes that were used to constrain the estimate of block grades.

After compositing, block model attributes like lithology, alteration, fault block, and grade envelopes were backtagged from the model to the drill hole composites.

| 14.3.4 | Variography – Mitchell Deposit

A variety of grade and indicator variograms were generated for the Mitchell deposit using MineSight® and Sage2001 software. Figure 14.18 shows gold (upper) and copper (lower) down-hole correlograms based on 2 m long drill hole composites.

Figure 14.19 shows key directional correlograms for gold (left panel) and copper (right panel) based on 15 m drill hole composites. Nested spherical models were used in modeling the Mitchell gold and copper correlograms. The gold and particularly the copper directional correlograms show relatively long ranges.

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Figure 14.18 Down-hole Grade Correlograms – Mitchell Deposit

Source: (RMI, 2019)

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Figure 14.19 Directional Grade Correlograms – Mitchell Deposit

Source: (RMI, 2019)

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| 14.3.5 | Grade Estimation Parameters – Mitchell Deposit

Gold, copper, silver, and molybdenum block grades were estimated for the Mitchell deposit using a multi-pass inverse distance squared (IDW) method. Grades were estimated using 15 m long drill hole composites. The gold grade wireframes were used as the primary constraint in estimating gold and silver grades. Independently developed copper and molybdenum wireframes provided the primary constraint for copper and molybdenum estimation, respectively. In addition to the grade shell domains, two structural domains were defined by the Mitchell thrust fault which was treated as a hard contact in the estimation plan.

For the key gold grade wireframe zones, both gold and silver grades were estimated using four passes featuring increasingly larger search ellipses. Once estimated blocks were flagged as final and were not overwritten by subsequent estimation passes. For the first three passes, a minimum of three, a maximum of eight composites were required with no more than two composites per drill hole. This requirement resulted in blocks that were estimated by at least two different drill holes. The first pass used an ellipse measuring 125 m x 125 m x 30 m. The second pass was doubled to 250 m x 250 m x 60 m. The third pass ellipse dimension was tripled to 375 m x 375 m x 90 m. The fourth pass was treated as a cleanup run to provide a local estimate around isolated drill holes.

The number of composites, drill holes, average distance of drill holes used, and distance to the closest drill hole used were stored during the estimation process. These attributes were used in conjunction with wireframes in defining Mineral Resource categories.

Similar to gold, copper grades were estimated by a multi-pass inverse distance squared method. Copper grade envelopes and the two aforementioned structural domains were used as the primary constraint in the estimation plan. A three-pass strategy was used where increasingly larger search ellipses were used. The first pass ellipse measured 150 m x 150 m x 30 m and required a minimum of three, a maximum of eight composites and no more than two composites from each drill hole. The second estimation pass ellipse dimensions were doubled (i.e. 250 m x 250 m x 60 m) with the same drill hole composite criteria. The final estimation pass was extended to 500 m x 500 m x 120 m. The last copper estimation pass was used to provide an estimate for blocks that were not informed by at least two drill holes. Like gold, the number of copper composites, drill holes and distances to data were stored in the model.

Molybdenum block grades were estimated using a two-pass inverse distance squared estimation method. The first pass used a spherical search ellipse measuring 300 m and only required a minimum of one drill hole. The second pass ellipse was allowed to overwrite results from the first pass provided at least two drill holes were found within a 250 m x 250 m x 50 m ellipse that required a minimum of three, a maximum of eight composites with no more than two per drill hole.

Nearest neighbor grade models were generated simultaneously with the inverse distance models using the same constraint requirements (i.e. grade shell and structural block criteria). Results from the nearest neighbor models were used in verifying that the grade model was globally unbiased.

| 14.3.6 | Grade Model Validation – Mitchell Deposit

Estimated block grades (gold, silver, copper, and molybdenum) were compared to drill hole composite grades in cross section and level plan views. Figure 14.20 and Figure 14.21 are north-south cross sections drawn through the Mitchell block model. The location of Mitchell Section 11 shown in Figures 14.20 and 14.21 is illustrated in Figure 10.4. Figure 14.22 and Figure 14.23 are block model level maps drawn at the 760 m elevation through the Mitchell block model showing estimated block grades and drill hole composite grades for gold and copper, respectively. The constraining Mineral Resource pit and block cave shape are shown on the block model cross sections.

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Figure 14.20 Mitchell Cross Section 11 - Gold

Source: (RMI, 2019)

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Figure 14.21 Mitchell Cross Section 11 - Copper

Source: (RMI, 2019)

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Figure 14.22 Mitchell760 m Level Plan - Gold

Source: (RMI, 2019)

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Figure 14.23 Mitchell 760 m Level Plan - Copper

Source: (RMI, 2019)

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The grade models were also validated by comparing the IDW block grades against NN models that were generated for gold, copper, silver, and molybdenum.

Grade swath plots were generated for rows (east-west), columns (north-south) and levels (elevations) through the block model comparing the IDW and NN models at a zero cutoff grade. Figure 14.24 shows grade swath plots by elevation for gold (upper panel) and copper (lower panel). These swath plots show results from all resource blocks.

Figure 14.24 Mitchell Swath Plots by Elevation

Source: (RMI, 2019)

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| 14.3.7 | Resource Classification – Mitchell Deposit

Mineral Resources were assigned to the estimated blocks through a combination of coding blocks with a wireframe representing Indicated Resources and the number of drill holes and distance to drilling. The Indicated Resource wireframe was constructed by generating a gold equivalent probability model. Blocks with an estimated probability in excess of 50% for a 0.50 g/t gold equivalent cutoff grade were used as a guideline in drawing shapes in cross section that were reconciled in plan view for creation of the final wireframe solid. Blocks inside of the Indicated shape were then coded as Measured Resources if the blocks were estimated by two or more holes with one of the holes located within 50 m of the block or one drill hole within 17 m of the block. Blocks not coded as Measured or Indicated were then flagged as Inferred Resources based on distance to drilling and the number of holes used. Blocks were flagged as Inferred if they were estimated by one hole within 75 m or two holes within 175 m of the block. In general, the drill hole spacing for Measured and Indicated Resources is about 75 to 100 meters while the spacing for Inferred Resources varies between 100 and 150 meters, with wider spacing with depth.

| 14.4 | Iron Cap Deposit

The Iron Cap grade model was updated following the completion of the 2018 drilling campaign. A total of 99 diamond core holes totaling about 66,740 m were used in the update. Previous Iron Cap models used trend plane searches and inverse distance estimation methods. The end-of-year 2018 model which is the basis for Mineral Resources that are the subject of this Technical Report is based on OK methods that used conventional search ellipses.

| 14.4.1 | Grade Distribution – Iron Cap Deposit

Basic descriptive statistics were generated for various logged attributes like lithology and alteration. The interpretation of those results did not identify any single or combined attributes that adequately defined the principal controls of mineralization. Independently generated gold, copper, and molybdenum wireframes were developed by Seabridge’s geologic staff. Table 14.8 summarizes uncapped gold assay distributions at four cutoff grades for five gold wireframe domains. Table 14.9 summarizes uncapped copper assay distributions at five cutoff grades.

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Table 14.8 Distribution of Gold by AUZON – Iron Cap Deposit

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Table 14.9 Distribution of Copper by CUZON – Iron Cap Deposit

| 14.4.2 | Assay Grade Capping – Iron Cap Deposit

Cumulative probability plots were used to identify high-grade outliers for gold, copper, silver, and molybdenum based on the original assay samples. The assays were initially examined with respect to logged lithology and alteration types. However, the final analysis of high-grade outliers was completed using the grade wireframes that were used to interpolate block grades. Figure 14.25 shows gold (upper) and copper (lower) cumulative probability plots that define the distribution of assays.

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Figure 14.25 Gold-Copper Cumulative Probability Plots – Iron Cap Deposit

Source: (RMI, 2019)

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| 14.4.3 | Drill Hole Compositing – Iron Cap Deposit

The capped drill hole assays were composited into 10 m long composites starting from the drill hole collar keying off of the grade wireframes to start/stop the compositing at wireframe contacts using MineSight® software. Various geologic data were assigned to the drill hole composites using the majority rule method. Two sets of composites were generated, one for precious metals (gold and silver) and one for base metals (copper and molybdenum). Most of the original assay data were in the range of 1.5 to 3.0 m long, with the majority being 2 m long. Based on the scale of the deposit and model block dimensions, 10 m long composites were deemed to be an appropriate length for estimating Mineral Resources.

| 14.4.4 | Geologic Constraints – Iron Cap Deposit

Lithologic, alteration, structural domains, and metal grade envelopes were constructed for the Sulphurets deposit by Seabridge and the Qualified Person responsible for this section of this Technical Report. Leapfrog software was used to generate preliminary wireframes that were further modified to account for various structural and lithologic criteria.

After reviewing gold and copper distribution by lithology and alteration it was determined that independently constructed grade wireframes would be more appropriate for estimating block grades. Three structural domains were also used in the interpolation plan as hard boundaries.

| 14.4.5 | Variography – Iron Cap Deposit

Spatial continuity was investigated by generating both grade and indicator correlograms using Sage2001 software. Down-hole correlograms were used to establish nugget effects using 2 m long drill hole composites. Directional correlograms were generated using 10 m long drill hole composites in 30-degree vectors in plan and section. The autofit function in Sage2001 was used to model the 37 directional correlograms.

Figure 14.26 shows gold (upper) and copper (lower) down-hole correlograms. Figure 14.27 shows three pertinent directional correlograms for gold (left panel) and copper (right panel) based on Sage2001 correlogram.

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Figure 14.26 Down-hole Grade Correlograms – Iron Cap Deposit

Source: (RMI, 2019)

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Figure 14.27 Directional Grade Correlograms – Iron Cap Deposit

Source: (RMI, 2019)

| 14.4.6 | Grade Estimation Parameters – Iron Cap Deposit

The Iron Cap deposit was modeled using MineSight® software, a widely recognized commercial mine engineering software package. The non-rotated model uses NAD83 UTM coordinates with 15 m x 15 m x 15 m blocks.

Block gold, copper, silver and molybdenum grades were estimated by three primary methods: IDW, OK, and NN. The OK model was selected for resource declaration after comparing that model with the IDW model and NN models.

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Independently constructed gold, copper, silver, and molybdenum wireframes provided the principal constraint in the grade estimation plan. In addition to the grade envelopes, three structural domain wireframes were developed to reflect different mineralized strike/dip regimes which allowed for specific search ellipses to be incorporated tor each structural block. The gradeshells were treated as soft contacts in the grade estimation plan where lower grade samples were allowed to inform the next highest grade shell but the higher grade samples were not allowed to inform adjacent lower grade zones.

A four-pass strategy was used for estimating block gold, silver, copper, and molybdenum grades for both the IDW and OK models. The first estimation pass used a 175 m x 175 m x 35 m search ellipse that required a minimum of five drill hole composites, a maximum of nine composites with no more than three composites per drill hole. That criteria resulted in blocks being estimated with at least two drill holes. All blocks estimated by the first pass were flagged and not overwritten by the subsequent estimation passes. The second pass used a 350 m x 350 m x 70 m search that required a minimum of five drill hole composites, a maximum of nine composites with no more than three composites per drill hole. The third estimation pass used a 500 m x 500 m x 100 m search ellipse and the same number of composites as pass one and two. A final cleanup pass used the same search ellipse as the third pass but only required a minimum of one drill hole.

NN models were generated for each metal at the same time the ordinary kriged grade was estimated. The NN models were used to access potential global biases in the grade estimates.

The number of composites and drill holes used to estimate each block along with distances to data (average and closest) were stored and used to help in the determination of Mineral Resource categories.

| 14.4.7 | Grade Model Verification – Iron Cap Deposit

Estimated block grades (gold, silver, copper, and molybdenum) were compared to drill hole composite grades in cross section and level plan views. Figure 14.28 and Figure 14.29 are northwest-southeast cross sections drawn through the Iron Cap block model (refer to Figure 10.5 for the location of cross section 12). Figure 14.30 and Figure 14.31 are block model level maps drawn at the 1200 m elevation through the Iron Cap block model showing estimated block grades and drill hole composite for gold and copper, respectively.

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Figure 14.28 Iron Cap Cross Section 12 – Gold

Source: (RMI, 2019)

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Figure 14.29 Iron Cap Cross Section 12 – Copper

Source: (RMI, 2019)

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Figure 14.30 Iron Cap 1200 m Level Plan – Gold

Source: (RMI, 2019)

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Figure 14.31 Iron Cap 1200 m Level Plan – Copper

Source: (RMI, 2019)

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Grade swath plots were generated for rows (east-west), columns (north-south) and levels (elevations) through the block model comparing the OK and NN models at a zero cutoff grade. Figure 14.32 shows swath plots for gold (upper) and copper (lower) by elevation. These swath plots show Indicated and Inferred Resource grades.

Figure 14.32 Iron Cap Gold-Copper Swath Plots by Elevation

Source: (RMI, 2019)

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| 14.4.8 | Resource Classification – Iron Cap Deposit

Mineral Resources were assigned to the estimated blocks by constructing 3D solids for Indicated and Inferred Resources. These shapes were based on mineralized continuity as defined by exploration drill hole results. The average drill hole spacing for Indicated and Inferred blocks is approximately 75 m and 125 to 150 m, respectively. The drill hole spacing is wider at depth for Inferred Resources.

| 14.5 | Bulk Density

Bulk density values were assigned to each deposit based on an analysis of available specific gravity analyses that were performed on drill core during the core logging process. Bulk density determinations were routinely performed on core samples at a frequency of one sample per 100 m of drilling. While rare, apparent anomalously high or low determinations were excluded for analysis. Most of bulk density data were assigned to each deposit by average densities by modeled lithology or alteration. Overburden and ice were assigned densities of 2.0 g/cm3 and 0.9 g/cm3. Bulk density values ranging between 2.7 g/cm3 and 2.8 g/cm3 were assigned to the four mineralized deposits.

| 14.6 | Resource Criteria

Based on calculated block NSR values, conceptual resource open pits and block cave underground shapes were generated for each model. Metal prices, costs (mining, processing and G&A), slope angles, and NSR cutoff grades are summarized in Table 14.10.

Table 14.10 Key Mineral Resource Parameters

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| 14.7 | Summary of KSM Mineral Resources

Mineral Resources were established for the various KSM mineralized zones using conceptual open pit and block cave shapes which were used to establish reasonable prospects for eventual economic extraction as outlined in the CIM Definition Standards for Mineral Resources and Mineral Reserves (CIM, 2014).

The conceptual open pit and underground mining shapes were generated for each resource area based on calculated block model NSR values. The NSR values were generated for each deposit by and reviewed by MMTS. The Qualified Person responsible for this section of this Technical Report generated conceptual pits for the Kerr, Sulphurets, and Mitchell deposits using MineSight® software and Lerchs-Grossmann algorithms. Golder Associates developed conceptual block cave footprints using the block NSR values and Geovia’s PCBC Footprint Finder software. The footprint polygons were extruded vertically based on guidance from Golder Associates.

Conceptual open pit shells were generated for the KSM deposits using the parameters summarized in Table 14.10 using MineSight®. Conceptual block cave shapes were defined for the Kerr, Mitchell, and Iron Cap deposits by Golder Associates using the parameters summarized in Table 14.10. Golder used PCBC software to define hypothetical draw point elevations at a Cdn$16 shutoff. The draw point extraction elevations were extruded vertically to create 3D solids that were used for resource tabulation. Conceptual caves were clipped against surface topography (Iron Cap) or conceptual resource pits (Kerr and Mitchell). Mineral Resources are summarized by resource category and mineralized zone in Table 14.11 at Cdn$9 and Cdn$16 NSR cutoffs for surface and underground resources, respectively.

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Table 14.11 KSM Mineral Resources

Notes: Mineral Resources have an effective date of December 31, 2019. The 2019 Sulphurets drill holes were not used in the construction of the resource model but were used to validate the interpretations of the model. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. It is reasonably expected that the majority of Inferred Mineral Resources could be upgraded to Indicated Mineral Resources with continued exploration. The Mineral Resources tabulated in Table 14.11 are inclusive of Mineral Reserves. Numbers may not add due to precision and roundoff of tonnes and grade.

| 14.8 | General Discussion

The QP responsible for this section of this Technical Report is not aware of any specific environmental, permitting, legal, taxation, socio-economic, marketing, political or other relevant factors, other than what is identified in this Technical Report, that could materially affect the Mineral Resource estimates that are the subject of this section.

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| 15.0 | Mineral Reserve Estimates

| 15.1 | Introduction

Mineral Reserves are based on modifying factors applied to Measured and Indicated Mineral Resources within a PFS mining study.

| 15.2 | Open Pit Reserve Parameters

Open pit Mineral Reserves use whole block grades. Open pit mining loss and dilution assumptions are shown in Table 15.1 and Table 15.2. The derivation of loss and dilution assumptions are described in Section 16.0.

Table 15.1 Pit Mining Loss and Dilution

Pit | Total Loss (%) | Dilution (%)
Mitchell | 2.2 | 0.8
Sulphurets | 5.3 | 3.9
Kerr | 4.5 | 3.2

The dilution tonnes are added as a percentage of ore tonnes from Table 15.1 at the average grade of mineralized material within the pits that is below cut-off as presented in Table 15.2.

Table 15.2 Grade of Dilution Material by Pit Area

| Mitchell | Kerr | Sulphurets
Au (g/t) | 0.21 | 0.12 | 0.26
Cu (%) | 0.04 | 0.08 | 0.04
Ag (g/t) | 1.52 | 0.64 | 0.70
Mo (ppm) | 52.0 | 2.5 | 16.0
NSR (Cdn$/t) | 6.70 | 4.70 | 6.20

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The open pit minimum NSR cut-off grade is based on an estimated process operating cost of Cdn$9.00/t. Process operating costs include plant processing (including crushing/conveying costs where applicable), G&A, surface service, tailing construction, and water treatment costs. The NSR grade used for mine planning is described in Section 16.0. A variable cut-off grade strategy has been used; a higher cut-off grade of Cdn$20.00/t is used until the end of Year 5 to maximize the NPV. During this time material between Cdn$9.00/t and Cdn$20.00/t is stockpiled and some of it is reclaimed through the mine life at the average grade of the stockpile. The premium cut-off grade in the early years of the mine schedule assists in minimizing the initial capital payback time. The cut-off grade by mine area is as follows:

| ● | Mitchell Open Pit NSR Cut-off Grade – Cdn$9.00/t to Cdn$20.00/t

| ● | Sulphurets Open Pit NSR Cut-off Grade – Cdn$9.00/t to Cdn$20.00/t

| ● | Kerr Open Pit NSR Cut-off Grade – Cdn$9.00/t.

| 15.3 | Underground Mining Reserve Parameters

The underground Mineral Reserves have been determined using block grades from the Mineral Resource model with mining dilution and losses being determined as an integral part of the caving mining analysis undertaken using GEOVIA’s PCBC™ software. The NSR grade used to determine value of the mineralized rock mucked from the drawpoints is described in Section 16.0. The site operating costs (mining and process) used in the analysis are presented in Table 15.3. The first part of the analysis determined the elevation of the production level and the shape of the production footprint at which the net value (NSR less site operating cost) of the mineralized rock to be mucked was a maximum.

Table 15.3 Site Operating Cost – Drawpoint Shut-off

| Mitchell (Cdn$/t) | Iron Cap (Cdn$/t)
Underground Mining | 6.00 | 7.00
Process | 9.00 | 9.00
Total | 15.00 | 16.00

The second part of the analysis determined a production and grade schedule based on mineralized rock mucked at the drawpoints that had a net positive value. If the net value during the mucking process is negative at any stage of the mining process, it is “shut-off”. Both the first and second parts of the analysis incorporate rock that is mucked as diluted, and the shutting-off of uneconomic drawpoints results in losses of resources. Dilution includes Mineral Resources that have grade but are sub-economic (less than drawpoint shut-off). Inferred Mineral Resources and non-mineralized rock are assumed to have zero grade. Underground mining dilution estimates from the PCBC™ assessments for the Mitchell and Iron Cap deposits are shown in Table 15.4.

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Table 15.4 Underground Mining Dilution

Deposit | Sub-economic Dilution (%) | Zero Grade Dilution (%) | Total (%)
Mitchell | 2 | 13 | 15
Iron Cap | 11 | 9 | 20

Drawpoint shut-offs (where the material being mucked becomes uneconomic) use the site operating cost (mining and process) shown in Table 15.3.

Process operating costs presented in Table 15.3 include plant processing (including crushing/conveying costs where applicable), G&A, surface service, tailing construction, and water treatment costs.

| 15.4 | Mineral Reserves

Proven and Probable Mineral Reserves are summarized in Table 15.5 and match the production plan described in Section 16.0. The Qualified Persons are not aware of any other risks, other than those identified in this Report, that could materially affect the Mineral Reserve estimates.

Table 15.5 Proven and Probable Reserves

| Ore (Mt) | Diluted Grades | Contained Metal
Au (g/t) | Cu (%) | Ag (g/t) | Mo (ppm) | Au (Moz) | Cu (Mlb) | Ag (Moz) | Mo (Mlb)
Proven | Mitchell Open Pit | 460 | 0.68 | 0.17 | 3.1 | 59 | 10.1 | 1,767 | 45 | 60
Kerr Open Pit | 0 | 0.00 | 0.00 | 0.0 | 0 | 0.0 | 0 | 0 | 0
Sulphurets Open Pit | 0 | 0.00 | 0.00 | 0.0 | 0 | 0.0 | 0 | 0 | 0
Mitchell Underground | 0 | 0.00 | 0.00 | 0.0 | 0 | 0.0 | 0 | 0 | 0
Iron Cap Underground | 0 | 0.00 | 0.00 | 0.0 | 0 | 0.0 | 0 | 0 | 0
Total Proven | 460 | 0.68 | 0.17 | 3.1 | 59 | 10.1 | 1,767 | 45 | 60
Probable | Mitchell Open Pit | 481 | 0.63 | 0.16 | 2.9 | 66 | 9.7 | 1,677 | 44 | 70
Kerr Open Pit | 276 | 0.22 | 0.43 | 1.0 | 3 | 2.0 | 2,586 | 9 | 2
Sulphurets Open Pit | 304 | 0.59 | 0.22 | 0.8 | 52 | 5.8 | 1,495 | 8 | 35
Mitchell Underground | 453 | 0.53 | 0.17 | 3.5 | 34 | 7.7 | 1,648 | 51 | 34
Iron Cap Underground | 224 | 0.49 | 0.20 | 3.6 | 13 | 3.5 | 983 | 26 | 6
Total Probable | 1,738 | 0.51 | 0.22 | 2.5 | 38 | 28.7 | 8,388 | 138 | 147
Proven + Probable | Mitchell Open Pit | 941 | 0.65 | 0.17 | 3.0 | 63 | 19.8 | 3,444 | 89 | 130
Kerr Open Pit | 276 | 0.22 | 0.43 | 1.0 | 3 | 2.0 | 2,586 | 9 | 2
Sulphurets Open Pit | 304 | 0.59 | 0.22 | 0.8 | 52 | 5.8 | 1,495 | 8 | 35
Mitchell Underground | 453 | 0.53 | 0.17 | 3.5 | 34 | 7.7 | 1,648 | 51 | 34
Iron Cap Underground | 224 | 0.49 | 0.20 | 3.6 | 13 | 3.5 | 983 | 26 | 6
Total Proven + Probable | 2,198 | 0.55 | 0.21 | 2.6 | 43 | 38.8 | 10,155 | 183 | 207

Note: All Mineral Reserves stated above account for mining loss and dilution.

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| 15.5 | Factors that could affect the mineral reserve estimate

Mineral Reserves are based on the engineering and economic analysis described in Sections 16 to 22 of this Report. Changes in the following factors and assumptions could affect the Mineral Reserve estimate:

| ● | assumptions on weather and climate

| ● | effects of climate change

| ● | metal prices

| ● | interpretations of mineralization geometry and continuity of mineralization zones

| ● | interpolation assumptions

| ● | geotechnical and hydrogeological assumptions

| ● | block caving geomechanical assumptions

| ● | operating cost assumptions

| ● | process plant and mining recoveries

| ● | ability to meet and maintain permitting and environmental license conditions

| ● | ability to maintain the social license to operate.

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| 16.0 | Mining Methods

| 16.1 | Introduction

The Mining Methods section of the 2016 PFS has been reproduced and summarized in this Technical Report without change to the mine design. A PFS-level production schedule, based on an annualized average 130,000 t/d mill feed rate, has been developed for the PFS based on a combined open pit and underground mine plan. Pit phases at Mitchell, Kerr, and Sulphurets deposits are engineered based on an economic pit limit analysis. Underground mining has been adopted at Iron Cap and below the Mitchell open pit to reduce the volume of waste generated from the mine.

| 16.1.1 | Production Rate Consideration

The 2016 PFS is based on the open pit and underground mine plans, to combine to an annual throughput of 130,000 t/d. This throughput was assessed and approved during the EA review process completed in 2015. Studies indicate there is an opportunity for improving the PFS NPV by increasing the mill throughput above 130,000 t/d early in the mine life.

The entire open pit and underground mining operation results in a mine life based on the 2016 PFS of approximately 53 years.

| 16.2 | Open Pit Mining Operations

| 16.2.1 | Introduction

The open pit mine planning work for this study is based on previous work including design criteria from the Application/EIS (Rescan 2013).

In addition to the geological information used for the block model, other data used for mine planning include the base economic parameters (metal prices, etc.), mining cost data derived from supplier estimates and data from other projects in the local area, recommended prefeasibility pit slope angles (PSAs), projected metallurgical recoveries, plant costs, and throughput rates.

| 16.2.2 | Mining Datum

The 2016 PFS design work is based on NAD83 coordinates. Historical drill hole information is based on various surveys with different sets of control that have been converted to NAD83. Topography is described in Section 12.1.6.

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| 16.2.3 | Open Pit Production Rate Considerations

The ore production rate is maintained at 130,000 t/d up to Year 35. At the introduction of underground mining at Mitchell in Year 23 and continuing through Iron Cap ore production commencing in Year 32, the open pit production is adjusted so that the combined open pit and underground production matches the maximum mill throughput. The underground mine plans are described in detail in Section 16.3.

After Year 35, the underground production becomes the base production plan with reduced mill throughput to conform to the release of ore from the block caves. The open pit mine plan is further adjusted to provide a uniform feed tonnage at the reduced rates.

| 16.2.4 | Open Pit Mine Planning 3D Block Model

Three of the four MineSight 3DBMs as described in Section 14.0 of this PFS were updated since the 2016 PFS was issued. Only the Mitchell block model remains unchanged. The effect of those updates has been verified to not be material with respect to open pit designs and associated reserve tonnes and grades that were reported in 2016. As such, the open pit designs and reserves of the 2016 PFS can now be deemed as being based on the Resource Model described in Section 14.0 of this PFS.

The block heights represent a suitable bench height for large-scale mining shovels, and the block dimension are suitably sized for long-range planning.

Net Smelter Return

NSR per tonne (net of off-site concentrate treatment and smelter charges and including on-site mill recovery) is estimated for each block and is used as a cut-off item for break-even ore/waste selection, as well as for the grade bins used to optimize cash flow in the open pit production scheduling. It is also used for the underground mine planning as described in Section 16.3.

NSR is estimated using net smelter price (NSP) and process recovery as shown in the equation below. The NSP is based on base case metal prices; US dollar exchange rate; and typical off-site losses, transportation, smelting, and refining charges. The terms of a final smelter schedule will be negotiated during the course of the mine development. The major smelter terms used to estimate NSP are specified in Table 16.1, not including minor terms for deductions/losses, payables, price participation, etc.

Where:

| – | Cu = copper grade (%) from the CUIDW 3DBM item
| – | Au = gold grade (g/t) from the AUIDW 3DBM item
| – | Mo = molybdenum grade (ppm) from the MOIDW 3DBM item
| – | Ag = silver grade (g/t) from the AGIDW 3DBM item
| – | RecCu = copper recovery (%)

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| – | RecAu = gold recovery (%)
| – | RecMo = molybdenum recovery (%)
| – | RecAg = silver recovery (%)
| – | NSPCu = net smelter price for copper (Cdn$/lb)
| – | NSPAu = net smelter price for gold (Cdn$/g)
| – | NSPMo = net smelter price for molybdenum (Cdn$/lb)
| – | NSPAg = net smelter price for silver (Cdn$/g)

Table 16.1 Major Smelter Terms Used in the NSR Calculation

| Amount | Unit
Copper Concentrate
Smelting | 75 | US$/dmt
Au Refining | 8.00 | US$/oz
Ag Refining | 0.50 | US$/oz
Off-site Costs | 236 | Cdn$/wmt
Moly Concentrate
Roasting | 2.00 | US$/lb
Other Off-site Costs | 5,298 | Cdn$/wmt
Gold Dore
Au Refining + Transport | 2.00 | US$/oz

Copper-to-gold ratio in mill feed and estimated concentrate grades varies by KSM mining area. Off-site costs and NSPs are therefore different for each mining area. The metal prices and resultant NSPs used at this early stage of the study are shown by pit area in Table 16.2 and Table 16.3.

Table 16.2 Metal Prices for Reserve NSR Calculation

| Metal Price (US$)
Cu | 2.70/lb
Au | 1,200/oz
Ag | 17.50/oz
Mo | 9.70/lb
Exchange Rate (US$:Cdn$) | 0.83

Table 16.3 Estimated NSP by Mining Area

| Cu NSP (Cdn$/lb) | Au NSP (Cdn$/g) | Ag NSP (Cdn$/g) | Mo NSP (Cdn$/lb)
Mitchell | 2.82 | 41.6 | 0.551 | 6.5
Kerr | 2.68 | 40.7 | 0.529 | 6.5
Iron Cap | 2.82 | 41.6 | 0.551 | 6.5
Sulphurets | 2.82 | 41.6 | 0.551 | 6.5

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Metallurgical recoveries used for the NSR calculation are based on test work conducted by G&T and evaluated by Tetra Tech and are described in Section 13.0.

Mining Loss and Dilution

The 2016 PFS involves a large gold-copper porphyry deposit and the orebody occurs relatively continuously within the cut-off grade shells. The pits will be mined with large shovels and trucks at an ore mining rate of 130,000 t/d. As is typical of large porphyries, blast hole assays will be used to determine the waste/ore boundaries for material designations on the pit bench for daily operations.

The Mineral Reserves used for scheduling are calculated from grades in the 3DBM using detailed pit designs with the appropriate mining recoveries and dilutions applied. The recoveries and dilutions convert the in situ ore tonnages into ROM delivered tonnage to the mill. The ROM delivered tonnage (i.e., what the mill will actually “see”) is used to determine the appropriate production schedule.

There are three main parts to recovery and dilution:

| ● | dilution of waste into ore where separate ore and waste blasts are not possible

| ● | loss of ore into waste where separate ore and waste blasts are not possible

| ● | general mining losses and dilution due to handling (haul back in truck boxes, stockpile floor losses, etc.)

In addition to the whole block dilution and the general mining losses and dilution, allowance is made for the contacts between ore and waste on the mining bench as defined by the NSR cut-off. This is affected by the size of the ore areas on the bench and the relative amount of edges. On a block-by-block basis, this is determined by the number of waste neighbours an ore block has or vice versa for waste. For this 2016 PFS, the Mitchell area has more massive ore zones on a bench than the other areas; therefore, contact dilution for this area is less. For this 2016 PFS, MMTS has estimated a mining loss and dilution factor that varies by pit area. Mining loss and dilution assumptions by pit area are provided in Section 15.0.

Since the dilution material on the contact edge of the blocks described above is mineralized, it will have some grade value. The dilution grades are estimated by determining the grades of the envelope of waste in contact with ore blocks inside the pit delineated area. These dilution grades are estimated by statistical analysis of grades in blocks with NSR less than the cut-off NSR. The dilution grades are shown in Section 15.0.

| 16.2.5 | Pit Slope Design Angles

Overview

BGC has provided open pit slope design parameters for the three proposed open pits of the KSM 2016 PFS: Kerr, Sulphurets, and Mitchell. The design parameters are based on geotechnical site investigations, available local and regional geological data, and well-established geotechnical design methods used to estimate the 2016 PFS design pit slope angles.

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BGC has identified geotechnical rock mass units associated with the primary rock and alteration types, based on the results of the site investigation and geological interpretations by Seabridge. Major geological structures (faults and foliation) have been included in the geotechnical slope stability analyses for each pit. Slope stability analyses were conducted using industry standard limit-equilibrium software, finite element analysis software, and in-house proprietary BGC tools.

BGC completed hydrogeological studies for each of the proposed pits, and numerical simulations of pit dewatering/depressurization have been carried out. BGC interpreted hydrostratigraphic units, estimated hydraulic conductivity and storage parameter values, and formulated a conceptual hydrogeologic model for the study area. The conceptual model was used as the basis for developing a numerical hydrogeologic model. The calibrated numerical model was used to evaluate the effort required to depressurize the open pit slopes to satisfy geotechnical constraints identified in the open pit slope designs. Preliminary dewatering/depressurization plans, including the number of vertical wells, horizontal drains, and the extraction rates required to achieve sufficient depressurization of the rock mass were developed to support the costing study. In addition, the need for a dewatering adit and associated drainage gallery was identified and simulated to achieve the depressurization targets of the upper north slope of the Mitchell pit.

BGC reviewed the proposed pit areas and surrounding terrain for potential geohazards, including the identification of snow avalanche paths and potential landslides, utilizing aerial photographs and satellite imagery. BGC completed ground-truthing of potential geohazards; the preliminary design of mitigation structures were completed by those responsible for the various 2016 PFS facilities at risk from the identified geohazards.

Mitchell Pit Design

The proposed Mitchell pit will be located within a glacially modified valley and targets a mineral deposit located in the valley floor, resulting in 1,200 m high ultimate slopes. This scale of the Mitchell pit north and south slope heights will rival some currently operating, very mature pits elsewhere in the Americas.

A multi-component site investigation program was completed to provide data for the Mitchell pit design work. Approximately 4,100 m of geotechnical drilling was completed, distributed over 10 core holes. BGC geotechnically logged the holes. Optical and acoustic televiewer surveys were completed in each hole to provide geological discontinuity orientations for rock slope design. Packer testing was undertaken in each hole, and vibrating wire piezometers were installed. Photogrammetric mapping of sections of the north and south valley walls was completed to provide additional data on the rock mass fabric of the study area.

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A laboratory testing program was completed, consisting of the following tests:

| ● | uniaxial compressive strength (16 tests)

| ● | Brazilian tensile strength (31 tests)

| ● | small scale direct shear testing (8 tests)

| ● | grain size and index testing (4 tests)

| ● | specific gravity (44 tests).

An appropriate quantity of quality data was collected to characterize the geological units of the study area and support PFS-level slope designs.

The structural geology of the Mitchell study area is defined by faults, foliation, and rock mass fabric (joints, etc.). In the Mitchell area the major faults are the Mitchell and Sulphurets thrust faults and the Brucejack fault. The proposed Mitchell pit has been divided into four geotechnical domains, based on the different structural geology fabrics in the area; discontinuity sets and geotechnical units for each domain are identified for use in the slope designs. Design sectors are based on the anticipated main orientations of the proposed pit walls, as determined from previous pit optimization studies.

Recommended inter-ramp slope angles vary from 34° to 54° based on wall orientation, overall wall height, geotechnical domain, and controls on slope stability. Inter-ramp slope heights are limited to 150 m, after which a geotechnical berm (or ramp) with a minimum width of 20 m is required. The inter-ramp height limits and geotechnical berms provide flexibility in the mine plan to mitigate potential slope instability; access for slope monitoring installations; and working space for in-pit wells, drains, and other water management infrastructure. All final pit slopes are assumed to be excavated using controlled blasting. Depressurization of the proposed pit slopes requires a combination of vertical wells, horizontal drains, and a dewatering adit with drainage galleries. The east and west overall slopes of the proposed Mitchell pit are within the range of slope heights that have been achieved in other porphyry metal mines in the world.

Sulphurets Pit Design

The proposed Sulphurets pit will be located on a glacially modified ridge between the Mitchell and Sulphurets valleys. The proposed mine plan would result in ultimate pit slopes with maximum heights of approximately 650 m, and a footprint of approximately 2 km x 1 km, with the long axis of the pit trending parallel to the strike of the STF.

A site investigation program including geotechnical drilling and hydrogeological testing was completed in 2010. Data from five geotechnical drill holes (consisting of approximately 1,950 m of drilling) was used to divide the Sulphurets Zone into three geotechnical domains: the hanging wall of the STF, the footwall of the STF, and an altered (crackled) zone associated with and defined by the STF. Additional joint and bedding sets have also been identified.

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Laboratory testing of core samples from the completed geotechnical drilling included:

| ● | uniaxial compressive strength (13 tests)

| ● | Brazilian tensile strength (20 tests)

| ● | small scale direct shear tests of natural discontinuities (5 tests)

| ● | index testing of discontinuity infilling material (3 tests).

The rocks of the Sulphurets Zone are typically moderately strong when weathered and strong when fresh. The RQD of the rocks of the Sulphurets Zone varies from fair to good, generally increasing in quality with depth below surface or distance from the STF.

The slope designs assume final walls will be excavated using controlled blasting, consistent with the approach proposed for the Mitchell pit. The recommended inter-ramp slope angles vary from 36° to 50° based on wall orientation, overall wall height, rock mass quality, and structural controls on slope stability. Inter-ramp slope heights are limited to 150 m after which a geotechnical berm (or ramp) with a minimum width of 20 m is required. Depressurization of the pit slopes is required and should be achievable with a combination of vertical wells and horizontal drains.

Kerr Pit Design

The proposed Kerr open pit is located on the south side of the Sulphurets Valley near the height of land and above the Sulphurets Glacier. The proposed mine plan will result in ultimate pit slopes approximately 600 m high, with a proposed pit footprint of approximately 2 km x 0.5 km.

A site investigation program including four geotechnical drill holes (consisting of approximately 1,500 m of drilling) and hydrogeological testing was completed in 2010. Data from the site investigation was used to divide the Kerr Zone into two geotechnical domains: a central altered zone and a surrounding unaltered zone; both are composed primarily of volcanic rocks. The structural geology of the Kerr Zone includes sets of west and east dipping normal faults (dipping greater than 60°) as well as bedding and joints.

Laboratory testing of core samples from the geotechnical drilling included:

| ● | uniaxial compressive strength (10 tests)

| ● | Brazilian tensile strength (14 tests)

| ● | small scale direct shear tests of natural discontinuities (4 tests)

| ● | index testing of discontinuity infilling material (3 tests).

The rocks of the altered zone are typically medium-strong but are highly fractured with poor RQD values. The rocks of the unaltered zone are strong to very strong, with good to excellent RQD values.

The slope designs assume that final walls will be excavated using controlled blasting. The recommended inter-ramp slope angles vary from 34° to 50°; based on overall wall height, wall azimuth, rock mass quality, and geological structures. Inter-ramp slope heights are limited to 150 m after which a geotechnical berm (or ramp) with a minimum width of 20 m is required. Depressurization of the pit slopes is required and should be achievable with a combination of vertical wells and horizontal drains.

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Slope Design Implementation

Achieving the proposed design criteria will require depressurization of the pit walls through the use of vertical wells and horizontal drains. Geological structures may affect bench and inter-ramp scale slope stability, and therefore, depressurization of these structures will be required.

Based on groundwater modelling results, approximately 76 in-pit wells will be required over the life of mine for the Mitchell pit. The total drilling length for the vertical wells is estimated to be approximately 15,200 m. In addition, a 3.5 km adit and drainage gallery will be required for the Mitchell pit north wall, and approximately 876 km of horizontal drains will be required to aid in depressurization of the pit slopes over the mine life. The average annual groundwater extraction rate for Mitchell pit depressurization measures is estimated to be approximately 12,600 m3/d throughout the life of the pit.

The average annual groundwater extraction rate for the Kerr pit depressurization measures is estimated to be approximately 1,300 m3/d. Approximately 36 vertical wells with a total drilling length of 7,200 m will be required throughout the life of the pit. In addition, it is estimated that approximately 110 km of horizontal drains will be required to aid in depressurization of the pit slopes over the life of the pit.

The average annual groundwater extraction rate for the Sulphurets pit depressurization measures is estimated to be 1,100 m3/d. Approximately 34 vertical wells with a total drilling length of 6,800 m will be required throughout the life of the pit. In addition, it is estimated that approximately 187 km of horizontal drains will be required to aid in depressurization of the pit slopes over the life of the pit.

Further characterization of the hydraulic properties of the bedrock at the feasibility stage and dewatering and depressurization response must be monitored throughout mining operations to determine if targets are being met.

Monitoring of pit slope displacements at various scales will be required. Inter-ramp and overall scale slopes should be monitored for deformations. The slope deformation monitoring system designed for the Mitchell pit will meet or exceed the size and complexity of those systems currently in operation at other large open pits elsewhere. The monitoring system should include multiple robotic-theodolites and survey prisms, mobile slope stability radar units, slope inclinometers, piezometers, and extensometers. The system would be computerized and use radio telemetry or a similar technology to provide real-time data to on-site geotechnical and mining staff. Similar monitoring systems would also be required for the Sulphurets and Kerr pits; the requirements of those systems would be scaled according to the proposed wall heights for those pits.

It will be important to manage geological hazards during mining operations. Additional engineered structures adjacent to the pit, or modifications to the pit slope geometry, may be required to mitigate the risk of snow avalanches. In addition, the 2016 PFS area has been recently de-glaciated and large-scale slope deformation features have been identified in the Mitchell and Sulphurets valleys.

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Nine large landslides in the study area will require management during construction and operations. Conceptual management plans detailing monitoring and mitigation measures were prepared as part of the 2016 PFS. Of particular note with respect to the open pits are: the Snowfield Landslide situated on the south slope of the Mitchell Valley and east of the Mitchell open pit, and the Kerr landslide situated on the south slope of the Sulphurets Valley and below the elevation of the proposed Kerr open pit.

The overall landslide management plan for the 2016 PFS uses a risk-based approach to determine the level of monitoring required for each landslide. The management plan for the Snowfield Landslide is comprehensive due to its proximity to the Mitchell open pit. The plan includes surface and subsurface deformation monitoring, surface water management, pumping wells, and a depressurization adit.

| 16.2.6 | Economic Pit Limits, Pit Designs

Pit Optimization Method

The economic pit limit is selected after evaluating Lerchs-Grossmann (LG) pit cases.

The assessment is carried out by generating sets of LG pit shells by varying revenue assumptions to test the deposit’s geometric/topographic and pit slope sensitivity.

The ultimate pit limit is typically determined by estimating the pit size where an incremental increase in-pit size does not significantly increase the pit resource. The selected pit limit is chosen where the economic return starts to significantly drop off. Economics of the selected pit limits are also tested to determine that they are economically viable.

LG Pit Assumptions

Inputs to the updated LG pit limit assessment shown in Table 16.4 are based on the previous PFS studies as a starting point for the 2016 design work.

Table 16.4 LG Pit Limit Primary Assumptions

Assumption | Value
Mining Cost | Cdn$1.90/t
Process, G&A, Site Services, Water Treatment | Cdn$9.00/t
Process Recoveries | See Section 17.0
Pit Slope Angle | Variable See Section 16.2.5
Metal Prices | See Table 16.2

LG pits are generated by varying prices in the range from 30% to 150% of the base NSR.

LG Economic Pit Limits

Pit shell cases are created by varying the input LG prices. Figure 16.1 to Figure 16.3 summarize the revenue sensitivity cases for the Mitchell, Sulphurets, and Kerr pits, respectively.

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The potential economic pit limits are shown for each mining area where inflection points occur, where an incremental increase in-pit size does not significantly increase the pit resource, signifying where an incremental increase in the pit resource will only result in marginal economic return.

In the Sulphurets pit area, the inflection point represents the significant potential economic pit limits and is selected.

The selected economic pit limits for Mitchell and Kerr are smaller than the potential pit limit for the following reasons.

Seabridge’s objective to reduce open pit waste mined has been achieved by selecting for Mitchell, a pit limit with 40% less mill feed than the potential economic pit limit. Waste mined in the selected pit limit is 3.0 Bt less than the potential economic pit limit leaving the lower material for inclusion in the underground mine plan below the pit.

Figure 16.1 Mitchell Sensitivity of Ore Tonnes to Pit Size

Source: MMTS, 2016

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Figure 16.2 Sulphurets Sensitivity of Ore Tonnes to Pit Size

Source: MMTS, 2016

Figure 16.3 Kerr Sensitivity of Ore Tonnes to Pit Slope and Pit Size

Source: MMTS, 2016

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For Kerr the pit limit has been selected for a pit size to ensure that waste placement backfilled into Sulphurets pit does not exceed the available capacity as described in Rescan (2013).

The selected open pit limits are summarized below:

| ● | Mitchell – open pit/underground: 60% Price Case

| ● | Sulphurets – inflection pit case: 90% Price Case

| ● | Kerr – inflection pit case: 75% Price Case.

A plan view and north-south section views of the LG pits for the open pit mining areas are shown in Figure 16.4 through Figure 16.7.

Figure 16.4 Plan View of the KSM LG Pit Limits

Source: MMTS, 2016

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Figure 16.5 Mitchell Open/Underground Pit and Economic Pit Limit – North-South Section at East 422950, Viewed from the East

Note: NSR values are in Cdn$ per tonne

Figure 16.6 Sulphurets Economic Pit Limit – North-South Section at East 421725, Viewed from the East

Note: NSR values are in Cdn$ per tonne

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Figure 16.7 Kerr Economic Pit Limit – East-West Section at North 6258800, Viewed from the South

Note: NSR values are in Cdn$ per tonne.

| 16.2.7 | Detailed Pit Designs

PFS-level pit designs demonstrate the viability of accessing and open pit mining the Measured and Indicated Mineral Resources at the KSM site. Pit designs use the selected LG pit limits as guides for estimated geotechnical parameters, suitable road widths, and minimum mining widths based on efficient operation for the size of mining equipment chosen for the 2016 PFS.

Haul Road Widths

Haul road widths are designed to provide safe, efficient haulage and to comply with the BC Mines Regulations’ minimum width specifications and safe operating practice.

Design Standards

Detailed design parameters for pits and RSFs are provided by BGC and KCB, respectively, according to their geotechnical testing and evaluations (Sections 16.2.5 and 18.2).

Minimum Mining Width

A minimum mining width between pit phases is allocated to maintain a suitable mining platform for efficient mining operations. This is established based on equipment size and operating characteristics. For this study, the minimum mining width generally conforms to 50 m, which provides sufficient room for 2-sided truck loading but, due to the configuration of merging pits, it is sometimes less.

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Access Considerations

As stated in the design criteria summary, haul road widths are dictated by equipment size. One-way haul roads must have a travel surface more than twice the width of the widest haul vehicle. Two-way roads require a running surface more than three times the width of the widest vehicle planned to use the road. All roads have a maximum grade of 8% suitable for mines with winter conditions.

Variable Berm Width

Pit designs for KSM are designed honouring overall PSAs, a nominal bench face angle (60° to 70°) and variable safety berm widths with a minimum 8 m width. Due to the low overall PSAs and double benching between berms, berm widths are generally greater than 15 m. Where haul roads intersect designed safety benches, the haul road width is counted towards the safety berm width for the purpose of calculating the maximum overall PSA.

Bench Height

The KSM pit designs are based on the digging reach of the large shovels (15 m operating bench) with double benching between high wall berms; therefore, the berms are separated vertically by 30 m. Single benching will be employed, if required, to maximize ore recovery and maintain the safety berm sequence as warranted.

LG Phase Selection

The LG selected pit cases discussed previously are used to evaluate alternatives for determining the economic pit limit and the optimal push-backs or phases before commencing detailed design work. LG pits provide a geometrical guide to detailed pit designs. Among the details to be added are roads and bench access, the removal of impractical mining areas with a width less than the minimum, and to ensure the pit slopes meet the detailed geotechnical recommendations.

There are smaller pit shells within the economic pit limits that have higher economic margins, due to their lower strip ratios or better grades than the full economic pit limit. Mining these pits as phases from higher to lower margins maximizes revenue and minimizes mining cost at the start of mining operations.

The description of the detailed pit designs and phases in this section uses the following naming conventions:

| ● | The letters M, S, and K signify Mitchell, Sulphurets, and Kerr, respectively.

| ● | The digit signifies the pit phase number.

| ● | A suffix of ‘i’ indicates that the reserve tonnage for the phase is incremental from the previous phase. If there is no ‘i’ specified, it is cumulative within the pit, up to the phase indicated.

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Mitchell Pits

For Mitchell pit, the LG “selected” pit limit as described above has been used as the economic pit limit guide for the detailed design. Mitchell pit has five incremental phases. Pit phase M1 enables the mine to have sufficient exposed ore with a six month pre-strip period. Phases M2i and M3i mine south and north respectively to provide low strip ratio ore to the mill during the pay-back period. The final two pit phases, M4i and M5i, are high strip ratio that mine to the ultimate pit limit in the south and then in the north.

The Mitchell pit phases have been designed to mine vertically through the Snowfield Landslide on the southeast side of the pit and not undermine it.

A plan view of the Mitchell pit phases is shown in Figure 16.8.

Figure 16.8 Plan View of Mitchell Pit Phases

Sulphurets Pits

The mine plan for the Sulphurets area includes 4 mining phases, which are designed using the LG “potential” economic pit limit as the ultimate pit limit guide.

S1 is a quarry that provides non-potentially acid generating (NPAG) monzonite for construction of the WSD during the pre-production period. Some of the S1 pit is outside the LG shell pit limit guide in order to provide sufficient NPAG for construction rock. S2i and S3i are low strip ratio starter pits at Sulphurets. S4i is the final Sulphurets pushback.

A plan view of the Sulphurets pit phases is shown in Figure 16.9.

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Figure 16.9 Plan View of Sulphurets Pit Phases

Kerr Pit

The Kerr deposit is mined within the LG selected pit limit, as described above, with two pit phases: a starter pit K1 and an ultimate pit K2. All ore and waste is hauled to a primary crusher on the east side of the pit and conveyed to the Mitchell Valley using a rope conveyor, across Sulphurets valley and then onto a tunnel conveyor through the Sulphurets ridge (SMCT) to the OPC.

Initial access to the Kerr pit is established with a service road built from the bottom of Sulphurets Valley to the east side of the Kerr pit (where the crusher will be located) at the 1,460 m elevation. Access to the highest benches of Kerr will be established with a small service road, and the upper benches will be dozed down to approximately 1,800 m where haul truck access can be established to the crusher. A plan view of the Kerr pit phases is shown in Figure 16.10.

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Figure 16.10 Plan View of Kerr Pit Phases

| 16.2.8 | Open Pit Mine Plan

LOM Open Pit Production Schedule

The open pit mine production schedule is developed with MS-SP, a comprehensive long-range schedule optimization tool for open pit mines used to produce a LOM schedule that increases the NPV of the PFS.

The open pit sequence is scheduled to optimize revenues and open pit development costs. The underground mining production schedule discussed in Section 16.3 is generated based on the development requirements for each mining area, the size, and capacity of the individual Mitchell and Iron Cap block caves and then integrated into the total property production schedule. The ore production from each of the open pit and underground mining components are inserted where they provide the best contribution to the PFS economics. After inserting the underground ore production into the LOM sequence, the open pit ore targets are then adjusted to meet the mill capacity.

At start-up, all production comes from open pit sources, producing higher grades from lower cost areas (both operating and capital). In the later years of the schedule, the base ore production is from the underground, from Mitchell first and then Iron Cap is phased in. After Year 35, the mill throughput rate is reduced to match the switch to continuous underground production, and the open pits are mined to supplement the ore tonnes produced from the block caves to meet the mill requirements and to improve overall head grades. In the final years of the production schedule, the stockpile accumulated during the open pit operations is used to augment the underground production. The combined LOM schedule, including open pit, underground, and stockpile reclaim, is presented in Section 16.4. The following describes the open pit sequencing to match the combined open pit underground mine plan.

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In the open pit mine schedule, “Time 0” refers to the mill start date; full mill feed production capacity is expected in Year 2. The production schedule specifies:

| ● | pioneering: Year -6, -5, -4

| ● | pre-production: Year -3, -2, -1

| ● | first year of operations: Year 1

| ● | LOM operations: at full capacity Year 2 onward.

Open Pit Mine Load and Haul Fleet Selection

The mine load and haul fleet are selected prior to production scheduling. Previous studies and similar projects in the area have shown that the lowest cost per tonne fleet of cable shovels and haul trucks that are currently being used for large hard rock open pit mines are the 100 t bucket class shovel matched with the 360 t truck. These unit sizes are proven in operating mines around the world. Diesel hydraulic shovels (85 t bucket class) are added to the fleet when a more mobile loading unit is needed. Suitable drill sizes (311 mm hole size) are chosen to match this size of truck/shovel fleet. The following performance and costs are estimated based on the use of these large-scale mining equipment.

Productivities of the selected equipment include shovel loading times and truck haul cycle estimates for multiple pit-to-destination combinations.

Schedule Criteria

In order to optimize the PFS NPV, grade bins have been specified (based on NSR block values); the MS-SP optimizer develops a cut-off grade strategy to increase the PFS NPV. This increases mill head grades, and therefore revenues, early in the production schedule. Pit/phase precedencies are specified based on the logical progression of phase geometry (no undermining) but also determined by the timing of water diversions, bench access issues, and RSF phase sequencing. Early Sulphurets waste production is initially based on WSD construction requirements and later based on RSF rock drain requirements. Kerr pit is mined after Sulphurets pit so that Kerr waste can be backfilled to the mined out Sulphurets pit.

Because of these complexities, each pit area is scheduled in MS-SP independently and then combined in a master LOM schedule.

The primary program objective in each period is to maximize the NPV. The MS-SP NPV calculation is guided by estimated operating and capital costs, process recoveries, and metal prices. Key production schedule assumptions are shown in Table 16.5.

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Table 16.5 Production Schedule Assumptions

Assumption | Value
Operating Days Per Year | 355
Hours Per Day | 21
Daily Mill Throughput | 130,000 t/d
Haul Truck Speed Limit | 50 km/h
Haul Truck Operator Efficiency | 90%
Haul Truck Operating Efficiency | 85%
Dump and Manoeuvre Time | 1.5 min
Shovel Loading Time | 35 s/pass
Shovel Spot and Wait Time | 10 s
Shovel Operator Efficiency | 84%
Shovel Operating Efficiency | 85%

Allowance has been made for the severe snowstorms or when poor visibility requires the mine to partially or completely shut down. The cumulative lost production hours are applied as lost days in the schedule.

Cut-off Grade Optimization

The pit phase designs and sequencing are typically from higher grades to lower, to mine the higher mill feed grades early in the schedule and thereby increase the PFS revenues in the earlier years. This can be further enhanced by stockpiling low and mid-grade. The stockpiled material is then milled at the end of the production schedule. However, stockpiling also results in increased mining cost per tonne milled. Additionally, oxidation can cause significant metallurgical recovery loss in the stockpile. At some point, the cost of mining more material and the recovery loss will exceed the incremental benefit from the higher grade milled. A variable cut-off grade strategy has been applied for the KSM production schedule to improve the project economics by mining the best grade ore first, varying the stockpiling, and smoothing out the haul fleet to optimize the NPV.

Rock Storage Facilities

The RSFs are located as close to the mining areas as possible. Mitchell and Sulphurets waste is placed in Mitchell and McTagg RSFs. Kerr waste rock is backfilled into Sulphurets pit.

Further details on the RSF design are available in Section 18.0.

Construction Methods

Several different construction methods will be used for waste placement: top-down, bottom-up, and wraparounds. Top-down platform heights are limited to a maximum height of 300 m. Bottom-up lifts are 30 m to 50 m high, or less if geotechnically required. Wraparounds are built onto the face of an existing RSF, creating a series of terraces to facilitate intermediate haul roads and lower the overall slope angle of high dumps, which will reduce the end of mine closure costs.

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Foundation Preparation

Design work for RSF foundation preparation will be performed as required at the feasibility-level design stage. Prior to mine development, soil will be salvaged from the footprint area where soil is suitable for reclamation purposes. Soils will generally not be salvaged on slopes steeper than 26° due to practical limitations on equipment access and operator safety. Soils salvaged from the RSF footprints will be stockpiled in the Ted Morris Valley.

The waste in the valley bottoms will be placed in low height lifts to confine and consolidate weaker foundation material. If required, loose tills and clays will be removed. Once the foundation is prepared, the basal drain is placed on top at the required lift height. The basal drain requirements are described in Rescan (2013).

RSF Monitoring and Planning

The long-term operation of the RSFs will be similar to that of the large, steep-terrain RSFs that have been in operation for many years in southeast BC Rocky Mountain coal mines. These operations involve high-relief RSF phases with clear dumping in single lifts of up to 400 m. Clear dumping is a technique whereby truck loads are dumped directly over the crest of the dump face; the load is not dumped short and then pushed over the edge. The clear dumping technique maintains a stable dump platform but requires well-established monitoring and operating practices. Foundation preparation also needs to be assured.

RSF Access Roads

Pioneering access to each pit and subsequent phases use roads with a maximum 15% grade; these are constructed using balanced cut and fill wherever possible. Pioneering roads are 10 m wide and enable major mining equipment to reach the top benches of each pit phase and start mining. These are built for pit access and not for hauling. After the pioneering road is established to the top benches of each pit phase, bench waste from the upper portions of each pit phase is used to fill full-width haul roads at a maximum gradient of 8% at the 38 m double lane width, to connect with permanent surface roads and high wall roads in the long-term road network. This road network connects the mining areas with the primary crusher and stockpile areas for ore and the RSF areas for waste.

As described earlier, the terraced RSFs on the south side of the Mitchell Valley provide level access to the south Mitchell Valley RSF platforms.

Final RSF Configuration

The RSFs for the PFS will have overall slope angles of 26° to 30°. The final post-closure configuration will be adapted in accordance with the closure plan as identified within Rescan (2013) and described in Section 20.7. Costs for this work are included during the later years of the operation, when the waste strip ratio drops to low levels and ancillary equipment then becomes available for other duties.

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Low-grade ROM Stockpile

Lower-grade ore is stockpiled throughout the mining schedule to follow the cut-off grade strategy, and then is reclaimed in later years. This increases the grade of the ore feed to the plant in the early years of the schedule but also is used to even out the waste mining requirements as required during periods of high pre-stripping for some of the pit phases. The low-grade ore stockpile is placed to the west of the Mitchell OPC. Provision has been made for an HDPE pipeline diversion around the surface of the stockpile, which can be moved as required.

Open Pit Mine Pre-production Detail

Development and pre-production activities include the following:

| ● | expose sufficient ore for start-up

| ● | establish mining areas that will support the equipment required to achieve ore production and annual mill feed requirements on a sustainable basis

| ● | provide material required for construction in the mine area

| ● | in early development, the mining fleet will excavate colluvium from a borrow source in the Mitchell Valley to provide construction fill for the Mitchell OPC

| ● | during pre-production, Mitchell pit phase M1 is mined to 885 m and M2 is mined to 1,290 m in the Mitchell Valley; and Sulphurets pit phase S3 is mined to an elevation of 1,485 m

| ● | development and pre-production must be completed by the end of Year -1. This will expose sufficient ore for commissioning and to sustain full mill production rate of 130,000 t/d of mill feed after that. The mine layout at the end of pre-production is shown in Figure 16.11.

| 16.2.9 | Open Pit Production

Year 1 to 20 – Open Pit Mining

The following is a summary of mining activity in Years 1 to 5:

| ● | Mining in Year 1 to 5 focuses on delivering the grade required to payback initial capital

| ● | All Mitchell waste and Sulphurets waste material is placed in the Mitchell RSF

| ● | Sulphurets construction rock (NPAG monzonite) waste is hauled to be used for the basal and selenium drains beneath the Mitchell/McTagg RSF

| ● | Mitchell and Sulphurets ore is hauled directly to the Mitchell primary crushers. From Year 2, the Sulphurets ore is hauled to the Sulphurets crusher and crushed ore is then conveyed to the OPC through a tunnel

| ● | An ROM ore stockpile is built in the area to west of the Mitchell OPC

| ● | The Mitchell RSF is built in lifts at an overall slope of 2:1 with an access road in the final face.

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The open pit mine layout at the end of Year 5 is presented in Figure 16.12.

The following is a summary of mining activity in Years 6 to 20:

| ● | By Year 6, the M3 and S3 phases are mined out and mining begins on the higher strip ratio phases of Mitchell and Sulphurets

| ● | Stockpile material is reclaimed to supplement mill feed during periods where mining is limited by the large volume of waste pre-stripping or vertical advance rate

| ● | By Year 10, the Mitchell RSF is full and waste placement begins in the McTagg RSF

| ● | Mitchell and Sulphurets pits are mined out by the end of Year 20. All waste is placed on the McTagg RSF.

Year 20 to 30 – Open Pit Production and Transition to Mitchell Underground Mining

The open pit mine production sequence is adjusted to meet plant feed requirements as the Mitchell block cave mine is brought online. As discussed in more detail in Section 16.3, the Mitchell block cave mine is ramped up to full production in this time interval.

Kerr open pit is mined out by Year 30. Waste from Kerr pit is conveyed to Sulphurets and backfilled into the mined out Sulphurets pit. Ore from Kerr is conveyed to the OPC.

Direct mining from the open pits is completed by the end of Year 30.

Year 30 to 53 – Stockpile Reclaim to Supplement Underground Mining

Open pit mining after Year 30 is limited to stockpile reclaim for supplementing Mitchell and Iron Cap block cave production (see Section 16.3).

Once the stockpile is removed, a closure channel is established around the Mitchell RSF by placing moraine material and NPAG riprap on berms along the north and west toes of the Mitchell RSF. The open pit mine layout at LOM is shown in Figure 16.13.

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Figure 16.11 End of Pre-production (Year -1)

Source: MMTS (2016)

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Figure 16.12 End of Year 5

Source: MMTS (2016)

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Figure 16.13 Open Pit Life of Mine

Source: MMTS (2016)

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| 16.2.10 | Open Pit Mine Operations

The open pit and underground operations are considered as separate operations in this study, with their own facilities and management, technical, and operating personnel. Future detailed planning should be able to reduce costs by integrating some of the support services, staff and facilities. The following description is for the open pit operations. Underground operations details are provided in Section 16.3.

KSM open pit mining operations will employ bulk mining methods and large capacity equipment. Services and support are well established in BC and in the local area for this kind of operation.

The use of large-scale equipment will minimize unit operating costs, reduce the on-site labour requirement of a remote site, and dilute the fixed overhead costs for the open pit operations.

Organization

Mine operations is organized into three areas: direct mining, mine maintenance, and general mine expense (GME).

In this study, direct mining and mine maintenance are planned as an owner-operated fleet with the equipment ownership and labour being directly under operations. It may be possible to contract out some of the direct mining activities under typical mine stripping contracts, and maintenance and repair contracts (MARC). For this study the mine will employ the blasting crew but, the supply and onsite manufacturing of blasting materials will be contracted out. All infrastructure required for the blasting supply contractor will be provided by the operations.

Direct Mining Activities – Open Pit

The direct mining area accounts for the drilling, blasting, loading, hauling, and pit maintenance activities in the mine.

Drilling

Areas will be prepared on the bench floor blast patterns in the in situ rock. Drill ramps will be dozed on original mountain side surfaces, between benches.

Blasthole drills will be fitted with GPS navigation and drill control systems to optimize drilling.

Diesel hydraulic and electric rotary drills (311 mm bit size) will be used for production drilling, both in ore and waste.

Diesel hydraulic percussive drills with a hole size of 6.5 inches (165 mm) will operate production benches for controlled blasting techniques on high wall rows, pioneering drilling during pre-production, and development of initial upper benches.

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Blasting

Powder Factor

An appropriate powder factor has been used to provide adequate fragmentation and digging conditions for the shovels. Similar large open pit projects in the KSM area use a powder factor of 0.32 kg/t for competent rock. Based on a blasting study by Orica, a power factor of 0.35 kg/t is being used. Future Feasibility Study planning can investigate further mine to mill performance with respect to blasting.

Explosives

A contract explosives supplier will provide blasting materials and technology. Due to the remote nature of the operation, an explosives manufacturing plant will be built on site when emulsion is required. For this study, the owner provides a serviced site and all facilities to the explosives contractor who manufactures and delivers the prescribed explosives to the blast holes and supplies all blasting accessories.

It is anticipated that roduction up to and including Year 1 will not require emulsion. After Year 1 it is assumed that half of the holes will use a 70/30 emulsion/ammonium nitrate-fuel oil (ANFO) mix explosive (“wet” product) and half of the holes will use a 35/65 emulsion/ANFO mix (“dry” product). Higher use of ANFO, and possible use of borehole liners to keep the ANFO dry, can be investigated in future studies to reduce blasting costs.

Blasting accessories will be stored in magazines adjacent to the mining areas suitably located to meet federal and provincial regulations and to avoid potential geohazards.

Explosives Loading

Loading of the explosives will be done with bulk explosives loading trucks provided by the explosives supplier. The trucks should be equipped with GPS guidance and should be able to receive automatic loading instructions for each hole from the engineering office. The GPS guidance will be a necessity to be compatible with stakeless drilling.

A smaller “goat” truck is needed for small development areas as well as for squaring-off blast patterns when the mine roads have been closed due to excessive snow fall. “Goat” trucks are similar to a logging skidder with high manoeuvrability and are a specific adaptation for open pit operations in mountainous and high snow fall areas.

Blast holes will be stemmed to avoid fly-rock and excessive air blasts. Crushed rock will be provided for stemming material and will be dumped adjacent to the blast pattern. A loader with a side dump bucket is included in the mine fleet to tram and dump the crush into the hole.

Blasting Operations

The blasting crew will comprise mine employees and will be on day shift only. The blasting crew will coordinate drilling and blasting activities to ensure a minimum of two weeks of broken material inventory is maintained for each shovel. Blasters will require hand-held GPS to identify holes for pattern tie-in as blast patterns will not be staked. A detonation system will be used that consists of electric cap initiation, detonating cord, surface delay connectors, non-electric single-delay caps, and boosters.

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Blasting assumptions are summarized in Table 16.6. These parameters are typical for other mines in the Western Canada and will be re-evaluated in the future with a detailed blasting study, using site-specific rock strength parameters.

Table 16.6 Blasting Assumptions

Blasting Pattern – Ore and Waste | Specifications
Spacing | 8.5 m
Burden | 8.5 m
Hole Size | 12¼ inches
311 mm
Explosive In-Hole Density | 1.25 g/cc
Explosive Average Downhole Loading | 95.0 kg/m
Bench Height | 15 m
Collar | 6 m
Loaded Column | 11 m
Sub-drill | 2 m
Charge per Hole | 1,046 kg/hole
Rock SG | 2.77 t/m3
Yield per Hole | 3,002 t/hole
Powder Factor | 0.35 kg/t

Loading

Ore and waste will be defined in the blasted muck pile by the OCS. A fleet management system will assist in optimizing deployment and utilization of the mine fleet

Three 85 t dipper diesel hydraulic shovels and three 100-t dipper electric cable shovels have been selected as the primary digging units. The diesel hydraulic shovels are selected for flexibility and mobility in accessing the narrow top pit benches.

Minimum bench widths of 50 m are designed to ensure sufficient operating for double-sided loading of trucks at the shovels. Where single-sided loading will be necessary and reduced productivity for the shovel will be encountered, such as the upper benches of the pit phases ancillary equipment will be deployed to prepare the digging areas for higher shovel productivity. This can entail dozing small benches down slope to the next bench, trap dozing, and other dozing activities.

Hauling

Ore and waste will be hauled by 360 t off-highway haul trucks. Haul productivities have been estimated from pit centroids at each bench to designated dumping points for each time period.

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Pit Maintenance

Pit maintenance services include haul road maintenance, open pit mine dewatering, transporting operating supplies, relocating equipment, and snow removal.

Haul road maintenance is paramount to low haulage costs; dozer and grader hours have been allocated to maintain the haul road network throughout the LOM production schedule.

A fleet of ancillary service vehicles are allocated to install and service the in-pit sump pumps and the high wall horizontal drains. This includes connecting these pumps to the pit dewatering pipeline system. This crew will also service and supply mobile light plants.

A fleet of service equipment is allocated for summer season construction and will be used in winter for snow clearing and spreading crushed rock for traction control. This includes scrapers and loaders. The snow fleet will be manned by mine operations staff in normal winter conditions with operators taken from reduced activities such as dust control and summer field programs. During severe storms, personnel to operate the standby snow fleet will be drawn from truck and shovel operations as the long-haul fleets shut down. This will ensure priority fleets remain operating.

A rock crusher for road grading material is included.

Open Pit General Mine Expense Area

The GME area accounts for the supervision, safety, environment, and training for the direct mining activities as well as technical support from mine engineering and geology functions. Open pit mine operation supervision will extend down to the shift foreman level and trainers.

GME costs also include engineering consulting on an ongoing basis for specialty items, such as geotechnical and geo-hydrology expertise, and third-party reviews in the open pit mine area.

| 16.2.11 | Mine Closure and Reclamation

Details on mine closure and reclamation are available in Section 20.7

| 16.2.12 | Open Pit Mine Equipment

Mining equipment descriptions in this section provide general specifications so that dimensions and capacities can be determined from vendor specification documents.

Major Equipment

The production requirements for the major mining equipment over the LOM are summarized in Table 16.7. The current production schedule requires a maximum haulage fleet of 60 trucks over the LOM.

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Table 16.7 Major Equipment Requirements

| Pre- production | Year 5 | Year 10 | Year 20 | Maximum
Drilling
Primary Drill – 311 mm Electric Drill | 0 | 4 | 5 | 5 | 5
Primary Drill – 311 mm Diesel Hydraulic Drill | 2 | 3 | 3 | 1 | 3
High Wall Drill – 150 mm Diesel Hydraulic Drill | 4 | 4 | 4 | 4 | 4
Loading
Primary Shovel – 40 m3 Diesel Hydraulic Shovel | 1 | 2 | 3 | 3 | 3
Primary Shovel – 56 m3 Electric Cable Shovel | 0 | 3 | 3 | 3 | 3
Construction Shovel – 12 m3 | 2 | 0 | 0 | 0 | 2
Hauling
Haul Truck – 360 t | 15 | 37 | 60 | 58 | 60
Construction Haul Truck – 90 t | 14 | 0 | 0 | 0 | 14

Drilling Equipment

The primary production drilling will be carried out in ore and waste with electric rotary drills with a 311 mm hole size. The production drills will be fitted with GPS navigation and drill control systems to optimize drilling. Production drilling assumptions are listed in Table 16.8.

Table 16.8 Open Pit Production Drilling Assumptions

Production Drill – Mineralized Material & Waste | Electric Rotary | Diesel Rotary
Bench Height | 15 m | 15 m
Subgrade | 2.0 m | 2.0 m
Hole Size | 311 mm | 311 mm
Penetration Rate | 40.0 m/h | 40.0 m/h
Hole Depth | 18 m | 18 m
Over Drill | 1.0m | 1.0 m
Setup Time | 2.0 min | 2.0 min
Drill Time | 27.0 min | 27.0 min
Move Time | 2.0 min | 2.0 min
Total Cycle Time | 31.0 min | 31.0 min
Holes per Hour | 1.94 | 1.94
Re-drills | 6% | 6%

A 150 mm diesel percussive drill is also specified for controlled blasting techniques on high wall rows in all pit phases, pioneering drilling during pre-production, and development of initial upper benches.

A detailed drill study is recommended for more advanced project studies to better determine the penetration rate that can be expected for the selected drills and the specific rock types that exist within the pit areas.

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Blasting Equipment and Facilities

A blast hole stemming unit will be required to load cuttings into the hole and stem the unloaded portion of the hole. This unit will be provided by the KSM operation. Blasting activities are detailed in Section 16.2.10.

Open Pit Loading and Hauling Equipment

The shovel-truck fleet selected for KSM is the 56 m3 dipper class of electric shovel, and the 360 t payload class of truck. A 40 m3 dipper class diesel hydraulic shovel is also required for difficult to access development benches and enables pre-production mining before power is established to the mine site. Loading and hauling is discussed in Section 16.2.10.

Open Pit Dewatering Equipment

The dewatering activities will include the following:

| ● | horizontal drain holes in bench faces

| ● | sloped pit floors as required

| ● | in-pit sumps

| ● | vertical dewatering wells

| ● | a dewatering tunnel behind the north high wall

| ● | water collection system

Pit water will be collected and transported to the WSF.

Open Pit Support Equipment

The mine support equipment fleet requirements are summarized in Table 16.9. The fleet size in Year 5 and Year 10 is shown as representative of the LOM requirement.

Table 16.9 Mine Support Equipment Fleet

Fleet | Function | Year 5 | Year 10
Hole Stemmer – 3 t | Blast Hole Stemmer | 2 | 2
Track Dozer – 430 kW | Shovel Support | 5 | 6
Rubber Tired Dozer – 350 kW | Pit Clean Up | 2 | 3
Fuel/Lube Truck | Shovel and Drill Fueling and Lube | 2 | 3
Wheel Loader Multipurpose – 14 t | Pit Clean Up | 2 | 3
Water Truck – 20,000 gal | Haul Roads Water Truck | 2 | 2
Track Dozer – 430 kW | Dump Maintenance | 3 | 3
Motor Grader – 400 kW | Road Grading | 4 | 4
Tire Manipulator | Tire Changes | 3 | 3

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Open Pit Ancillary Equipment

The mine ancillary equipment fleet is listed in Table 16.10. The fleet sizes in Year 5 and Year 10 are shown as representative of the LOM requirement.

Table 16.10 Open Pit Ancillary Equipment Fleet

Fleet | Function | Year 5 | Year 10
Track Dozer – 430 kW | Pit Support | 2 | 2
Float Tractor/Trailer – 189 t | Float Tractor and Trailer | 1 | 1
Hydraulic Excavator – 6 t | Utility Excavator | 2 | 2
Sump Pump – 1,400 gal/min | Pit Sump Dewatering | 6 | 6
Light Plant | Lighting Plant | 6 | 8
250 t Crane | Utility Crane | 2 | 2
Crew Cab | Supervision and Crew Transportation | 18 | 18
Ambulance | Ambulance | 1 | 1
Hydraulic Excavator – 4 t | Utility Excavator | 4 | 3
Mine Rescue Truck | Rescue Truck | 1 | 1
Crew Bus | Crew Bus | 5 | 5
Maintenance Truck – 1 t | Maintenance Truck | 5 | 5
Fire Truck | Fire Truck | 1 | 1
Screening & Crushing Plant – 12” max. | Road Crush and Stemmings | 1 | 1
Picker Truck | Maintenance + Overhauls | 2 | 2
Scraper – 37 t | Crush Haul for Winter Roads etc. | 5 | 5
Crane 40 t Hydraulic Extendable | Utility Crane | 2 | 2
Wheel Loader – 14 t | Crusher (Road Crush) Loader | 1 | 1
Snowcat | Winter Off Road Crew Transport | 6 | 6
40 t Crane | Utility Crane | 2 | 2
Forklift – 30 t | Forklift | 1 | 1
Forklift – 10 t | Forklift | 2 | 2
Service Truck | Service Truck | 5 | 5
Welding Truck | Welding Truck | 4 | 4
Powerline Truck | Powerline Maintenance | 2 | 2

Snow Fleet

All of the following snow fleet equipment is chosen to start operating during pre-production and continue to the end of mine life, unless otherwise noted.

| ● | Five scrapers with the ability to haul 37 t are included in the fleet. The scrapers are required to haul and spread crushed rock for traction control and remove snow from the haul roads and mine working areas as necessary. The scrapers are also used on occasion for small earthmoving jobs and reclamation projects outside of the snow season.

| ● | One wheel loader with an approximately 14 t bucket to clear snow from the plant area and truck shop, as well as ancillary routes within the mine. The wheel loader is also used to load the cone crusher at the crushing and screening plant.

| ● | Six snowcats to transport operators to equipment in a location that is inaccessible to the crew bus or vans because of heavy snowfall.

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The snow fleet has a low utilization as it is only required in winter. Other than the use of the scraper for summer construction projects and stockpiling road crush, operating the snow fleet equipment outside winter is not currently scheduled.

Open Pit Ancillary Facilities

Shops and Offices

In addition to providing an area for maintenance bays, tire shops, and a wash bay, the maintenance shop will also house:

| ● | a welding bay

| ● | an electrical shop

| ● | an ambulance

| ● | a first aid room

| ● | a first aid office

| ● | a machine shop area

| ● | a mine dry

| ● | a warehouse

| ● | offices for administration, mine supervision, and engineering/geology staff

| ● | a lunchroom and foreman’s office.

The recommended shop sizing for the open pit operations includes eight service bays, one welding bay, and three wash bays. This will accommodate the fleet for the LOM PFS production plan. The mine maintenance facility will also include a machine shop area, tool storage area, mine muster area, warehouse, and office complex. A separate tire bay facility will be required with an exterior heated pad to accommodate at least two trucks and a tire manipulator.

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| 16.3 | Underground Mining Operations

The Mitchell and Iron Cap block caves are located within the Mitchell Valley and are accessed by the MTT. Figure 16.14 shows a plan view of the block cave area and the relationship between the block caves, the MTT, and the North Pit Wall Depressurization Tunnel. A section view shows the same elements in Figure 16.15.

Figure 16.14 Plan View of the Mitchell and Iron Cap Block Cave Mines

Source: Golder, 2016

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Figure 16.15 Section View of the Mitchell and Iron Cap Block Cave Mines (Looking North)

Source: Golder, 2016

| 16.3.1 | Underground Mine Design Inputs

This subsection summarizes the block cave mine design inputs that are similar for both Mitchell and Iron Cap and includes a geotechnical review and caving analysis.

The quality of the rock mass at the Mitchell deposit is rated as good. No major structural features are identified that might influence the caving mechanism and the progression of the cave in any significant manner.

The Iron Cap deposit appears to be composed of strong, moderately fractured rock. Rock quality variations are most commonly attributed to variations in fracture frequency, as the strength of the rock mass does not vary significantly within the deposit. The fracture frequency is higher for Iron Cap than for the Mitchell deposit, resulting in a corresponding lower predicted median in situ block size of 2.5 m3, as compared to the Mitchell deposit. Several gaps in data are identified in the Iron Cap geotechnical and hydrogeological studies. These gaps will need to be addressed as part of future feasibility-level studies.

Caveability assessments for both the Iron Cap and Mitchell deposits have been completed using Laubscher’s and Mathews’ methods, which involve assessing caveability based on experience at other mining operations with similar rock quality. These methods indicate that the size (area) of a footprint required to initiate and propagate caving is between approximately 110 m and 220 m for both deposits. These dimensions are significantly smaller than the size of the deposit footprints that can potentially be mined economically by caving. This fact, together with the general large 3D shape of the deposits, suggests that both the Iron Cap and Mitchell deposits are amenable to block cave mining.

In situ stresses have been estimated at the Mitchell deposit using hydraulic fracturing tests. Based on high-induced stresses in the cave back, as predicted by numerical modelling, it is expected that stress-induced fracturing of the rock mass will contribute to caving. More sophisticated numerical analyses to confirm and quantify stress-related impacts are recommended as part of future studies.

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There have been no fracture propagation assessments applicable to preconditioning designs or in situ stress interpretations developed for the Iron Cap deposit. Measurements carried out in the Mitchell deposit may not accurately reflect the fracture propagation and stress environment at Iron Cap because of the effects of surface topography. Future drilling programs should include hydraulic fracturing tests.

A significant proportion of the rock at the Mitchell deposit is predicted to have block sizes greater than 6 m3. At Iron Cap, block sizes are predicted to be 2.5 m3. Without adopting some remediation measure, such large blocks will require significant secondary blasting, and a significant adverse impact on production and damage to the drawpoints that will require ongoing rehabilitation is likely. The cost estimates for the designs presented herein have considered remediation measures to accommodate large fragmentation.

The primary measure to accommodate the large fragmentation is to precondition the rock mass. The costs and scheduling to do this have been incorporated into this study for both cave mines. However, there are a number of uncertainties associated with preconditioning due to the limited number of caving mines where it has been applied and tested. The results from those mines employing preconditioning are encouraging, and there is sufficient experience in the industry to indicate that such fragmentation concerns do not represent a fatal flaw at either mine.

The uncertainty in the effectiveness of preconditioning to enhance fragmentation was addressed via production and cost risk mitigation measures. The average draw rate per column during steady state production is approximately half of maximum (165 mm/d), meaning there are roughly two drawpoints available for production for every one required to meet production targets. In addition, a fleet of mobile rock breakers and remote blockholers are included in the designs and costs to increase the time a drawpoint is producing, by decreasing the time it is blocked with oversize.

It is very difficult to quantify the effect of attrition as the rock is brought down within the cave except that experience has indicated that in caving mines operating under similar rock conditions to those at Iron Cap and Mitchell, fragmentation of rock drawn down more than approximately 100 m is generally good. For this study, it is assumed that fragmentation of the initial 100 m of draw height is approximately equal to the estimated in situ block size and, above this, only limited secondary blasting will be required.

The expected coarse fragmentation at Mitchell and Iron Cap will result in relatively large isolated drawcone diameters of 13 m or more, for a loading width of 5 m. The present experience in other operating mines is that a 15 m by 15 m drawpoint spacing performs well under these coarse fragmentation conditions. Some caving mines operating in good quality rock have successfully expanded the layout to 17 m by 17 m or 18 m by 15 m, but it was considered prudent at this stage of study to adopt the slightly more conservative 15 m by 15 m spacing.

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| 16.3.2 | Mitchell Underground

The Mitchell deposit extends approximately 1,500 m east-west (along strike) and 400 m to 1,400 m north-south and is between approximately 300 m and 900 m in the vertical dimension. The deposit is massive, reasonably continuous, and, in general, geometrically suitable to mine by block caving. The potential of mining the Mitchell deposit by a combination of open pit and underground methods was investigated previously (Golder 2011, Golder 2012) and these studies concluded that it is possible to mine the upper portions of the Mitchell deposit by open pit methods and the deeper portions by block caving.

Mitchell Underground Mineral Reserves

The Mineral Resource block model used for the study contains gold, copper, silver, and molybdenum grades, as presented in Section 14.0, as well as the NSR values which are described in Section 16.2.4. The model contains Measured, Indicated, and Inferred grades, but the Inferred grades were set to zero and are not included in this PFS assessment. The Mineral Resources were constrained by the PFS pit and then evaluated using GEOVIA’s PCBC™ software, to determine the Mineral Reserves for a block cave mine. Footprints at elevations of 180 m and 235 m produced the most value. Considering the footprint elevation of 235 m from the 2012 PFS (Tetra Tech, 2012) and the similar geometries of the 2012 and 2016 footprints at this elevation, it was decided to maintain the footprint elevation at 235 m. This resulted in 454 Mt of block cave Mineral Reserves, as shown in Table 16.11.

Table 16.11 Mitchell Block Cave Mineral Reserves (Cdn$15/t NSR Shut-off)

Category | Tonnes (million) | Au (g/t) | Cu (%) | Ag (g/t) | Mo (ppm)
Probable1 | 454 | 0.53 | 0.17 | 3.5 | 34

| Notes: | 1Includes 10 Mt of mineralized dilution (the portion of Measured and Indicated material that is Cdn$0 < NSR < Cdn$16) and 59 Mt of non-mineralized dilution (material at zero NSR including the Inferred material set to zero grade).

The Mineral Reserves contain dilution that include Mineral Resources that have grade, but are sub-economic (less than drawpoint shut-off), Inferred Mineral Resources that are set to zero grade, and non-mineralized material that is zero grade rock. Dilution estimates for the Mitchell block cave are 2% of sub-economic material (10 Mt) and 13% of zero grade dilution (59 Mt), for a total of 15% (69 Mt).

Mitchell Underground Mine Design

The underground mine design is based on modelling using GEOVIA’s PCBC™ software and Footprint Finder module. Footprint Finder modelling indicates that the optimum footprint for the Mitchell deposit is approximately 728 m wide in the north-south direction, 1,022 m wide in the east-west direction, and 860 m vertically, with the footprint elevation established at 235 m. PCBC modelling indicates that the block cave could produce 20 Mt/a (55,000 t/d), requiring the development of 120 new drawpoints per year. The final mine design includes approximately 218 km of drifts and raises, including a 25% design allowance to account for the excavation of infrastructure, such as service bays, fueling stations, wash bays, sumps, and electrical substations, and to account for overbreak.

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The mine design comprises six main levels: preconditioning, undercutting, extraction, secondary breakage, haulage, and conveying (Figure 16.16). In addition, the design includes a service ramp to surface to provide access for personnel, equipment and materials, and a conveyor ramp to the MTT ore bin for excavated material to be loaded on the MTT train to the mill. The floors of the extraction drifts and drawpoints are designed to be concreted, which will increase the speed and productivity of the load-haul-dump (LHD) vehicles, as well as reduce equipment maintenance. The six levels of the mine design will be accessed through internal ramps beginning on the extraction level. These ramps will be strategically positioned to maintain access to the levels during caving and to meet ventilation requirements.

Figure 16.16 Section View of the Mitchell Block Cave Mine Design (Looking South)

Source: Golder, 2016

There are 34 extraction drifts on the extraction level, and each drift is designed with three ore passes. This will reduce the average LHD haul distance to approximately 100 m and improve productivity. The ore passes from neighbouring extraction drifts will feed a stationary rockbreaker on the secondary breaking level, which will reduce the size of the material further, and feed it to the haulage level via passes with chutes. A train on the haulage level will haul the material to centrally located gyratory crushers, where it will be crushed and conveyed to the surface.

In 2012, BGC evaluated the surface disturbance and ground deformation caused by block caving the Mitchell deposit (BGC 2012), and the analysis is still applicable to this study. It was found that the MTT and Mitchell OPC are outside the zone of disturbance resulting from caving mining.

Mitchell Underground Equipment and Major Infrastructure

The proposed mobile diesel equipment is typical of that used in underground mines and will comprise units directly related to moving ore to the crushers (7 m3 LHDs, secondary rockbreakers, and the train), development equipment (4.6 m3 LHDs and 18 m3 trucks), as well as ANFO loaders and ground support machines. In addition, service equipment is included for construction, supervision, engineering, and mine maintenance activities. At peak production, Mitchell will require a fleet of approximately 55 units of mobile underground equipment.

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The mine workforce includes both staff and labour positions and the size varies according to the stage of the mine life, with a peak quantity of 379 personnel in Year 30 (in PFS years). Groundwater inflows are very small compared to surface runoff and will be readily handled by the proposed dewatering system.

The majority of the main ventilation infrastructure will be located on the extraction level. It will consist of two fresh air raises, two fresh air drifts, a fresh air ring drift, multiple internal ventilation raises, a return air drift, and two exhaust raises. An airflow of 860 m3/s is required for the Mitchell mine to achieve a production rate of 55,000 t/d, based upon the diesel equipment utilized, air velocity considerations, and an allowance of 20% per level for items such air loss around regulators, poorly installed or damaged ducting, and ventilating any inactive headings in the active mining areas. Heating the mine air in the winter months is included in the design and cost estimates. It is estimated that the Mitchell mine will require approximately 25.1 MWh of electricity at peak operation. The main contributors to this are the crushers, conveyor belts, ventilation fans, and dewatering pumps.

The mine dewatering system will require an average of 3.9 MWh, with a maximum of 29 MWh during a peak storm event, which is greater than that required to operate the entire mine under normal conditions. The strategy will be to shut down or reduce operations in the underground mine, along with other site facilities, during flooding events when the high-powered pumps are required. This will allow power to be diverted from normal operations to power the pumps.

The hydrological characterization of the site indicates that a 200-year runoff event could lead to a maximum one-day inflow of approximately 773,000 m3 of water, even with the construction of diversion ditches beyond the crest of the pit. To accommodate this inflow, the dewatering plan includes significant pumping and storage capacity underground. Two, 6.0 km long, 7.5 m x 7.5 m dewatering tunnels have been designed to convey water from the mining area to beneath the water treatment plant where eight multi-stage centrifugal pumps will lift the water and transport it to the WSD. The system is designed to handle flows at variable combined rates up to 4 m3/s.

Mitchell Underground Schedules

The mine development schedule is separated into three phases; an initial pre-production phase, which develops the primary access ramp and conveyor drifts; a second ore production phase that creates sufficient openings to start and ramp-up production from the cave; and the final phase once the mine has reached steady-state production and the development fleet is only required to create sufficient openings to maintain production. The average length of required annual development is approximately 4,000 m, with peak development occurring during the second phase, when approximately 15,000 m/a is required. The pre-production development phase for Mitchell block cave is six years.

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The mine production schedule was developed using GEOVIA PCBC™ software. It is assumed that sloughing of peripheral waste rock will occur into the crater above the cave and cover the upper surface of the material being drawn down. This was modeled in PCBC by adding an infinite supply of waste material on top of the mineralized material. As material is drawn from the drawpoints, the waste will mix with mineralized material as dilution with zero grade, and the combined material will report to the drawpoint. Due to the large fragmentation that is estimated to report to the drawpoints at Mitchell, particularly during the early stages of mining, a draw rate of 200 mm/d was chosen as a maximum draw rate in the PCBC analysis. However, an average draw rate of only 108 mm/d is required to achieve production targets (the maximum draw rate modelled never exceeds 165 mm/d, so there are roughly twice as many drawpoints available as are required to meet production targets). Initially, it is assumed that a drawpoint can produce at 60 mm/d and that this will steadily increase until 50% of a column is mined. Then, the drawpoint will produce up to the set maximum of 200 mm/d. Mitchell is estimated to have a production ramp-up period of six years, steady state production at 20 Mt/a for 14 years, and then ramp-down production for another 7 years. Figure 16.17 presents the lateral development, rehabilitation, and production schedules in PFS years.

Figure 16.17 Mitchell Block Cave Mine Development and Production Schedules

Source: Golder, 2016

| 16.3.3 | Iron Cap Underground

The Iron Cap deposit extends approximately 1,200 m east-west (along strike), 700 m north-south, and 800 m in the vertical direction. It is understood that the deposit remains open at depth. Open pit mining methods were used to evaluate the mining potential of this deposit as part of an update to a PFS published in 2011. Similar to Mitchell, the location, dimensions, and dip of the mineralized material at Iron Cap indicates that it is a suitable candidate for block caving.

Iron Cap Underground Reserves

The Mineral Resource block model used for the study contains gold, copper, silver, and molybdenum grades as presented in Section 14.0, as well as NSR values described in Section 16.2.4. The model contains Measured, Indicated, and Inferred grades, but the Inferred grades were set to zero and are not included in this 2016 PFS. The Mineral Resources were evaluated using GEOVIA’s PCBC™ software to determine the Mineral Reserves for a block cave mine. A footprint at an elevation of 1,035 m produced the most value and resulted in 224.6 Mt of block cave Mineral Reserves as shown in Table 16.12.

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Table 16.12 Iron Cap Block Cave Reserves (Cdn$16/t NSR Shut-off)

Category | Tonnes (million) | Au (g/t) | Cu (%) | Ag (g/t) | Mo (ppm)
Probable1 | 224.6 | 0.49 | 0.20 | 3.6 | 13.0

| Notes: | 1Includes 25 Mt of mineralized dilution (the portion of Measured and Indicated material that is Cdn$0 < NSR < Cdn$16) and 20.2 Mt of non-mineralized dilution (material at zero NSR including the Inferred material set to zero grade).

The Mineral Reserves contain dilution that includes Mineral Resources that have grade but are sub-economic (less than drawpoint shut-off), Inferred Mineral Resources that are set to zero grade, and non-mineralized material that is zero grade rock. Dilution estimates for the Iron Cap block cave are 11% of sub-economic material (25 Mt) and 9% of zero grade dilution (20.2 Mt), for a total of 20% (45.2 Mt).

Iron Cap Underground Mine Design

The underground mine design is based on modelling using GEOVIA’s PCBC™ software Footprint Finder module. Footprint Finder modelling indicates that the optimum footprint for the Iron Cap deposit is at an elevation of 1,035 m and is approximately 600 m wide in the north-south direction and 400 m wide in the east-west direction. PCBC modelling indicates that the block cave could produce 15 Mt/a (average of 40,000 t/d), requiring development of 120 new drawpoints per year. The mine design requires approximately 87 km of drifts and raises, including a 25% design allowance to account for the excavation of infrastructure such as service bays, wash bays, fueling station, refuge stations, sumps and electrical substations, and to account for over break.

The Iron Cap mine design includes four main levels: preconditioning, undercutting, extraction, and conveying (Figure 16.18). The design also includes a return air drift located between the conveying and extraction levels. The floors of the extraction drifts and drawpoints are designed to be concreted, which will increase the speed and productivity of the LHD vehicles as well as reduce equipment maintenance.

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Figure 16.18 Section Looking East of the Iron Cap block Cave Mine Design

Source: Golder, 2016

Personnel, material, and supplies will access the Iron Cap mine through a drift driven from the Mitchell underground access ramp. Two fresh air portals and one exhaust portal are planned on the north slope of the Mitchell valley. These tunnels may act as an emergency egress. The fresh air tunnels will connect to surface and a perimeter drift will be constructed around the mine footprint to provide fresh air to the mine workings.

Excavated material will be hauled directly from the drawpoints to one of four gyratory crushers installed on the extraction level perimeter drift. The crushed material will be transported by one of two conveyor belts, which both feed a third conveyor that will transport the production material to a surge bin located above the Iron Cap MTT train tunnel.

Iron Cap Underground Equipment and Major Infrastructure

The proposed mobile diesel equipment is typical of that used in underground mines and will comprise equipment related to moving ore to the crushers (7 m3 LHDs and secondary rock breakers), development equipment (4.6 m3 LHDs and 18 m3 trucks), as well as the ANFO loaders and ground support machines. In addition, service equipment is included for construction, supervision, engineering and mine maintenance activities. At peak operation, Iron Cap will require a fleet of approximately 51 units of mobile underground equipment.

The Iron Cap mine workforce includes both staff and labour positions and the size varies according to the stage of the mine life with a peak quantity of 350 personnel in Year 38 (KSM production years).

The required airflow for the Iron Cap mine is 548 m3/s based upon the total diesel equipment used on each mining level, including a 20% design allowance for items such air loss around regulators, poorly installed or ripped ducting, and ventilating unused headings in active sections of the mine. It is estimated that the Iron Cap mine will require 9.1 MWh of electricity at peak operation.

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The hydrological characterization of the Iron Cap site indicates that a 200-year runoff event could lead to a maximum one-day inflow of approximately 292,000 m3 of water. The underground water management system at Iron Cap is currently designed to handle 4 m3/s. This caters for the estimated groundwater inflow and ice melt. The surface inflows will report to the drawpoints and will be managed in a similar manner to the groundwater inflows. To provide for drainage, the underground drifts will be graded so that water will naturally drain towards the MTT. Any flood water will be directed through the return airway drift and into the Mitchell NPWDA by a series of raises connecting the two tunnels. Pumps are not required to dewater Iron Cap.

Iron Cap Underground Mine Schedule

The mine development schedule is separated into three phases: an initial pre-production phase, which develops the primary access ramp and conveyor drifts; a second ore production phase that creates sufficient openings to start and ramp-up production from the cave; and a final phase once the mine has reached steady-state production and the development fleet is only required to create sufficient openings to maintain full production. The length of development required during the peak development period is approximately 10,000 m/a. The pre-production development period for Iron Cap is six years.

The mine production schedule was developed using GEOVIA’s PCBC™ software. It is assumed that sloughing of peripheral waste rock will occur into the crater and cover the upper surface of the material being drawn down. This was modeled in PCBC by assuming that an infinite supply of waste material is present on top of the mineralized material. As material is drawn from the drawpoints, the waste rock will mix with mineralized material as dilution with zero grade, and the combined material will report to the drawpoint.

The draw rates used in the PCBC modelling of Iron Cap are similar to those used at Mitchell, for similar reasons. During the early stages of mining, a draw rate of 200 mm/d was chosen as a maximum draw rate in the PCBC analysis. However, an average draw rate of only 110 mm/d is required to achieve production targets, so there are roughly twice as many drawpoints available as are required to meet production targets (the maximum draw rate modeled never exceeds 180 mm/d). Iron Cap is estimated to have a production ramp-up period of four years, steady state production at 15 Mt/a for 10 years, and then ramp-down production for another 9 years. Figure 16.19 presents the lateral development and production schedules in PFS years.

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Figure 16.19 Iron Cap Block Cave Mine Development and Production Schedules

Source: Golder, 2016

| 16.4 | Mine Production Schedule

The summarized production schedule results are shown in Table 16.13 and Figure 16.20, including both open pit and underground mining. The mine production plan starts in lower capital cost open pit areas using conventional large-scale equipment before transitioning into block cave underground bulk mining later in the mine life. Starting pits have been selected in higher grade lower strip ratio areas and cut-off grade strategy is used to enhance revenues for a minimum capital payback period. The cut-off strategy stockpiles lower grade open pit material early in the mine life. The Mitchell and Iron Cap underground block caves are brought into a development and production sequence to provide continuous mill feed during the LOM, while evening out the mine’s sustaining capital requirements. The open pit sequencing is then adjusted to augment the underground ore production to meet the full mill throughput requirements. The Mitchell underground block cave starts ore production in Year 23 and ramps up to full production by Year 29. Iron Cap block cave ore production starts in Year 32 and reaches full production by Year 36. At this point onward, all ore is supplied by the Mitchell and Iron Cap underground block caves (a total of 96,000 t/d). As the Mitchell block cave production starts to ramp down (starting in Year 46), the mill feed is further reduced to 62,000 t/d. In the final years of production (Year 49 and onwards), the underground ore is supplemented with material stockpiled from the open pits to maintain 62,000 t/d of mill feed.

Details of the production schedule can be found in the 2016 MMTS KSM PFS Engineering Report (MMTS, 2016).

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Table 16.13 Summarized Production Schedule – Open Pit and Underground

| Note: | 1Waste mined in the production schedule in Figure 16.20 includes re-handled waste and waste mined from borrow pit sources for construction purposes. 2The mill feed specified in Table 16.13 only includes ore from the Proven and Probable open pit and underground Mineral Reserves and does not include any Inferred Mineral Resources.

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Figure 16.20 KSM Mill Feed Production Schedule

Source: MMTS, 2016

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| 17.0 | Recovery Methods

| 17.1 | Introduction

The proposed KSM plant for the 2016 PFS will have an average process rate of 130,000 t/d. The process plant will receive ore from the Mitchell, Kerr, Sulphurets, and Iron Cap deposits. The planned mill life is approximately 53 years, excluding the production development stage and closure stage. The Mitchell deposit will be the dominant resource of mill feed for the process plant and will supply mill feed throughout the projected LOM. The ore from the Sulphurets deposit will be fed to the plant together with the ore from the Mitchell pit from Years 1 to 17, excluding Years 4, 5, 12 and 13, and with the ores from the other deposits during the last 4 years. The ore from the Kerr deposit, together with the ores from the other deposits, will be introduced to the plant during Years 24 to 34 and 53, while Iron Cap ore will be fed to the process plant from Year 32 to the end of mine life.

A combination of conventional flotation and cyanidation processes are proposed for the PFS. The processing flowsheet was developed based on the test results discussed in Section 13. In general, the minerlaization from the Mitchell, Sulphurets, Iron and Kerr deposits are amenable to the combined flotation and cyanidation process. HPGR is proposed for the PFS design because of its energy efficient comminution process. The process plant will consist of three separate facilities:

| ● | an ore primary crushing and handling facility at the mine site

| ● | an ore transportation system by trains through the MTT

| ● | a main process facility in the Treaty OPC area at the Plant Site, including secondary/tertiary crushing, primary grinding, flotation, regrinding, leaching/recovery, and concentrate dewatering.

These processes are shown in the simplified flowsheet in Figure 17.1 and are detailed in the following sections.

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Figure 17.1 Simplified Process Flowsheet

Source: Tetra Tech

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The ores from the Sulphurets and Kerr deposits and the upper zone of the Mitchell deposit will be extracted by open pit mining, while the ores from the Iron Cap deposit and the lower Mitchell deposit will be mined by underground block caving. The primary crushing facilities located at the Mitchell mine site will reduce the ROM particle size to approximately 80% passing 150 mm by gyratory crushers. Ores from the Sulphurets and Kerr deposits will be crushed at their respective sites, excluding the Sulphurets ore produced in Year 1, which will be hauled to and crushed at the Mitchell site. The ores from the lower Mitchell Zone and the Iron Cap deposit will be mined by block caving and be crushed in the underground mine. The crushed materials will then be transported by conveyors, or by a combination of conveyors and trains, to the MTT and loaded onto the MTT ore transport trains. The crushing circuit at the Mitchell surface site will include:

| ● | primary crushing by two 60 inch by 89 inch gyratory crushers

| ● | crushed ore transport conveyors.

The crushed ore will be transported by a train transport system through the MTT to the main process plant located at the Treaty OPC site, approximately 23 km northeast of the mine site. The main process plant will consist of the following process facilities:

| ● | secondary crushing by cone crushers in closed circuit with screens

| ● | tertiary crushing by HPGR in closed circuit with screens

| ● | primary grinding by ball mills

| ● | copper-gold/molybdenum bulk flotation with regrinding of bulk concentrate

| ● | copper-gold/molybdenum separation depending on molybdenum grade of mill feed

| ● | copper-gold concentrate and molybdenum concentrate dewatering

| ● | gold CIL cyanide leaching of scavenger cleaner tailing and pyrite rougher concentrate

| ● | gold recovery by carbon elution and production of gold doré

| ● | cyanide recovery, and then cyanide destruction of washed CIL residue prior to disposal of the residue in the lined pond within the TMF.

The TMF, located southeast of the main process plant, is designed to store flotation tailing and CIL tailing, which will be stored in the lined tailing pond of the TMF.

The mill feed produced from the Mitchell crushing facility or from the block caving sites will be transported via the MTT train system to the coarse ore stockpile at the Treaty OPC site. The stockpile will be located at the exit portal of the MTT tunnel and will have a live capacity of 60,000 t. The coarse ore will be reclaimed and be further crushed by five cone crushers (four in operation and one on standby) and then four HPGRs in closed circuit with vibrating screens.

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The screen undersized material from the HPGR circuit will be fed to four ball mills in a closed circuit with hydrocyclones. Ore solids will be reduced to a particle size of 80% passing 125 µm to 150 µm.

The products from the primary grinding circuits will be fed into copper-gold/molybdenum rougher/scavenger flotation circuits, consisting of two operation circuits in parallel. The copper rougher flotation concentrates from the flotation circuits will be reground to a particle size of 80% passing, approximately 20 µm in the tower mills.

The reground rougher concentrate will then be upgraded in a cleaner flotation circuit with three stages of copper cleaner flotation, producing a copper-gold or copper-gold/molybdenum concentrate with an average grade of 25% Cu. Depending on the molybdenum content in the copper-gold/molybdenum concentrate, the bulk concentrate may be treated by flotation to produce a molybdenum concentrate and a copper-gold concentrate. The molybdenum concentrate will be leached using the Brenda Mines procedure to reduce copper and lead contents.

The final copper concentrate(s) will be dewatered by a combination of thickening and pressure filtration to approximately 9% moisture before being transported to the Stewart port site for ship loading and delivery to copper smelters, while the molybdenum concentrate will be further dried prior to being shipped in bags to the port at Prince Rupert for delivery to molybdenum smelters.

The copper-gold/molybdenum rougher scavenger flotation tailing will be subjected to further flotation, producing a gold-bearing pyrite concentrate. The final pyrite flotation tailing will be sent to the TMF for storage. The pyrite concentrate will be reground in tower mills to a particle size of 80% passing approximately 20 µm.

The reground gold-pyrite concentrate and the first copper cleaner tailing from the copper-gold/molybdenum cleaner flotation circuit will be separately leached in a CIL cyanidation plant to recover the contained gold. The sulphide pulp will be pre-oxidized by aeration prior to cyanidation. Dissolved gold will be adsorbed onto activated carbon in the CIL circuit.

The loaded carbon from the two streams will be combined and gold will be stripped from the carbon by a conventional Zadra pressure stripping process; the gold in the pregnant solution will be recovered in the subsequent electrowinning process. The barren solution from the elution circuit will be circulated back to the leach circuit. The gold sludge produced from the electrowinning circuit will be smelted using a conventional pyrometallurgical technique to produce gold-silver doré bullion.

The residues from the leach circuit will be pumped to a conventional counter-current decantation (CCD) washing circuit. The solution from the circuit will be sent to a cyanide recovery circuit using a combination of a SART process and an AVR process. The AVR process will recover the free cyanide from the solution by acidifying and stripping the solution and then absorbing the stripped hydrogen cyanide gas by a sodium hydroxide solution to recover the cyanide for reuse.

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The washed residues will be treated by an sulphur dioxide/air cyanide destruction process to destroy the residual weak acid dissociable (WAD) cyanide. The leach residues will then be further treated by carbon adsorption to remove dissolved copper. The copper will be stripped off from the activated carbon and precipitated by sodium sulphide as copper sulphide which will be blended into copper concentrates for sale.

The treated residues will then be transported by pipeline to the lined CIL pond of the TMF. The sulphide leach residues will be stored under water at all times to prevent the oxidation of sulphides.

The Treaty process plant layout and the primary grinding and flotation facilities are depicted in Figure 17.2.

| 17.2 | Major Process Design Criteria

The concentrator is designed to process an average of 130,000 t/d. The major criteria used in the design are shown in Table 17.1.

Table 17.1 Major Design Criteria

Criteria | Unit | Value
Average Daily Process Rate | t/d | 130,000
Operating Year | d | 365
Primary/Secondary Crushing
Availability – Primary Crushing | % | 70
Availability – Secondary Crushing | % | 85
Primary Crushing Product Particle Size, P80 | mm | 150
Secondary Crushing Product Particle Size, P80 | mm | 45
HPGR/Grind/Flotation/Leach
Availability | % | 94
Milling and Flotation Process Rate | t/h | 5,762
Mill Feed Size, P80 | mm | 2.0
Primary Grind Size, P80 | µm | 125-150
Bond Ball Mill Work Index - Design | kWh/t | 16
Bond Abrasion Index | g | 0.293
Concentrate Regrind Size, 80% Passing
Cu/Au Rougher/Scavenger Concentrate | µm | 20
Au-Pyrite Concentrate | µm | 20
Gold-bearing Materials Leach Method | - | CIL
Feed Mass to CIL Circuit | t/d | 14,600

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Figure 17.2 Treaty Process Plant Layout

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| 17.3 | Process Plant Description

| 17.3.1 | Primary Crushing

There will be five primary crushing sites throughout the plant operation life at the Mitchell, Kerr, and Sulphurets sites, as well as in the underground block caving mines. The proposed mill feed rates from these deposits are shown in Section 16.4.

At the Mitchell OPC site, primary crushing will mainly consist of two 60 inch by 89 inch gyratory crushers, two apron feeders, and one train load surge bin feed conveyor. The ROM material feeding to the gyratory crushers will be from the Mitchell pit and the first year Sulphurets pit, and will be approximately 80% passing 1,200 mm. The oversize materials will be fragmented by a rock breaker. The gyratory crushers will reduce the ROM material to a particle size of 80% passing 150 mm or less. The products from each gyratory crusher will be fed to one 1.83 m wide by 37 m long conveyor via one 2.13 m wide by 10 m long apron feeder. The crushed ore from the two conveyors will be fed to a 2.13 m wide by 450 m long train load surge bin feed conveyor, which will be located inside of the Mitchell surge bin feed conveyor tunnel. The surge bin is designed to have a live capacity of 30,000 t (two pockets, each 15,000 t). The ore from the Mitchell open pit will feed to the mill during Years 1 to 22, 34, 35, 48, and 49.

For the underground block cave operation, the ore from the lower Mitchell zone will be mined by block caving and crushed on site to 80% passing 150 mm or finer. The crushed ore will be conveyed to the 30,000-t surge bin where the crushed ore will be blended with the materials from the Mitchell and Sulphurets pits and then loaded from the surge bin into the train cars and transported to the end of the MTT at the Treaty site. The ore will be fed to the mill during Years 19 to 49.

The Sulphurets ore will supplement the mill feed from Years 1 to 17, excluding Years 4, 5, 12 and 13, and with the ores from the other deposits during the last four years. The ROM ore from the Sulphurets pit will be trucked to the Mitchell site and crushed at the Mitchell crushing facility in Year 1. Starting from Year 2, the ore produced will be crushed by a 60 inch by 89 inch gyratory crusher at the Sulphurets mine site. The crushed ore will be conveyed to the Mitchell site via the 3.0 km SMCT to a 10,000-t Sulphurets/Kerr coarse ore stockpile. The stockpiled ore will then be reclaimed and trucked to the Mitchell crusher dumping pockets, where the ore will pass through the crushers and be sent to the 30,000-t surge bin together with the crushed Mitchell ore.

The ore from the Kerr deposit, together with the ores from the other deposits, will be introduced to the plant during Years 24 to 34 and 53, the ROM ore and waste rock from the Kerr pit will be crushed by two 60 inch by 89 inch gyratory crushers at the Kerr mine site. The crushed ore will then be conveyed to Mitchell through a 2,480 m cross valley rope conveyor to the Sulphurets site, followed by the 3.0 km overland conveyor through the SMCT to the Sulphurets/Kerr coarse ore stockpile at the Mitchell site. The ore from the stockpile will be trucked to the Mitchell crushing facility or to the 10-Mt surge stockpile for later reclaiming and delivery to the Mitchell crushing facility. Similarly, the reclaimed ore will be trucked to the Mitchell crusher dumping pockets, where the ore will pass through the crushers and be sent to the 30,000-t surge bin together with the crushed Mitchell ore.

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Waste rock from the Kerr mine will be conveyed from the Kerr to the Sulphurets pit via the rope conveyor. The waste rock will then be backfilled into the mined Sulphurets pit for storage.

The Iron Cap ore will be mined by block caving and crushed on site to 80% passing 150 mm or finer. The crushed ore will be conveyed to a surge bin located at the end of the Iron Cap and MTT connection tunnel. The ore will supplement the mill feed from Year 32 to the end of mine life.

| 17.3.2 | Coarse Ore Transport From Mitchell Site to Treaty Site

The crushed ore will be reclaimed by two automatic train loading systems from the two coarse ore surge bins at the Mitchell site and transported to the plant site by a train transport system through the MTT. The ore from the Iron Cap site will be loaded into the train cars at the Iron Cap underground site from the Iron Cap surge bin and transported to the Treaty site via an Iron Cap and MTT connection tunnel and then through the MTT. Loading chutes under the ore surge bins will feed ores into awaiting trains that will transport the ores to an unloading station at the Treaty end of the MTT. Because the loading and unloading systems, including surge bins are located underground, the arrangement would mitigate potential freezing issues. The train cars will dump ore into a live underground unloading bin. Two apron feeders will unload the bin onto a conveyor to transport the ore to the top of the Treaty coarse ore stockpile. Loading chutes will be controlled remotely, and unloading chutes will operate autonomously. No onboard operators will be required within the tunnels during train system operation.

Each train will consist of one, 140-t electric locomotive and sixteen 42-m3 belly dump ore cars that will have the capacity to deliver 800 t/h from Mitchell to Treaty based on 90-minute cycle times. On average, eight trains will deliver approximately 130,000 t/d of ore to meet the process plant requirements. An additional four trains will be on standby to provide for mechanical availability or to handle an increase in plant feed of up to 10,000 t/h. The transport system is detailed in Section 18.4. Dust collecting systems will be installed at the loading and unloading points to collect fugitive dust.

| 17.3.3 | Coarse Material Handling

The crushed ore from the trains will be continuously and automatically unloaded from the bottom discharge ore cars into the bin underneath, with each car taking an average of nine seconds to unload. The train will be driven via traction drives across the unloading station at a maximum speed of 2.5 km/h.

At the bottom of the surge bin, the ore will be reclaimed by two apron feeders and then onto a conveyor belt that will transport the ore to the surface and feed the Treaty coarse ore stockpile with a live capacity of 60,000 t at the Treaty OPC site. Apart from the conveyor tunnel, a vertical escape tunnel that joins the unloading station and the surface will be constructed for emergency egress.

Seabridge Gold Inc. | 17-8 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

The ore will be reclaimed from the stockpile by six, 1.8 m wide by 8.5 m long apron feeders and conveyed in two lines to the secondary crushing circuit. A dust collecting system will be installed at each of the transfer points to collect fugitive dust. The reclaim tunnel will be heated to prevent potential freezing during operation in winter.

| 17.3.4 | Secondary Crushing

The reclaimed coarse ore will be conveyed to the secondary crushing facility and fed to four vibrating screens. Each screen oversize will feed a secondary cone crusher. Each secondary crusher is in closed circuit with a screen. The cone crusher product will return to the screen feed conveyor. A spare cone crusher is provided in the circuit for when any of the other four cone crushers require maintenance.

Screen undersize product that is finer than 50 mm will be delivered by conveying to an enclosed surge stockpile with a 60,000 t live capacity. The circuit will consist of the following key equipment:

| ● | five cone crushers, each with an approximately 2.4 m diameter mantle and driven by a 750-kW motor or equivalent

| ● | five 3.7 m wide by 7.3 m long double deck vibrating screens (one unit on standby).

| 17.3.5 | Tertiary Crushing Material Conveyance/Storage

The crushed ore from secondary crushing will be reclaimed from the 60,000 t stockpile by six 1.5 m by 7.6 m reclaim apron feeders and fed into two 1.37 m-wide HPGR feed conveyors. These conveyors will deliver the ore to two tertiary crusher HPGR feed surge bins, each with a live capacity of 400 t. Similarly, the stockpile reclaim tunnel will be heated to prevent potential freezing during winter operation.

| 17.3.6 | Tertiary Crushing

The reclaimed ore will be further crushed by four HPGR crushers. Four belt feeders will withdraw the reclaimed ore from the two HPGR feed surge bins and feed each of the four HPGR crushers separately. Each HPGR crusher is in a closed circuit with a 4.0 m wide by 8.0 m long double deck vibrating screen. Discharge from the HPGR crushers will be wet-screened at a cut size of 6 mm. The screen oversize will return to the feed conveyor of the HPGR feed bin while the screen undersize will leave the crushing circuit and report to the ball mill grinding circuits. The four HPGR crushing lines will have a total process capacity of 5,762 t/h. The key equipment is as follows:

| ● | four HPGR crushers, each equipped with two 2,900 kW motors

| ● | four 4.0 m wide by 8.0 m long vibrating screens

| ● | four 1.5 m wide by 10.0 m long belt feeders.

Seabridge Gold Inc. | 17-9 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 17.3.7 | Primary Grinding

The grinding circuit will employ conventional ball mills to grind the HPGR product to a particle size of 80% passing 125 to 150 µm. All the primary grinding circuits are designed to have a nominal processing rate of 5,762 t/h.

The primary grinding circuit will include four milling circuits, which are made up of the following equipment:

| ● | four 7.6 m diameter by 11.9 m long (25 ft by 39 ft) ball mills, each mill driven by two 7.0 MW synchronous motors

| ● | six 700 mm by 650 mm centrifugal slurry pumps (4 in operation and 2 on standby), each equipped with a 1,650 kW variable speed drive

| ● | four hydrocyclone clusters, each with twelve 710 mm diameter hydrocyclones.

Each ball mill will be in a closed circuit with a cluster of twelve 710 mm diameter hydrocyclones. The hydrocyclone underflow will gravity-flow to the ball mill feed chute, while the overflow of each hydrocyclone cluster with a solid density of 37% weight/weight (w/w) will gravity-flow to one of the four copper-gold-molybdenum rougher flotation trains.

Lime will be added to each mill as required. Flotation collectors will be added to the hydrocyclone feed sumps or to the hydrocyclone overflow collecting sumps.

| 17.3.8 | Copper, Gold and Molybdenum Flotation

Copper-Gold/Molybdenum Bulk Rougher/Scavenger Flotation

There will be two copper-gold-molybdenum bulk rougher flotation circuits. The overflow from two of the four hydrocyclone clusters from the primary grinding circuits will separately feed the two flotation trains, each consisting of six 300 m3 flotation cells. The flotation reagents used will include lime, A208, 3418A, fuel oil, and MIBC. A bulk copper-gold/molybdenum rougher flotation concentrate, approximately 6% of the flotation feed by weight, will be reground. The flotation tailing will be sent to the pyrite flotation circuit.

Copper-Gold/Molybdenum Bulk Concentrate Regrinding

The copper-gold/molybdenum bulk concentrate will be reground to a particle size of 80% passing 20 µm in a regrind circuit consisting of three tower mills, each with an installed power of 2,240 kW, and a 250 mm diameter hydrocyclone cluster. The overflow of the hydrocyclones will gravity-flow to the bulk copper-gold/molybdenum cleaner circuit, while the underflow of the hydrocyclones will return to the regrinding mills by gravity flow.

Copper-Gold/Molybdenum Bulk Concentrate Cleaner Flotation

The hydrocyclone overflow will be cleaned in three stages. In the first stage of cleaner flotation, six 100 m3 tank cells will be used; for the second and third stages, three 50 m3 tank cells and two 50 m3 tanks will be used separately. First cleaner flotation tailing will be further floated in two cleaner scavenger flotation cells, each with 100 m3 capacity. The concentrate product from the cleaner scavenger flotation will be sent to the first cleaner cells and the tailing will report to the gold leaching circuit. The tailing from the second and third cleaner flotation stages will be returned to the head of the preceding cleaner flotation circuit. Final copper-gold/molybdenum bulk concentrate will be sent to copper-gold/molybdenum bulk concentrate thickener.

The same reagents used in the rougher flotation circuit will be employed in the cleaner flotation circuits.

Seabridge Gold Inc. | 17-10 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Copper-Gold and Molybdenum Separation

Depending on molybdenum content, the final copper-gold/molybdenum concentrate may be further processed to produce a copper-gold concentrate and a molybdenum concentrate. The separation will employ a conventional process, which will include copper suppression by sodium sulphide and four stages of molybdenum cleaner flotation and regrinding. The circuit will include the following key equipment:

| ● | one 15 m diameter high rate thickener

| ● | six 30 m3 conventional mechanical flotation cells

| ● | one 1.5 m diameter by 4.5 m high column cell

| ● | one 1.1 m diameter by 4 m high column cell

| ● | two 1.0 m diameter by 4 m high column cells

| ● | one nitrogen gas generator

| ● | one regrinding stirred mill.

The copper-gold/molybdenum bulk concentrate will be thickened prior to the copper-gold/molybdenum separation. The thickener underflow will be diluted and conditioned with sodium sulphide and gravity flow into the molybdenum rougher flotation cells. The rougher flotation tailing will be scavenged by flotation and the scavenger concentrate will return to the rougher flotation head while the tailing will be the final copper-gold concentrate reporting to the copper-gold concentrate dewatering circuit.

The resulting rougher molybdenum concentrate will be classified by a hydrocyclone. The hydrocyclone underflow will be reground by a stirred mill and joined with the hydrocyclone overflow reporting to the molybdenum cleaner flotation circuit. Four stages of cleaner flotation were designed to upgrade the molybdenum rougher flotation concentrate to marketable grade. The tailing of each cleaner flotation will be returned to the head of the preceding molybdenum cleaner flotation circuit while the first cleaner tailing will be sent to the molybdenum rougher flotation conditioning tank. To reduce sodium hydrosulphide consumption, the molybdenum flotation cells will be aerated by nitrogen gas, which will be generated on site by a nitrogen generator.

The final cleaner flotation concentrate will be leached to reduce the copper content, if it is higher than 0.2%. The leached product will be dewatered in a molybdenum concentrate dewatering facility.

Seabridge Gold Inc. | 17-11 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 17.3.9 | Concentrate Dewatering

The upgraded copper-gold concentrate will be thickened in a 15 m diameter high rate thickener. The thickener underflow will be directed to the copper-gold concentrate pressure filter to further reduce the water content to 9% moisture. The copper-gold concentrate will be stockpiled on site and then transported by trucks to a port facilities at Stewart, where the concentrate will be stored and loaded into ships for ocean transport to overseas smelters.

The average copper concentrate produced is estimated to be approximately 940 t/d or 344,000 t/a.

The molybdenum concentrate will be dewatered using a similar process to the copper-gold concentrate. The filtered concentrate will be further dewatered by a dryer to 5% moisture content, before being bagged and transported to the processors. The key equipment used in the dewatering processes will include:

| ● | copper-gold concentrate dewatering:

| - | one 15 m diameter high rate thickener

| - | one 8 m diameter by 7 m high concentrate stock tank

| - | two 160 m2 pressure filters.

| ● | molybdenum concentrate dewatering:

| - | one 2 m diameter high rate thickener

| - | one molybdenum concentrate leaching system

| - | one 4 m2 pressure filter

| - | one 2.5 t/h dryer.

| 17.3.10 | Gold Recovery From Gold-bearing Pyrite Products

Gold-bearing Pyrite Flotation

The tailing of the copper-gold/molybdenum rougher flotation circuits will be further floated in a pyrite flotation circuit. The pyrite rougher flotation will consist of two parallel lines, each line with six 300 m3 pyrite rougher flotation cells.

Tailing from the pyrite rougher flotation will gravity-flow, or be pumped to the TMF located southeast of the main process plant.

Gold-bearing Pyrite Concentrate Regrinding

The pyrite concentrate will be reground to a particle size of 80% passing approximately 20 µm in three 2,240 kW tower mills. A hydrocyclone cluster consisting of twenty-six 250 mm diameter hydrocyclones will be incorporated with the mills in a closed circuit. The hydrocyclone overflow will report to the gold leach circuit or the copper-pyrite separation circuit.

Depending on copper content, the reground materials may be subjected to a flotation process to separate copper minerals from the other minerals. The copper concentrate will be sent to the copper-gold/molybdenum cleaner flotation circuit, while the flotation tailing will report to the gold leach circuit.

Seabridge Gold Inc. | 17-12 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Gold Leach

The reground gold-bearing pyrite product and the first cleaner scavenger tailing from the copper-gold/molybdenum bulk flotation circuit will be separately thickened to a solids density of 65% in two 35 m diameter high rate thickeners.

The underflow of each thickener will be pumped to two separate cyanide leaching lines. Each line will consist of two pre-treatment tanks and five cyanide leaching tanks. In the pre-treatment tanks, the thickener underflow will be diluted with barren solution to approximately 45% w/w and aerated. Lime will be added to increase the slurry pH to approximately 11.

The pre-treated slurry will be leached by sodium cyanide to recover gold in a conventional CIL circuit. The leach circuit will consist of five agitated tanks, which are 15 m diameter by 15 m high. The tanks will be equipped with in-tank carbon transferring pumps and screens to advance the loaded carbon to the preceding leach tank.

The loaded carbon leaving the first CIL tanks of the two leaching lines will be transferred to the carbon stripping circuit, while the leach residue will be blended and sent to subsequent processes including residue washing, cyanide recovery, and cyanide destruction circuits.

The key equipment in the leach circuit will include:

| ● | two 35 m high rate thickeners

| ● | four 9 m diameter by 10 m high aeration tanks

| ● | ten 15 m diameter by 15 m high CIL leach tanks equipped with in-tank carbon transferring pumps and screens

| ● | one 3 m wide by 4 m long carbon safety screen.

Compressed air will be provided for the leaching process from four dedicated oil-free air compressors.

Carbon Stripping and Reactivation

The loaded carbon will be treated by acid washing and the Zadra pressure stripping process for gold desorption.

The loaded carbon will be acid washed prior to being transferred to two elution vessels. The stripping process will include the circulation of the barren solution through a heat recovery heat exchanger and a solution heater. The heated solution will then flow up through the bed of the loaded carbon and overflow near the top of the stripping vessels. The pregnant solution will flow through a back pressure control valve and then be cooled by exchanging heat with the barren solution prior to reporting to the pregnant solution holding tank for subsequent gold recovery by electrowinning. The barren solution from the electrowinning circuit will then return to the barren solution tank for recycling.

Seabridge Gold Inc. | 17-13 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

The stripping process will include barren and pregnant solution tanks, two 3 t acid wash vessels, two 3 t stripping vessels, four heat exchangers, and two solution heaters and associated pumps.

Prior to reactivation, the stripped carbon will be screened and dewatered. The reactivation will be carried out in an electrically heated rotary kiln at a temperature of 700°C. The activated carbon will be circulated back to the CIL circuit after abrasion treatment and screen washing.

The carbon reactivation process will include a reactivation kiln, a carbon quench tank, and a carbon abrasion tank equipped with an attrition agitator, a reactivated carbon sizing screen, a carbon storage bin, and fine carbon handling associated equipment.

Gold Electrowinning and Refining

The pregnant solution from the elution system will be pumped from the pregnant solution stock tank through electrowinning cells where the gold and silver will be deposited on stainless steel cathodes. The depleted solution will be subsequently reheated and returned to the stripping vessel. The electrowinning circuit will have a capacity to process 80 kg/d of gold-silver doré bullion and will include two 3.5 m3 electrowinning cells, direct current rectifiers, cathodes, anodes, and a pressure filter.

Periodically, the stainless steel cathodes will need to be cleaned to remove precious metal residues by pressure washing. The cell mud will fall into the bottom of the electrowinning cells and pumped through a pressure filter for dewatering on a batch basis. The filter cake will be transferred to the gold room for drying and smelting after it is mixed with melting flux. A 125 kW induction furnace will be used for gold-silver refining. The area will be monitored by a security surveillance system.

| 17.3.11 | Treatment of Leach Residues

Leach Residue Washing

The residues from the CIL circuit will be pumped to a two stage conventional CCD washing circuit. The CCD circuit will consist of two 40 m diameter high rate thickeners. The thickener overflow from the first stage washing will be pumped to the cyanide recovery system. The underflow (washed residues) of the second thickener will be sent to the cyanide destruction circuit prior to being pumped to the TMF.

Cyanide Recovery

The overflow of the first leach residues washing thickener will be sent to a cyanide recovery circuit where the copper will be removed and the cyanide will be recovered from the solution by a SART/AVR process.

The SART/AVR cyanide recovery process will be carried out in a negative pressure system generated by a vacuum system.

Seabridge Gold Inc. | 17-14 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

The CCD overflow will be acidified by sulphuric acid. Sodium hydrosulphide will be added to precipitate the heavy metals in the solution, especially the copper. The precipitates will be blended with the copper-gold concentrate for sale. The solution will then be pumped to two volatilization towers in series. The solution together with pressurized air will be sprayed in the towers to provide a high liquid surface area to promote volatilization.

The gas phase will be directed through an absorption tank, in which a caustic solution is circulated counter-current to the gas to absorb hydrogen cyanide. The regenerated cyanide solution will be returned to the leach circuit.

The cyanide-depleted solution from the volatilization tower will be settled in a 10 m diameter clarifier. The metal species will precipitate in the clarifier, while the clarified solution will be circulated to the leach residues washing circuit and the leach circuit after the solution is treated with lime to a pH above 9.5. The precipitates will be blended with the copper-gold concentrate for sale.

Cyanide Destruction

The remaining cyanide in the washed leach residues from the second washing thickener will be decomposed by a sulphur dioxide (SO2)/air oxidation cyanide destruction process. Sodium metabisulphite will be used as the sulphur dioxide source. The equipment will include a 6 m diameter by 6 m high pre-aeration agitation tank, three 11 m diameter by 12 m high sulphur dioxide oxidation tanks, and a wet alkaline scrubbing system. Compressed air will be provided for the oxidation process. The treated residues will be sent to the copper removal treatment circuit.

Copper Removal

A copper removal circuit is proposed to remove the dissolved copper from the treated residue slurry if the copper level in the slurry from the sulphur dioxide/air cyanide destruction circuit is higher than the requirement. Activated carbon will be added to the residue slurry after the slurry is treated by cyanide destruction. The copper removal treatment will be carried in two stages in two reactors. The loaded carbon will be removed from the first stage of the copper removal reactor, while the fresh carbon will be added into the section stage of the copper removal reactor. The copper loaded carbon will be stripped by acid washing and the copper in the washing solution will be precipitated by sodium sulphide. The precipitate produced will be blended with the copper-gold concentrate for sale. The treated residues will be sent to the lined CIL residue storage pond in the TMF.

| 17.3.12 | Tailing Management

The flotation tailing and the treated CIL residues will separately gravity-flow or be pumped to the TMF located southeast of the main process plant. The flotation tailing and CIL residue will be stored in separate areas within the TMF.

Seabridge Gold Inc. | 17-15 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

The CIL residue will be deposited in a lined CIL residue storage pond. The residue will be covered with the supernatant to prevent oxidation of the sulphide minerals. The residue will be eventually covered by the flotation tailing. The supernatant from the CIL residue pond will be reclaimed by pumping to the CIL circuit for reuse. The excess water will be sent to a hydrogen peroxide (H2O2) water treatment plant to further remove impurities before it is sent to the north or south tailing ponds.

There will be two flotation tailing pipelines directing the flotation tailing to the TMF. The flotation tailing from one of the tailing pipelines will be classified to produce coarse tailing sands by two stages of hydrocyclone classification. The coarse fraction will be used to construct the tailing dam and the fines will directly report to the TMF together with the tailing from the other line. The supernatant from the tailing impoundment area will be reclaimed by a reclaim water barge and sent to the process water tank by two stages of pumping. The reclaimed water will be used as process water for flotation circuits.

One energy recovery system will be installed on one of the rougher flotation tailing lines, which will deliver the tailing to the north dam, to generate electricity.

A separate barge equipped with reclaim water pumps will be installed in the flotation tailing storage pond to reclaim water from the TMF supernatant for the tailing classification operations (to provide dilution water for hydrocycloning) and for the excess water discharge via the Treaty Creek diffuser. Discharge will occur during a five month window, beginning during spring runoff when the creek flows are highest. A floating skimmer will be installed. If required, flocculant will be added from the floating skimmer to improve the settlement of any suspended solids before the excess water is discharged.

| 17.3.13 | Reagents Handling

The reagents used in the process will include:

| ● | Flotation: PAX, 3418A, A208, fuel oil, MIBC, lime (CaO), NaHS, and sodium silicate (Na2SiO3)

| ● | CIL and Gold Recovery: lime, sodium cyanide (NaCN), activated carbon, sodium hydroxide (NaOH), hydrochloric acid (HCl) and flux

| ● | Cyanide Recovery and Destruction Reagents: metabisulphite (MBS), copper sulphate (CuSO4), sulphuric acid (H2SO4), lime, NaOH, activated carbon, sodium hydrosulphide (NaHS)

| ● | Others: flocculant, antiscalant, H2O2.

All the reagents will be prepared in a separate reagent preparation and storage facility in a containment area. The reagent storage tanks will be equipped with level indicators and instrumentation to ensure that spills do not occur during operation. Appropriate ventilation and fire and safety protection will be provided at the facility.

The liquid reagents (including fuel oil, A208, 3418A, MIBC, HCl, H2SO4, H2O2 and antiscalant) will be added in the undiluted form to various process circuits via individual metering pumps.

Seabridge Gold Inc. | 17-16 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

All the solid type reagents (including PAX, NaHS, Na2SiO3 if required, NaOH, NaCN, CuSO4, and MBS) will be mixed with fresh water to 10% to 25% solution strength in the respective mixing tank, and stored in separate holding tanks before being added to various addition points by metering pumps.

Lime will be slaked, diluted into 15% solid milk of lime, and then distributed to various addition points through a closed and pressurized loop.

Flocculant will be dissolved, diluted to less than 0.5% strength, and then added to various thickener feed wells by metering pumps.

Flux will be added directly in solid form.

| 17.3.14 | Water Supply

Three separate water supply systems will be provided to support the operation; a fresh water system, a process water system for grinding/flotation circuits and a process water system for CIL/gold recovery circuits.

On average, based on the preliminary water requirement and water balance estimates, approximately 39,400 m3/day of process make-up water and 1,300 m3/day of fresh water will be required for the processing plant.

Fresh Water Supply System

Fresh and potable water will be supplied to two 12 m diameter by 9 m high storage tanks from nearby wells and local drainage runoff areas. One tank will be located at the Plant Site and the other at mine site. Fresh water will be used primarily for the following:

| ● | fire water for emergency use

| ● | cooling water for mill motors and mill lubrication systems

| ● | potable water supply

| ● | reagent preparation.

By design, the fresh water tanks will be full at all times and will provide at least 2 h of firewater in an emergency. The minimum fresh water requirement for process mill cooling and reagent preparation is, on average, estimated to be approximately 250 m3/h.

The potable water from the fresh water source will be treated (by chlorination and filtration) and stored in a covered tank prior to delivery to various service points.

Process Water Supply System

Two process water systems will supply the process water for the process plant. The water for each circuit will be from different sources, as follows:

| ● | water for Grinding/Flotation Circuits: reclaimed water from the flotation tailing pond, copper-gold/molybdenum concentrate thickener overflow and the CIL feed thickener overflow, as well as fresh water. The dominant process water will be the supernatant fluid from the flotation tailing impoundment area

Seabridge Gold Inc. | 17-17 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| ● | water for CIL Leaching/Gold Recovery Circuits: reclaimed water from the CIL storage pond, barren solution and fresh water. As required, the water reclaimed from the flotation tailing pond may also be used in these circuits.

The water reclaimed from the flotation tailing impoundment area will be sent to a 25 m diameter by 15 m high process water surge tank by two stages of pumping systems, while the bulk concentrate thickener overflow will be directed to the primary grinding circuits. The process water tank will be located approximately 25 m higher than the process plant base elevation. The water will flow to the various service points by gravity. A booster pump station is provided at the Plant Site to pump water to the various distribution points where high pressure water is required.

The water from the CIL residue storage pond will be pumped to an 8 m diameter by 8 m high process water surge tank located at the Plant Site. The water will service the CIL leach/gold recovery circuits. Any excess water from the CIL residue storage pond will be treated at the H2O2 WTP located at the Plant Site. The treated water will be sent to the north or south tailing ponds.

The overall site water management is detailed in Section 18.2.

| 17.3.15 | Air Supply

Plant air service systems will supply air to the following areas:

| ● | flotation circuits – low pressure air for flotation cells by air blowers

| ● | leach circuits – high pressure air by dedicated air compressors

| ● | cyanide recovery and destruction circuits – high pressure air by dedicated air compressors

| ● | filtration circuit – high pressure air for filter pressing and drying of concentrate by dedicated air compressors

| ● | crushing circuit – high pressure air for the dust suppression (fogging) system and other services by an air compressor

| ● | plant service air – high pressure air for various services by two dedicated air compressors

| ● | instrumentation – instrument air at mine siteand Plant Site will come from the local air compressors and will be dried and stored in dedicated air receivers.

| 17.3.16 | Assay and Metallurgical Laboratory

The assay laboratory will be equipped with necessary analytical instruments to provide routine assays for the mine, process, and environmental departments.

The metallurgical laboratory, with laboratory equipment and instruments, will undertake all necessary test work to monitor metallurgical performance and to improve the plant production and metallurgical results.

Seabridge Gold Inc. | 17-18 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 17.3.17 | Process Control and Instrumentation

The plant control system will consist of a Distributed Control System (DCS) with PC-based Operator Interface Stations (OIS) located in the following three control rooms:

| ● | Mitchell site primary crusher control room

| ● | MTT train transport control room

| ● | Treaty plant site control room.

The plant control rooms will be staffed by trained personnel 24 h/d.

A crushing control room at the Sulphurets pit will be added in Year 2. The Sulphurets pit crushing control room will be relocated to the Kerr pit crushing plant in Year 20.

In addition to the plant control system, a closed-circuit television (CCTV) system will be installed at various locations throughout the plant including the crushing facility, the stockpile conveyor discharge point, the slurry pumping tunnel, the tailing facility, the concentrate handling building, and the gold recovery facilities. CCTVs in these areas will be monitored from the local control room and the central control room.

An automated train dispatching system will be utilized to achieve a safe and efficient flow of trains through the system, with no on-board operators. The system employing full radio-based train spacing and speed supervision on the whole railway system will be supervised from a control room located in the train maintenance shop. The train control system will operate using a wireless communications system (Wi-Fi) that must be in place for the entire track. While wireless communications are the current state of the art technology for train control communications, it is recognized that more efficient and reliable communications may be developed in the future.

Process control will be enhanced with the installation of an automatic sampling system. The system will collect samples from various streams for on-line analysis and the daily metallurgical balance.

To protect operating staff, cyanide monitoring/alarm systems will be installed in the cyanide leaching area as well as at the cyanide recovery area and destruction area. A sulphur dioxide monitor/alarm system will monitor the cyanide destruction area as well.

Seabridge Gold Inc. | 17-19 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 17.4 | Yearly Production Projection

In general, the mineral processing is designed to use conventional flowsheet and mature technologies for the process plant. The flowsheet proposed is relatively simple and mining will start with conventional open pit operation with an exposed ore body on the surface. HPGRs will be used for comminution circuit, which is expected to have fewer potential rock hardness issues. Cerro Verde mine used a similar comminution and flotation plant initially for its approximately 120,000 t/d plant, which has since been successfully expanded to 360,000 t/d. They leveraged experience from the existing operation and ramped up to approximately 95% of the name plate rate in approximately 6 months. Also notable is that larger cone crushers and HPGRs greater than those which KSM will employ have been in operation.

It is estimated that the plant may take approximately twelve months to reach design capacity after the plant is wet commissioned.

According to the metallurgical projections described in Section 13.7 and the current mine schedule, metal recovery and concentrate grades for the plant operation life are projected on a yearly basis, as indicated in Table 17.2.

As shown by the test results, it is anticipated that on average the impurity contents in the copper concentrates would be below the penalty limits as outlined for most of the smelters, although in short periods the impurity content may slightly exceed the penalty limits as outlined for some of the smelters. The projected copper concentrate quality is shown in Table 17.3.

In general, the molybdenum concentrate separated from copper and molybdenum bulk concentrate will be leached on site to remove copper, lead and other impurities. The anticipated molybdenum content is approximately 50%. The main impurities such as copper and lead are estimated to be lower than 0.2% and 0.3%, respectively.

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Table 17.2 Projected Metallurgical Performance

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Table 17.3 Projected Copper Concentrate Quality

Element | Unit | Content
Range | Years 1 to 10 Average | LOM Average
Cu | % | 23 – 28 | 25 | 25
Au | g/t | 28 – 109 | 61 | 43
Ag | g/t | 123 – 505 | 174 | 190
Mo | % | 0.03 – 0.21 | 0.15 | 0.10
ST | % | 28 – 41 | 34 | 34
S-2 | % | 26 - 36 | 32 | 32
Fe | % | 24 – 35 | 29 | 29
Sb | ppm | 199 – 1,966 | 993 | 1,008
As | ppm | 460 – 2,760 | 1,769 | 1,403
Co | ppm | 42 – 97 | 47 | 61
Cd | ppm | 33 - 172 | 88 | 92
Bi | ppm | 15 – 156 | 27 | 48
Hg | ppm | 1.2 – 10 | 2.8 | 3.6
Ni | ppm | 49 – 233 | 76 | 91
F | ppm | 75 - 399 | 241 | 193
Cl | % | 0.01 – 0.02 | 0.01 | 0.01
Se | ppm | 62 – 148 | 74 | 95
P | ppm | 67 – 536 | 156 | 194
Pb | % | 0.1 – 1.0 | 0.5 | 0.4
Zn | % | 0.2 – 0.9 | 0.4 | 0.5
SiO2 | % | 2.3 – 11 | 6.4 | 5.7
CaO | % | 0.2 – 0.8 | 0.5 | 0.5
Al2O3 | % | 0.5 – 3.9 | 2.1 | 1.7
MgO | % | 0.1 – 0.5 | 0.3 | 0.3
MnO | % | 0.01 – 0.04 | 0.01 | 0.02
InSol | % | 3.1 – 12 | - | -

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| 18.0 | Project Infrastructure

| 18.1 | Site Layout

There will be two separate areas of infrastructure associated with the 2016 PFS: the Mine Site and the PTMA. The Mine Site is the centre of mining activity and includes the primary crushing facilities (Mitchell OPC). Process facilities will be located at the Treaty OPC in the PTMA, approximately 23 km northeast of the Mine Site. Twinned tunnels (the MTT) will be constructed from the north side of the Mine Site into the upper reaches of the Treaty OPC. Along the MTT route, a topographical low (valley) is designated as the Saddle Area, approximately 17 km from the Mitchell portal, where the MTT will be accessed via a construction adit. The TMF is located in a valley comprising the upper catchments of North Treaty and South Teigen creeks, southeast of, and adjacent to the Treaty OPC.

The Mine Site layout at the end of construction is shown in Figure 18.1.

The Treaty OPC area is shown in Figure 17.2 and the TMF area is shown in Figure 18.2.

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Figure 18.1 Mine Site Layout after Initial Construction

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Figure 18.2 Ultimate TMF Layout

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| 18.2 | Tailings, Mine Rock, and Water Management

| 18.2.1 | Introduction

This section addresses the geotechnical designs for tailings and mine rock management, as well as for site-wide water management. Designs were updated in 2016 to incorporate incremental improvements and changes to water management structures in the Mine Site and TMF areas, in response to commitments made during the EA process review.

The 2016 PFS design updates in the TMF area included:

| ● | addition of a discharge pipeline system and a diffuser located in Treaty Creek to route operational-period discharges to Treaty Creek

| ● | addition of the North Cell Closure Spillway

| ● | addition of three energy dissipation dams to facilitate road crossings and to contain potential debris flows at significant stream crossings along the South Diversion Channel (this channel follows the TCAR to the Treaty OPC)

| ● | relocation of the TMF seepage collection dams downstream to more effectively intercept seepage.

The 2016 PFS design updates at the Mine Site included:

| ● | relocation of MDT Inlets upstream to improve diverted water quality, and improved inlet design

| ● | a shift of the WSF CDT outlet 200 m upstream to avoid an avalanche area, and design of a closure gate and permanent tunnel plug for the WSF CDT

| ● | modification of discharge from the WSF to use pumped discharge in lieu of discharge pipes passing under the WSD

| ● | expansion of the water treatment plant to a maximum capacity of 7.5 m3/s treatment capacity once the HDS WTP is fully built

| ● | improved water treatment discharge strategy designed to mimic natural flows within the Sulphurets drainage basin

| ● | addition of a Selenium WTP

| ● | placement of the Kerr waste rock in the Sulphurets Pit

| ● | expanding the contact water management systems to handle a 200-year flood

| ● | the inclusion of the MVDT

| ● | design updates for the Sludge Storage Facility and Water Treatment Sludge Storage Building.

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KCB re-assessed additional site climate and hydrology data recorded through 2015. These analyses determined similar values to those adopted for the 2012 PFS (KCB, 2012; KCB, 2013a; Tetra Tech, 2012) for an average year. As a result, the water management design basis remained unchanged from 2012.

Details of TMF, RSF, and SWM prefeasibility design for the 2016 PFS are provided in the following 2016 KCB reports:

| ● | Mine Area Water Management Addendum Report

| ● | Tailings Management Facility Design Addendum Report

| ● | Rock Storage Facility Addendum Report

| ● | Tunnels and Temporary Water Treatment Addendum Report.

| 18.2.2 | Mine Site Characterization

No additional geotechnical site investigations have been completed for the WSD or RSF areas since the 2012 PFS (Tetra Tech, 2012). The 2012 PFS provides a results summary of KCB Mine Site geotechnical site investigations completed up to 2012 (KCB, 2009; 2010; 2011; and 2012).

Seabridge has completed regional geological mapping that resulted in a Mine Site geology map, presented in Figure 18.3.

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Figure 18.3 Mine Site Mapped Geology

Mine Site Climate

Much of the annual precipitation in Mine Site and PTMA occurs as snowfall between October and May, while peak rainfall is associated with storms coming in from the Pacific between August and October. Major elevation variations and numerous glaciers help create diverse climatic conditions across the site.

Two climatic regions are present within the site development areas: the western region in the Sulphurets watershed (Mine Site) and the eastern region in the Treaty-Teigen watersheds (PTMA). The two regions are 23 km apart and have differing climates. The two areas are separated by the Johnstone Icefield (ranging from 1,800 m to 2,200 m in elevation).

Significant orographic and rain shadow effects were recorded in the KSM area as part of the 2012 baseline study. In 2012, KCB and ERM performed extensive analysis of climate variations in Mine Site and PTMA for engineering design and EA purposes (Rescan, 2013). Algorithms were developed based on the UBC watershed model to estimate effects of variation in precipitation with altitude, and to adjust glacier and snow melt rates in response to climatic variations.

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In 2016, metrological and hydrological data collected since 2012 was reviewed by KCB (KCB, 2016a; KCB, 2016b). The result is that no significant trends have been observed in the additional data for average site parameters, as compared to the data prior to 2012.

Mine Site Temperature

Weather data recorded at the Sulphurets weather station between 2007 and 2015 indicate the following:

| ● | mean annual temperature is approximately 0°C

| ● | mean monthly temperatures range from -13°C in December to 14°C in July

| ● | temperature extremes range from -31°C to 30°C

| ● | mean daily temperatures are above freezing from May to October and below freezing from October to May.

Mine Site Precipitation and Hydrology

The estimated mean annual precipitation is 1,652 mm at Sulphurets weather station (elevation of 880 masl). Annual lake evaporation is estimated at 400 mm. Runoff at the Mine Site is influenced by the effects of both seasonal snowmelt and glacial melt. Mitchell and McTagg glaciers are losing significant ice mass on an annual basis. Runoff from glacier-influenced catchments is therefore larger than the annual precipitation over these catchments. Effects of glacial meltwaters are included in the analysis of flows and extreme events. Monthly precipitation, evaporation, and runoff distribution for the Mine Site, as well as the runoff distributions for the glacier catchments, is provided in Table 18.1.

Precipitation listed in Table 18.1 is representative of the Sulphurets weather station at 880 masl and is derived from site data between 2008 and 2011. Additional site data recorded since 2012 has not changed the values adopted for design. Assessment of tipping bucket rain gauge data indicates that precipitation increases at higher elevations within the Mine Site at a nominal rate of +5% per 100 m.

The 2012 and 2016 KCB design reports present detailed analyses of climate and hydrology data for the Mine Site (KCB, 2012a; KCB, 2016b) and PTMA (KCB, 2012; KCB 2016a).

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Table 18.1 Climate Data for the Mine Site (Sulphurets Creek Climate Station)

Data Area | Jan | Feb | Mar | Apr | May | Jun | Jul | Aug | Sep | Oct | Nov | Dec | Total
Precipitation (mm) 1 | 215 | 50 | 50 | 66 | 99 | 83 | 115 | 149 | 264 | 297 | 132 | 132 | 1,652
Pond Evaporation (mm)2 | 0 | 0 | 0 | 0 | 86 | 93 | 99 | 80 | 43 | 0 | 0 | 0 | 400
Site Runoff Distribution (%) | 0 | 0 | 1 | 1 | 4 | 14 | 35 | 17 | 17 | 7 | 3 | 1 | 100
Mitchell Glacier Runoff (mm) | 37 | 37 | 73 | 73 | 110 | 404 | 1,248 | 917 | 514 | 183 | 37 | 37 | 3,670
McTagg Glacier Runoff (mm) | 66 | 33 | 33 | 33 | 197 | 625 | 954 | 526 | 461 | 197 | 99 | 66 | 3,290

| Notes: | 1Weather station at 880 masl elevation. 2Estimated pond evaporation based on pan evaporation data.

Source: KCB

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Mitchell Glacier Recession

Seabridge has been monitoring the recession of site glaciers by analyzing historical air photos, Light Detection and Ranging (LIDAR) data, ongoing global positioning system (GPS) and remote sensing measurements of the glacier extents. Glaciers in the Mine Site area continue to recede; recession rates for the Mitchell Glacier toe area recession since 2008 have reached as much as 65 m/a and total recession has exceeded 2012 estimates. As a result, the Mitchell pit area is now ice free. The locations of the initial stage proposed surface contact water inlets in the toe area of Mitchell Glacier are now also free of ice.

| 18.2.3 | TMF Site Characterization

TMF Site Investigations

No additional site investigations have been conducted at the TMF since the effective date of the 2016 PFS, with the exception of mapping and sampling for till borrow material (KCB, 2019a).

Tailings Characterization and Laboratory Testing

No additional tailings testing has been conducted for TMF design purposes since 2012. The Treaty process plant will produce two tailings streams: the bulk rougher flotation tailings1 representing about 90% of the ore (by dry weight) and a fine, sulphide-rich cleaner tailings comprising 10% of the ore. The sulphide stream will be cyanide leached using the CIL method and then processed for gold recovery. A two-stage cyanide destruction circuit is proposed, using the Inco sulphur dioxide process, followed by hydrogen peroxide treatment2.

The flotation tailings is classified as NPAG and will be cycloned to produce sand fill for construction of the tailings dams during the summer months. The fine cyclone overflow tailings will be discharged along the upstream crest of the tailings dams. The entire flotation tailings stream will be discharged along the dam crests during the winter months.

The CIL residue tailings is a high-sulphide concentration material and is classified as PAG. This material will be deposited under water in the CIL Residue Storage Cell in the centre of the TMF and kept saturated to mitigate the onset of acid generation.

TMF Area Climate

Additional climate and hydrological data collected since 2012 was reviewed (KCB, 2016a) and the conclusion is that no significant trends in average parameters are apparent. As a result, the climate and hydrology parameters used for design of the TMF in 2012 have been retained.

1 Referred to as “Flotation Tailings” in this report.

2This stream is referred to as “CIL Tailings” in this report.

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TMF Area Temperature

Weather data recorded at the TMF area from the Teigen Creek weather station between 2009 and 2011, with additional data now available through 2015, shows the following:

| ● | mean annual temperature is approximately 0°C

| ● | mean monthly temperatures range from -8°C from December to February, to 11°C in July

| ● | temperature extremes range from -27°C to 29°C

| ● | mean daily temperatures are above freezing from May to October and below freezing from October to May.

The monthly precipitation and runoff distribution are provided in Table 18.2.

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Table 18.2 Climate Data for the TMF (Teigen Creek Climate Station)1

Data Area | Jan | Feb | Mar | Apr | May | Jun | Jul | Aug | Sep | Oct | Nov | Dec | Total
Precipitation (mm) | 151 | 110 | 123 | 82 | 55 | 69 | 82 | 82 | 165 | 206 | 164 | 82 | 1,371
Pond Evaporation (mm)2 | 0 | 0 | 0 | 0 | 75 | 81 | 86 | 70 | 38 | 0 | 0 | 0 | 350
Site Runoff Distribution (%) | 1 | 1 | 1 | 3 | 16 | 32 | 19 | 8 | 9 | 7 | 2 | 1 | 100

| Note: | 1Weather station at elevation 1,085 masl. 2Estimated pond evaporation based on pan evaporation data.

Source: KCB

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| 18.2.4 | Rock Storage Facilities

A rendering of the ultimate mine site layout, including the McTagg and Mitchell RSFs is provided in Figure 16.13. Design of the RSFs has not changed since the 2016 PFS.

At the PFS level, there are three primary RSF design considerations:

| ● | foundation conditions

| ● | maximum lift height

| ● | closure slope criteria.

Conservative RSF designs were developed in collaboration with MMTS to address the aforementioned design considerations using existing data. MMTS designed the RSF layouts, with geotechnical guidance on slope stability and geotechnical recommendations from KCB.

The RSFs will be built in progressive lifts (bottom-up construction) to initially confine toe areas and consolidate foundations to improve stability and reduce downslope risks.

| 18.2.5 | Mine Site Water Management

Figure 18.4 illustrates ultimate water management structures as existing at the end of mine life showing diversion tunnel routes and operational phase surface diversions. Catchment boundaries are indicated with blue dashed lines.

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Figure 18.4 Mine Site Ultimate Water Management Facilities

Source: KCB

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After mining is complete, perimeter closure channels will be constructed at the top and margins of the RSFs, and the channels widened for closure along the toes of the RSFs. Operational phase channels are shown on Figure 18.4; closure routings are similar, fresh (non-contact) water routings are in blue and the routing of contact water is in green.

Diversion Tunnels and Surface Diversions

The MDT Inlets were shifted approximately 500 m upstream since the 2012 PFS (Tetra Tech, 2012) as a result of the Application/EIS (Rescan, 2013) to improve water quality of the diverted water. The open pit phase non-contact water MDT was enlarged to allow a single tunnel to carry design diversion flows. Upon commencement of underground mining, the open pit phase tunnel will be twinned with a second tunnel of the same dimensions to increase diversion capacity. There have been no changes to the design of the MTDTs or the Mine Site surface diversions since the 2012 PFS (Tetra Tech, 2012).

Contact Water Collection Systems

The MVDT is a 5 km long, 5 m by 6 m tunnel that drains to the WSF and provides routing of contact water around the Mitchel RSF. The tunnel connects to the NPWDA, which will be added in Year 5, to accept pit wall drainage and local drainage of contact water from upstream of Mitchell pit and from the Snowfields area.

The Mitchell and McTagg RSFs include a Selenium Seepage Collection System that is designed to collect up to 500 L/s of seepage and convey this flow to the Selenium WTP. Water from other sources will be treated when seepage collected from the RSFs is less than 500 L/s. The collection and treatment of seepage from these RSFs, and other high selenium loading waters, will enable selective removal of selenium from flows with higher selenium concentrations, compared to lower concentrations within the WSF. The Selenium Seepage Collection System will be operational by Year 5.

The Selenium WTP will discharge to the WSF to allow further treatment for metals removal.

Water Storage Facility

Seepage from the Mitchell RSF, McTagg RSF, and Sulphurets Pit Backfill RSF requiring treatment for the removal of metals by the HDS process will be collected in the lower Mine Site by the WSD. The WSD is an asphalt core rock fill dam that will create the WSF pond, and will be large enough to handle seasonal freshet flows as well as volume accumulated from a 200-year wet year.

Figure 18.5 shows monthly average water treatment rates for flows from the WSF as blue bars plotted over the LOM. The installed ultimate HDS water treatment capacity of 7.5 m3/s is greater than the annual average or monthly peak flows to allow the treatment rate to vary seasonally with stream flow rates.

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Figure 18.5 Mine Site Monthly Water Treatment Rate

Source: KCB

The WSD design was updated after 2012 PFS (Tetra Tech, 2012) that included dam slope and internal zonation revisions (KCB, 2012d). The slopes and zonation of the dam are shown in section view in Figure 18.6.

The WSD will be located in the lower Mitchell Creek area and founded on competent rock. The WSD crest elevation is also unchanged from the ultimate dam height in the 2016 PFS and will be established at the full height of 716 masl (165 m height) before Year 1. An emergency spillway will be cut into rock on the southeast side of the dam. There will be appropriate freeboard for avalanche wave mitigation and flood routing.

Water in the WSF is predicted to be acidic, similar to existing seeps situated in the upper Mine Site. An asphalt core will be included in the dam to control seepage. Asphalt is inert with respect to acidic water. To control seepage, the WSD and WSF Seepage Dam foundations will be grouted. The depth of the WSD grout curtain is designed to vary from 25 m at the west abutment to as deep as 150 m at the east abutment if required. Grout hole spacing will be 2.5 m.

Fill for dam zones is specified such that critical zones of the dam (e.g., sections in contact with the core) and drain zones will be constructed with materials that have low potential to react with acidic water.

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During the Application/EIS (Rescan, 2013) review process, additional mitigations were developed to minimize seepage from the WSD. These design enhancements are included in the 2016 PFS. Changes include six seepage interception tunnels that lower groundwater levels between the WSD and the WSF Seepage Dam to reduce the driving force on seepage. The seepage collection tunnels will also facilitate foundation grouting both during construction and for remedial grouting after completion of the WSD, if required.

An asphalt-core seepage collection dam will be located downstream of the WSF. Water collected in this dam will be sent to the WTP via an HDPE pipeline.

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Figure 18.6 Water Storage Dam Sections

Source: KCB

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The WSD CDT routes Mitchell Creek around the dam footprint during construction. The diversion tunnel size is designed as 4.4 m by 5.0 m and is sized to pass flows from a 50-year storm event. To reduce exposure to avalanche hazards, the downstream portal was moved upstream from where shown in the EA resulting in a total tunnel length of 900 m.

Hydroelectric Potential of Diversions

Diverting Mitchell and McTagg creeks into tunnels creates an opportunity for hydroelectric power generation. Generated power can be used during mine operations or sold to the grid via the power lines through the MTT. During operations, the hydroelectric plants will reduce the power requirements of the mine. Upon mine closure, the hydroelectric plants will continue operating and will generate income and offset water treatment costs.

The later stages of the MTDTs include hydroelectric generation that comes into operation in Phase 2 (Year 10) with an installed capacity of 8.0 MW. In Phase 2 of the RSF layouts the tunnels will be raised to have inlets above the expansion of the McTagg RSF; during Phase 3 (Year 15) the inlets will be raised again once the RSF reaches its ultimate extent.

Characteristics of hydroelectric plants installed on the diversions are similar to run-of-the-river installations, in that they provide peak power during freshet flows.

| 18.2.6 | Water Treatment

Temporary Mine Area Construction Period Water Treatment

Temporary water treatment facilities and settling ponds are designed to meet current guidelines for mine sediment control and settling ponds.

During the construction period, six TWTPs for TSS and metal removal will be provided in the proposed mine area. Additional TWTPs will be located in the Saddle Area, and at the Treaty portal of the MTT. The TWTPs are intended to deal with drainage from existing mineralized zones, PAG cuts, tunnel portals, and runoff from PAG tunnel muck piles during the period before the permanent WTP is in operation.

The WSF and WTP will be in operation during the six-month pre-production period to capture sediment and runoff from mine area stripping and from fill placement during Mitchell OPC and haul road construction.

A total of eight TWTPs and associated collection ponds will operate during the construction period to manage potential metals, TSS, and ammonia in drainage from tunnel portals and from temporary stockpiles of tunnel muck near the portals and other flows of contact water (Table 18.3).

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Table 18.3 Temporary Water Treatment Plant Locations

TWTP | Location
TWTP #1 | WSD/HDS WTP Area
TWTP #2 | MTDT Outlet
TWTP #3 | MDT Outlet
TWTP #4 | Saddle MTT Portal Tunnel
TWTP #5 | MDT Inlet
TWTP #6 | MTT – Mitchell Portals
TWTP #7 | WSD CDT Inlet
TWTP #8 | Treaty MTT Portals

Source: ERM

High density Sludge Water Treatment Plant

The HDS WTP is designed to treat water that comes in contact with areas of disturbance from mining operations and natural seeps in the area. Design of the WTP is unchanged since the 2016 PFS.

Water will be collected in the WSF. Drainage from the Mitchell pit and Mitchell/McTagg RSFs will be directed by gravity to the WSF and contact water from the Sulphurets and Kerr pit areas will be routed to the WSF by gravity pipeline. The water from the WSF will be pumped over the WSD to the HDS WTP. The HDS WTP is designed with variable discharge rates in order to stage discharge to match the natural hydrograph, to ensure sufficient dilution capacity to minimize any effects on the receiving environment.

The HDS WTP maximum throughput capacity will be 7.5 m3/s; however, the maximum rate is only anticipated to be required for a three-month period in the summer. Water pumped from the WSF will pass through hydroelectric generators installed at the Energy Recovery Power Plant, immediately upstream of the HDS WTP.

The HDS WTP installed generation capacity will be 9 MW and the two installed turbines will be capable of passing a flow of up to 7.5 m3/s.

The site selection for the HDS WTP is based on a 50+ year mine life and post-closure treatment for 200 years. The HDS WTP will be located at an elevation of 520 m on a flat benched terrain above the flood plain near the confluence of Mitchell and Sulphurets creeks.

During operation, the sludge will be transported year-round by truck to the Mitchell OPC and onto the ore trains within the MTT, where it will be transported with the ore to the stockpile located at the Treaty OPC, fed through the Treaty process plant, and deposited with the tailings in the TMF.

The three principal reagents for the HDS WTP will be quick lime, dry flocculant, and sulphuric acid for pH adjustment to 7.5. Table 18.4 provides an estimate of annual reagent consumption based on an annual average of 71 Mm3 of water treated. The predicted total annual volume of water will vary from 63 Mm3 to 79 Mm3. After closure, the predicted long-term volume of water for treatment will be 64 Mm3/a.

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Table 18.4 Annual Reagent Consumption for the HDS WTP

Reagent | Feed Rate | Average Annual Water Treated (Mm3/a) | Total Annual Reagent Consumption
Quick lime to pH 10.5 | 0.83 kg/m3 | 71 | 59,000 t
Magnafloc 10 | 3 g/m3 | 71 | 215 t
H2SO4 to pH 7.5 | 11 mL/m3 of 36.8N H2SO4 | 71 | 780,000 L

Source: ERM

Selenium Treatment Plant

Design of the Selenium Water Treatment Plant (Selenium WTP) is unchanged since the 2016 PFS. BioteQ Environmental Technologies Inc. (BioteQ) demonstrated selenium removal of spiked Mitchell Creek feed water during a pilot-scale ion exchange water treatment study (BioteQ, 2015).

The pilot study demonstrated reduction of selenium concentrations from 120 ppb and 320 ppb feed water to less than 1 ppb (BioteQ, 2015). The Selen-IX™ Ion Exchange Circuit is designed to selectively remove selenate from the feed water with a high efficiency in order to obtain the 1 ppb discharge limit, while concentrating the selenium into a small volume of brine solution that is directed to the eluate treatment circuit. The eluate treatment circuit removes selenium from the spent regenerant (or eluate) solution produced by the ion exchange circuit with an electro-reduction process using iron and fixes the selenium into an iron-selenium solid that is easily separated from solution. The solution discharged from the eluate treatment circuit is largely free of selenium and can be recycled back to the ion exchange circuit for re-use in resin regeneration.

As the Selenium WTP is only designed to remove selenium, effluent from the Selenium WTP will report to the WSF for further treatment at the HDS WTP, prior to discharge to the receiving environment.

| 18.2.7 | Tailings Management Facility Design

The general layout of the TMF is shown in Figure 18.2. The TMF will be located within a cross-valley between Teigen and Treaty creeks. Three cells will be constructed: the North Cell and the South Cell will store flotation tailings, and the CIL Residue Cell (fully lined with a geomembrane) will contain PAG CIL residue tailings.

The cyclone sand dams will be constructed over earth fill starter dams using the centerline construction method with compacted cyclone sand shells and low-permeability glacial till cores. The Saddle and Splitter dam cores incorporate geomembranes to limit seepage from the CIL residue tailings. The dams will be progressively raised over their operating life to an ultimate elevation of 1,068 m.

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Seepage from the impoundment will be controlled with low-permeability zones in the tailings dams and dam foundation treatment. Seepage and runoff from the tailings dams will be collected downstream at seepage collection dams and pumped back to the TMF. The ponds behind the collection dams will also be used to settle solids eroded by runoff from the dam and fines from cyclone sand construction drain-down water.

Tailings Staging Plan

Tailings flows will be initially routed by gravity in slurry pipelines from the plant to the North Cell and CIL Residue Cell. Energy will be recovered during early years of operation of each cell from discharge of the tailings into the impoundment. Tailings will be pumped to the CIL Residue and South Cell when required during later stages of the operation.

Tailings will be retained by four cyclone sand tailings dams: the North Dam, Splitter Dam, Saddle Dam, and Southeast Dam. During operation, elevations of annual dam crest raises will be set to provide 12 months of tailings storage and to store the PMF with 1 m of freeboard.

Figure 18.7 illustrates the staging of the TMF. The North Cell will be constructed first and will store flotation tailings production for 25 years. During operation of the North Cell, floods will be routed south. A pipeline and surface channel will divert environmental maintenance flows of up to 2 m3/s from the East Catchment around the TMF into Teigen Creek. The North Cell will be closed after 25 years of operation and reclaimed over a five-year period. The CIL Residue Cell will be constructed and operated in parallel with the North Cell, and will be filled to about half its capacity with PAG CIL residue tailings. At Year 25, the South Cell goes into operation, providing flotation tailings storage for the remaining mine life. During this time, the East Catchment Tunnel will route East Catchment flood flows away from the South Cell. The CIL Residue Cell will be filled to ultimate capacity. The South Cell and CIL Residue Cell will then be closed and reclaimed over a five-year period.

Based on the mill ramp-up schedule, and the assumed density ranges possible at start up, the starter dams can store between 18 and 24 months of tailings. The earth fill starter dams at the North, Splitter and Saddle dam sites will be constructed to store a minimum of 8.4 Mm3 of water for mill start-up. The design operating PMF ranges from 42 Mm3 at start-up to 91 Mm3 at the ultimate dam elevation. The dams will then be raised annually by cycloning tailings sand. The beach will be built up to separate the reclaim pond from the dams by at least 700 m, increasing to 1,200 m at the ultimate dam elevation. The separation between the tailings dam and pond created by the beach increases the margin of safety against overtopping of the tailings dam, and reduces seepage through the tailings dam and underlying foundation.

A starter dam, at elevation 930 m, will be completed by Year 25 to allow deposition to begin in the South Cell. Between Year 25 and closure, the Southeast Dam will be raised to its ultimate elevation.

Seabridge Gold Inc. | 18-21 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Figure 18.7 TMF Staging Plan

Note: Raising of cyclone dams within each stage is not shown on these diagrams.

Source: KCB

Seabridge Gold Inc. | 18-22 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

TMF Dam Structures

Over an initial two-year construction period, three earth fill starter dams will be constructed at the North Cell and CIL Residue Cell (North, Splitter and Saddle) to provide start-up flotation and CIL residue tailings storage. These dams will be progressively raised over their operating life to an ultimate elevation, providing storage capacity of 2.3 Bt. A summary of the tailings dams is provided in Table 18.5.

TMF Starter Dams

Starter dams will be earth fill embankments, with shells of compacted random fill supporting the central glacial till cores. The glacial till cores will be keyed into the underlying foundations to cut off seepage through weathered near-surface soils and any pervious strata. A blanket drain is provided to lower the phreatic levels in the downstream shell. Riprap erosion protection will not be placed on the upstream slope due to the temporary exposure of the dam to the pond water. Figure 18.8, Figure 18.9 and Figure 18.10 show typical sections of the starter dams.

Main Tailings Dams

The North, Splitter, Saddle, and Southeast dams will be compacted cyclone sand dams with glacial till cores constructed by the centerline method. Dimensions of the dams are summarized in Table 18.5. Details of the North, Splitter, Saddle, and Southeast dam designs are shown in Figure 18.8, Figure 18.9, and Figure 18.10.

A system of finger drains will be installed at the base of the downstream shells of the North, Saddle and Southeast dams to keep water levels in the dam depressed. Main drains in the centre of the valley floor will collect and convey seepage to the toe of the dam. Smaller secondary drains will convey water laterally into the main drains.

Seabridge Gold Inc. | 18-23 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Table 18.5 Tailings Dam Summary

Dam | Starter Dam | Ultimate Dam
Crest Elevation (masl) | Maximum Height* (m) | Crest Length (m) | Random Fill Volume (Mm3) | Core Volume (Mm3) | Crest Elevation (masl) | Maximum Height* (m) | Ultimate Crest Length (m) | Cyclone Sand Volume (Mm3) | Core Volume above Starter (Mm3)
North Dam | 930 | 80 | 680 | 3.59 | 0.95 | 1,068 | 218 | 1,900 | 47.16 | 3.42
Splitter Dam | 935 | 61 | 890 | 3.74 | 1.08 | 1,068 | 194 | 1,930 | 31.31 | 3.75
Saddle Dam | 935 | 35 | 780 | 2.09 | 0.75 | 1,068 | 168 | 1,600 | 22.99 | 3.39
Southeast Dam | 930 | 101 | 890 | 12.32 | 1.72 | 1,068 | 239 | 1,400 | 60.45 | 3.20
Totals | - | - | 3,240 | 21.73 | 4.50 | - | - | 6,830 | 161.91 | 13.77

Note: *maximum height measured at dam crestline

Source: KCB

Seabridge Gold Inc. | 18-24 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Figure 18.8 North Tailings Dam

Source: KCB

Seabridge Gold Inc. | 18-25 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Figure 18.9 Saddle and Splitter Tailings Dams

Source: KCB

Seabridge Gold Inc. | 18-26 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Figure 18.10 Southeast Tailings Dam

Source: KCB

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

TMF Dam Construction

Table 18.5 summarizes fill requirements for the dams. For construction of the starter dams, general fill and core material will be excavated by a contractor fleet from local borrow sources (less than 2 km haul distance) that were identified at each dam site. The cyclone sand TMF dams will be raised using a fleet of dedicated mine equipment.

TMF Seepage Recovery Dams

Seepage recovery dams will be constructed of compacted glacial till in a similar manner as for the tailings starter dams, but with flatter 3H:1V upstream and downstream slopes.

TMF Area Water Management

Figure 18.11 shows the schematic water balance with water inputs and outputs from the impoundment.

Seabridge Gold Inc. | 18-28 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Figure 18.11 Schematic TMF Water Cycle

Source: KCB

Seabridge Gold Inc. | 18-29 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

TMF Diversion Channels

Two main diversion channels, the Northeast Diversion and the South Diversion, will be constructed around the TMF North Cell with additional diversions around the Treaty OPC to divert non-contact runoff water into a tributary of Teigen Creek at the north end of the TMF.

Diversion channels are shown on Figure 18.12.

TMF Area Extreme Flood Routing and Storage

The TMF cells are designed to be able to store extreme flood events without discharge. Specifically, the PMF can be stored, which is a flood resulting from a 30-day Probable Maximum Precipitation (PMP) storm, combined with a 100-year 30-day snow melt. The perimeter diversions are assumed to be inoperative during this extreme flood event.

During operations, water will be reclaimed from the ponds and routed back to the Treaty OPC, where it will be treated as part of the mineral separation process. Surplus water from the TMF will be discharged seasonally via the Treaty Creek Diffuser. Discharge will occur during an approximate period extending from May to mid-November, when the creek flows are highest.

Seabridge Gold Inc. | 18-30 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Figure 18.12 Ultimate TMF with Catchments and Diversion Channels

Source: KCB

Seabridge Gold Inc. | 18-31 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 18.3 | Tunnels

A number of tunnels will be excavated during both the pre-production period and the operating period. These tunnels will be classified as either infrastructure tunnels or water tunnels. Table 18.6 summarizes tunnels constructed in pre-production and Table 18.7 summarizes tunnels constructed in the operational phase. Tunnels associated specifically with block cave mining, are presented in Section 16.3.

The infrastructure tunnels provide for the transportation of ore, personnel, and supplies between the Mitchell, Sulphurets, and Kerr mining areas and the Treaty OPC. The principal infrastructure tunnel is the MTT, which transports all mined ore from the Mitchell OPC to the Treaty OPC, and personnel and freight between the PTMA and the Mine Site, via the train haulage system. Other infrastructure tunnels include load out, unloading, and freight sidings in the MTT, a spur off the MTT to Iron Cap, and an ore conveyor tunnel from the Sulphurets pit to the Mitchell OPC.

The water tunnels include the diversion tunnels as described in the SWM Plan and the slope drainage tunnels for the Mitchell high wall and the Snowfields landslide.

This section includes a description of the construction method, sequencing, and cost basis for the tunnels as applied to the PFS schedule and capital estimate. The MTT design cross section has been modified since the EA to accommodate the change to train haulage from the previous ore conveyor system; however the alignment remains the same.

Table 18.6 KSM Pre-Production Tunnels Summary

Tunnel | Description | Total Length (m) | Excavation Volume (m3)
Infrastructure Tunnels
MTT | Principal Alignment | 49,406 | 1,266,770
Treaty Ore Handling | 514 | 27,174
Mitchell Ore Handling | 943 | 43,784
Mitchell Freight and Fuel | 576 | 28,898
MTT Subtotal | 51,439 | 1,366,626
Water Tunnels
MDT – Open Pit Phase | Diversion Tunnel Inlets | 1,241 | 32,899
Access Tunnels | 870 | 20,402
Main Diversion Tunnel | 7,000 | 185,570
MDT Subtotal | 9,111 | 238,871
MTDT | Stage I | 8,020 | 253,038
MVDT and Mitchell OPC Decline | - | 5,424 | 182,787
CDT at WSD | Construction Diversion Tunnel | 900 | 18,360
SCT at WSD | Seepage Collection Tunnels | 1,780 | 19,402
Pre-production Tunnels Total | - | 76,674 | 2,079,084

| Note: | Tunnel volumes are based on construction volumes and account for drill hole “look out” contractor tolerance and up to 200 mm of shotcrete as required.

Source: KCB

Seabridge Gold Inc. | 18-32 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Table 18.7 KSM Operational Phase Tunnels Summary

Tunnel | Description | Total Length (m) | Excavation Volume (m3)
Infrastructure Tunnels
Iron Cap Connection | - | 972 | 35,983
SMCT | - | 3,025 | 108,419
Water Tunnels
MDT
MDT – Underground Phase | Diversion Tunnel Inlets | 1,095 | 29,028
Access Tunnels | 812 | 29,516
Main Diversion Tunnel | 7,000 | 185,570
MDT Subtotal | 8,907 | 244,114
MTDT | Stage II | 6,742 | 136,469
Stage III | 7,940 | 125,135
MTDT Subtotal | 14,682 | 261,604
NPWDA and SSDA | Main NPWDA | 3,000 | 100,260
Access Tunnels | 600 | 12,834
Inlet Tunnels | 589 | 26,855
SSDA | 1,104 | 22,345
NPWDA and SSDA Subtotal | 5,293 | 162,294
East Catchment Tunnel | at TMF | 4,000 | 67,733
Mitchell Block Cave Dewatering Tunnels | Underground Phase | 12,000 | 675,000
Operation Phase Tunnels Total | - | 48,879 | 1,555,147

| Note: | Tunnel volumes are based on construction volumes and account for drill hole “look out” contractor tolerance and up to 200 mm of shotcrete as required.

Source: KCB

| 18.3.1 | Mitchell-Treaty Tunnels

The MTT is an electric rail-based twinned haulage system to handle all ore from the Mitchell mining areas to the Treaty OPC as well as personnel transport, supplies, and services between the two operating areas. Personnel transport includes mine side shift exchange as well as daily personnel interchange. Supplies and services include rail-based fuel tanks, explosives, mine operating supplies, maintenance parts and supplies, and an electric power line.

MTT Design

Primary crushing of ore from the Kerr, Sulphurets, and Mitchell open pits will be done proximal to each open pit. The crushed ore will be transported through the MTT to the crushed ore stockpile located at the Treaty OPC, approximately 23 km to the east. Future underground ore from the Mitchell and Iron Cap block cave mines will connect with the Mitchell-Treaty ore transport system. The twinned configuration provides higher capacity with the haulage loop and rail cross-overs allow sections of the tunnel to be isolated for periodic tunnel maintenance. Under normal operations, the North Tunnel will be designated for westbound travel and the South Tunnel will be designated for eastbound travel.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

The MTT will comprise the following excavations:

| ● | two, 22,715 m long tunnels

| ● | 75, 30 m long cross-cuts, linking the two tunnels

| - | 36 of these will be equipped with refuge stations

| ● | nine track cross-overs to allow for flexibility and tunnel maintenance during operations

| ● | Saddle Adit – a 265 m long decline at 15% complete with waste transfer chamber (for use during construction)

| ● | various other excavations for ore bins, conveyor tunnels, escape raises, train unloading and freight/fuel handling areas.

MTT Tunnel Support and Advance Rates

The length of the tunnel has been broken into 4 general ground support classes to account for the different rock types encountered based on field investigations comprising mapping drilling and testing. These support class segments are accounted for the tunnel advance rates, and costs per meter for costing and scheduling.

MTT Mining Method

The tunnels will be constructed in accordance with BC Mines Act and Health, Safety and Reclamation Code for Mines in BC using conventional drill and blast techniques, and will follow the conditions contained within the License of Occupation for the MTT issued in September 2013 by the Government of BC.

The MTT will be excavated by a contractor using conventional drill and blast with rail haulage to the portals. Electrical mining equipment will be used, reducing ventilation requirements. Excavated rock will be temporarily stored at the portals in managed facilities and relocated to permanent rock disposal sites when the permanent waste management plan is in operation.

Towards the end of the MTT construction period, the contractor will lay the track and install the electrical supply and distribution system for the ore trains.

Water Management

Both MTT tunnels will be driven with a ditch in the floor and excavated in the corner that will contain a perforated pipe for the collection of tunnel water during the operations period; the pipe will be buried under the track ballast. The nominal grade of the MTT is 1.2% down from the Treaty portals to the Mitchell portals; therefore, in operation the water in the tunnels will flow to the Mitchell portal where it will be routed to the WSF.

Seabridge Gold Inc. | 18-34 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

During construction, tunnel water will be collected and treated by temporary water treatment plants located near the portals.

MTT Mining Sequence

The MTT is on the construction critical path and has therefore been broken into two segments to allow for concurrent development workplaces resulting in a shorter total tunnel construction period. This preliminary sequence will be accomplished using the Saddle, which is a transverse valley along the tunnel alignment, located approximately 6.1 km from the Treaty Portal of the MTT and 16.6 km from the Mitchell portal. An adit will be driven at a negative grade from the Saddle and access both the north and south MTT tunnels. It will allow the Mitchell-Saddle segment of the MTT to be driven from two headings from either end of the segment, with four independent crews at one time. Additionally, two crews will be advancing the Treaty-Saddle segments from the Treaty portals, one in the North Tunnel and one in the South Tunnel.

Crews from the shorter Saddle to Treaty segment of the MTT will finish early and will then be used to develop the other excavations. The North and South tunnels will advance together both from the Mitchell and Saddle headings. As the twin headings advance, cross-cuts will be developed every 300 m joining the two tunnels. The cross-cut closest to the face will be used for remucking, the next closest cross-cut will be used for the ventilation cross-over and the cross-cut before that will be sealed and equipped with a refuge station.

Ventilation During Construction

Since there are three sets of two headings advancing together, there will be three primary ventilation circuits. The primary circuit will be established by installing two fans in a bulkhead just inside the portal in the South Tunnel (and near the entrance to the Saddle adit), to provide fresh air under positive pressure through the South Tunnel, and through the ventilation cross-cut with exhaust out the North Tunnel. The secondary circuits will be established to intercept fresh air from the primary circuits in order to ventilate the advancing faces. This will be done by two auxiliary fans with flexible vent ducting installed in the South Tunnel on the fresh air side of the active ventilation cross-cut and blowing air to the advancing headings in each of the South and North tunnels. The air from the South Tunnel will exhaust via the ventilation cross-cut where it will meet with the exhaust air from the North Tunnel. This exhaust air stream will then flow out the portal. As the tunnel faces advance, a new remuck cross-cut will be established and the previous remuck cross-cut will now act as the new ventilation cross-cut. The previous ventilation cross-cut will be sealed and equipped with the advancing refuge station.

Ventilation During Operations

During normal operations air will be moved through the tunnel by the piston effect of the trains. Thus, ventilation in the North Tunnel will generally be from east to west and ventilation in the South Tunnel will generally be from west to east.

To allow for segments of the MTT to be isolated for maintenance, sets of ventilation doors with axial vane fans will be installed at the portals and at the track cross-overs. In this way fresh air will be supplied to the isolated sections of track where the crews will be working. Energizing or de-energizing of fans will be coordinated with train traffic so that the fans will not move air against each other or against closed vent doors.

Seabridge Gold Inc. | 18-35 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 18.4 | Mine To Mill Ore Transport System

At the Mitchell OPC, ore will be crushed and conveyed through a tunnel to two live underground bins within the MTT. Loading chutes under the bins will feed into trains that will transport the ore to an unloading station at the Treaty end of the MTT. The train cars will dump into a live underground unloading bin. Apron feeders will unload the bin onto a conveyor to transport the ore to the top of the Treaty COS.

Each train will consist of a 140 t electric locomotive and 16 x 42 m3 belly dump ore cars that have the capacity to deliver 800 t/h from Mitchell to Treaty based on 90-minute cycle times. On average, eight trains will be in operation to meet the process plant requirements with an additional four trains on standby to provide for mechanical availability or to handle an increase in plant feed of up to 10,000 t/h, when required to meet the process plant feed of 130,000 t/d. The transport system capacity was confirmed at PFS level by using a high-level train routing simulation that incorporates static modelling of the rail system and average cycle times between Mitchell and Treaty for delivery of ore, materials, and personnel.

The trains will travel on a conventional ballasted track structure with timber ties and operate via an electrical overhead catenary system. Trains will be controlled by an automated train control system managed from a remote control room. Loading chutes will also be controlled remotely and unloading chutes will operate autonomously. No onboard operators will be required within the tunnels during train system operation.

Figure 18.13 shows a plan view of the approximately 23 km long MTT dual track transport system. The South Tunnel will be primarily utilized for loaded trains travelling to Treaty OPC, and the North Tunnel will be utilized for empty trains travelling back to Mitchell OPC. The tunnels will run uphill from Mitchell OPC to Treaty OPC at a nominal 1.2% incline, such that tunnel drainage will flow to the Mine Site. Cross-overs are planned at both end points of the tunnel, as well as in two intermediate points to split the route into three sections. One section of the tunnel can be isolated when maintenance is required, and one-way traffic can be implemented and safely controlled by the train automation system. A simulation verified that these traffic flow restrictions will not compromise average daily train production requirements.

Seabridge Gold Inc. | 18-36 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Figure 18.13 MTT Dual Track Plan View (Distances in Metres)

Source: MMTS

Seabridge Gold Inc. | 18-37 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

The numerical demarcations in Figure 18.13 show the distance, in meters, along the MTT tunnel of noted features from Mitchell to Treaty. The Golder Alignment refers to the planned access from a location near the Mitchell portal to the underground block cave operations.

Figure 18.14 shows a typical cross-section through both of the MTT train tunnels.

Figure 18.14 MTT Train Transport Drift Section (Dimensions in Millimeters)

Source: MMTS

The 140 t locomotives are powered by an electrical overhead 25 kV alternating current (AC) catenary system that includes:

| ● | two, 25 kV medium voltage substations

| ● | three low-voltage substations equipped with a transformer 25 kV/0.6 kV/208 V, and 600 V/208 V control panel

| ● | 12 rectifiers, 6 in the North Tunnel and 6 in the South Tunnel

| ● | a Siemens SCAT SR aluminum overhead conductor rail system.

Seabridge Gold Inc. | 18-38 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Loaded trains will travel on an uphill slope, requiring a tractive force of approximately 275 kN. At a speed of 30 km/h, power consumption of 2,400 kW per train is expected and will require a current draw of 1,600 A from the catenary at an operational voltage of 1,500 V. Empty trains travelling downhill at 50 km/h will feed approximately 400 kW of energy back into the grid.

The locomotives, as well as the loading and unloading chutes, will carry their own fire suppression systems, and will not require a fixed system within the tunnel.

An automated train dispatching system will be utilized to achieve a safe and efficient flow of trains through the tunnels,

As shown in Figure 18.13, two parallel underground spur lines coming off of the main tunnel will be used for train loading at the Mitchell end of the MTT.

Trains will be continuously loaded without the need to stop between cars. Each 42 m3 car can be loaded in an average of 36 seconds. The mine ore cars adopted for this PFS are end-hinged cars with an overlapping apron between the cars.

From the bottom of the bin, ore will discharge into two apron feeders and onto a conveyor belt that will transport the ore to the surface and feed to the Treaty COS.

| 18.4.1 | MTT Freight and Personnel Transport

The train system will also be used for transportation of personnel and freight between the Treaty and Mitchell areas via the MTT. Freight and personnel transport will be scheduled on a daily basis, and the transport trains will be controlled by the automated train control system. Specially configured personnel and freight trains will transport personnel, freight, and fuel through the MTT, with marshalling and unloading areas at each end, separate from the ongoing ore transportation facilities. Personnel, freight, and fuel handling will only be scheduled during the day shift operations.

The train control system will ensure there is no haulage of fuel or explosives when personnel are being transported in the tunnels. In the event of an emergency, the personnel cars will be equipped with personal protective equipment (PPE) kits including self-rescuers. The twin tunnel configuration and cross-cuts provide alternate egress and rescue stations will be installed through the length of the MTT.

Treaty Staging Areas

Staging areas on surface, near the Treaty portal will be used to load personnel, freight, and fuel onto the specialty train cars. These staging areas will be road accessible and include a laydown area for all freight that is for transport through the MTT.

The personnel marshalling will be out of the North Tunnel and will include a structure to protect waiting passengers from the elements.

A train maintenance shop will be located in the freight marshalling area. The workshop facilities are enclosed in a building on the south side of the pad, located outside the Treaty portal. It also includes the control room for train system operations (see Figure 18.15).

Seabridge Gold Inc. | 18-39 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Figure 18.15 Treaty Personnel, Freight, and Fuel Staging and Marshalling

Source: MMTS

Mitchell Staging Areas

Three separate, enclosed, underground staging areas near the Mitchell portal will be used to offload passengers, freight, and fuel, respectively (shown in Figure 18.16). Loaded fuel tanks and other hazardous freight will be shuttled out of the tunnel as soon as possible.

Personnel will exit the Mitchell portal by bus or other light vehicles.

Freight and fuel staging areas will include gantry cranes to offload the train payloads onto awaiting flatbed tractor-trailer units. Freight will be driven out to its ultimate destination at the Mine Site.

Seabridge Gold Inc. | 18-40 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Figure 18.16 Mitchell Personnel, Freight, and Fuel Staging and Marshalling

Source: MMTS

| 18.5 | Site Roads

Currently, the KSM Site can only be accessed by helicopter. Helicopter support will augment the road pioneering work and construction camps set up.

Avalanche protection will be constructed where appropriate so that work can be safely carried out at the tunnel portals. Rock storage landforms will be developed adjacent to the Mitchell pit.

The Mine Site roads are outlined in the Figure 18.17, with the starter pits outlines shown in blue.

Seabridge Gold Inc. | 18-41 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Figure 18.17 Mine Site Roads

Source: MMTS

| 18.5.1 | Road Width

Haul road widths were designed to comply with the following BC Mines Regulations:

| ● | for dual lane traffic, a travel width of not less than three times the width of the widest haul vehicle used on the road

| ● | for single lane traffic, a travel width of not less than two times the width of the widest haul vehicle used on the road

| ● | a berm height of at least three-quarters the height of the largest tire on any vehicle hauling along the road, where a drop-off of greater than 3 m exists.

Ditches are included within the travel width allowance.

Seabridge Gold Inc. | 18-42 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 18.6 | Ancillary Buildings

Ancillary building construction considered for the PFS will be pre-engineered, stick-built, or modular structures, as applicable. The HVAC for these buildings will be designed to industrial standards. The following ancillary buildings and infrastructure are included in the PFS:

| ● | Treaty OPC:

| - | fuel storage facility

| - | fuel distribution station

| - | administration building

| - | assay and metallurgical laboratories

| - | warehouse and maintenance building

| - | concentrate storage/load-out building

| - | cold storage/reagent storage building

| - | first-aid building

| - | ore train storage yard, maintenance shop and loading/unloading facilities

| - | permanent operations camp

| - | potable water treatment plant

| - | sewage treatment plant

| - | incinerator

| - | substation and auxiliary power supply facilities

| - | construction laydown areas

| - | pre-construction fuel storage

| ● | EPCM and contractors’ offices, concrete batch plant, temporary construction camps, TWTPs, and numerous other construction related facilities

| ● | Mitchell OPC and lower Mine Site areas:

| - | truck shop including first aid facilities

| - | HDS WTP with sludge storage facilities

| - | Selenium WTP (operational by Year 5)

| - | diesel fuel storage and dispensing

| - | permanent operations camp

| - | temporary construction camps

| - | sewage treatment plants

| - | incinerator

| ● | off-site facility:

| - | new concentrate storage and loadout at the Stewart port facility.

Seabridge Gold Inc. | 18-43 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 18.6.1 | Treaty OPC

Fuel Storage and Distribution (Permanent and Construction)

The main fuel storage tanks at the Treaty OPC are sized to store 2,500,000 L, which accounts for 6 days of fuel demand. All fuel storage areas will be lined with containment berms and approved double-wall type tanks. Additional fuel stations will be located near the Mitchell OPC, truck shop, and at the Sulphurets and Kerr pits. Gasoline will also be similarly stored where required.

The majority of the fuel requirement is for mining activities at Mitchell; however, some fuel will be distributed to all Treaty OPC facilities via pipelines from the Treaty Fuel Storage Tank to the required facilities.

A pipeline from the Treaty fuel storage tank to the train marshalling area will be installed, with hook-ups for fast fuel transfer directly to the ISO fuel tanks at the marshalling area. Wherever possible, the ISO fuel tanks will not be removed from the train cars on the Treaty end.

Administration Building

The pre-engineered administration building will be approximately 1,000 m2 in plan area.

Assay, Metallurgical, Acid-base Accounting, and Geotechnical Laboratory

The pre-engineered laboratory will be a 815 m2 single-storey structure located in a separate building near the process plant at the Treaty OPC.

First Aid Buildings

The first aid buildings will be pre-engineered structures, located at both the Mine Site and Treaty OPC, equipped with first aid facilities and provide emergency vehicle storage.

Concentrate Storage

The on-site concentrate storage facility will be a pre-engineered structure, approximately 2,000 m2 in area. It will have a five-day storage capacity equating to approximately 4,600 t of concentrate.

Cold Storage/Reagent Storage Building

The cold storage/reagent storage building will be located at the Treaty OPC and will be approximately 1,200 m2 in area.

Permanent Camp

The Treaty operations camp will be located at the PTMA, approximately 600m southwest from the process plant. The camp components will include accommodation, office/recreation complex, kitchen/diner, parking, sanitary sewer, potable water treatment, wastewater treatment and disposal field.

Warehouse and Maintenance Building

An 800 m2 warehouse and maintenance pre-engineered building will be constructed at the Treaty OPC. It will be located adjacent to the cold storage facility.

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| 18.6.2 | Mine Site

Truck Shop

The Mine Site truck shop will be a pre-engineered building, approximately 9,500 m2 in area. The truck shop/mine dry will comprise eight maintenance bays, two light vehicle repair bays, a truck lube bay, a truck wash bay, a welding and machine shop, an electrical and instrument shop, a storage warehouse and a dry area.

Acid-base Accounting and Geotechnical Laboratory

The pre-engineered laboratory will be a 815 m2 single-storey structure located in a separate building near the truck shop.

Permanent Camp

The Mitchell operations camp will be located between Sulphurets and Gingras creeks, just north of the CCAR. Like the Treaty operating camp, the Mitchell operating camp is planned to be used for most of the construction period and the entire life of mine.

The camp components will include a helipad, sleeping dorms, parking, fuel storage and loading area, recreation facility, sewage treatment, fire/fresh water tanks, an emergency generator, laundry, and kitchen/diner.

Landfills

Two landfills are proposed to be permitted and developed, one for the Mine Site and one for the PTMA.

The Mine Site landfill, which will occupy approximately 6.5 ha, will be located within the Sulphurets laydown area.

The PTMA landfill, occupying approximately 8.4 ha, will be located near the Treaty operating camp.

Each landfill will include a land farm. The land farms will accept contaminated soils from spill clean-ups and leaks, while the landfill will be used to dispose of non-inert, dry industrial, and forestry waste.

| 18.7 | Sewage

The wastewater treatment system installed at the construction and operation camps in both Mine Site and PTMA will treat the anticipated maximum daily flow through a variety of processes to meet a secondary level of treatment as defined in the Municipal Wastewater Regulation (MWR) (BC Reg. 87/2012) of the Environmental Management Act (2003).

For temporary construction camps off site with occupancy less than 100, Type 3 effluent quality as defined in the Public Health Act (2008) sewerage guidelines will be applicable to the sewage requirements for these camps.

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| 18.8 | Communications System

Permanent communications utilizing fibre optic cables will be installed over the existing NTL, and the new 30 km spur transmission line to the Treaty OPC. Installation of fibre-optic cable on site will also allow for the installation of dedicated cellular service.

A fibre optic communication system will be installed in conjunction with the power distribution system in both the Treaty OPC and the Mine Site. A fibre-optic cable has been included in the MTT to provide communications between Treaty OPC and the Mine Site.

An ultra-high frequency (UHF) radio system will be used for mobile communications in both the PTMA and the Mine Site. Base stations and repeaters will be installed as necessary on ground and inside tunnels.

Treaty OPC wired telephone service will be provided by a Voice over Internet Protocol (VoIP) system. A local cell phone system is also planned, as is satellite television for the camps.

In addition, uninterruptible power supplies will be used to provide backup power to communication systems and critical control systems to facilitate orderly shutdown of process equipment and to back up computers and control systems.

| 18.9 | Fresh and Potable Water Supply

Fresh and potable water for the Treaty OPC will be supplied from nearby wells to an elevated storage tank approximately 12 m in diameter and 9 m in height. Fresh water will be used primarily as:

| ● | fire water for emergencies

| ● | cooling water for mill motors, mill lubrication systems and reagent preparation

| ● | the potable water supply.

| 18.10 | Power Supply and Primary Distribution

Power generation and transmission utilities in BC are regulated by the British Columbia Utilities Commission (BCUC), acting under the Utilities Commission Act. BC Hydro generates the majority of power in BC, although there are an increasing number of private Independent Power Producers (IPPs). BC Hydro owns and operates the major transmission and distribution system in BC and is the electric utility that would serve the KSM Mine via the recently constructed NTL.

The utility interconnection capital cost for the KSM Mine will be as set out in the facilities agreement, an arrangement approved by the BCUC in January 1991, pursuant to Order G-4-91 that sets out the rights and obligations of BC Hydro and the customer for construction, ownership, and operation of the facilities necessary for electric service.

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| 18.10.1 | Northwest Transmission Line

The 344 km long, 287 kV NTL runs from the Skeena substation near Terrace, BC, to a new substation near Bob Quinn Lake (Figure 18.18). This new transmission line was commissioned in the summer of 2014 and currently serves the AltaGas Forrest Kerr Hydroelectric Facility and the Red Chris Mine. A tap from this transmission line will service the KSM Mine.

Due to an overrun in the construction cost of the NTL, BC Hydro Tariff Supplement TS37, as approved by the BCUC, was put in place requiring NTL customers to share in the overrun cost. In accordance with TS37, based on a KSM contract (peak) demand of 200 MVA, the required contribution will be just over Cdn$209 million. This amount is separate from system reinforcement and is a required cash contribution. Payment of the tariff is not due until the start of commercial production, and BC Hydro offers the option of spreading the payments out over five years, with an applicable finance charge. This required cost contribution is included in the 2016 PFS sustaining capital costs.

Figure 18.18 NTL Route Map

Source: BC Hydro

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| 18.10.2 | Treaty Creek Switching Station

BC Hydro is responsible to deliver power from the transmission system to a customer at the point of delivery (POD). The customer is responsible to bring power from the POD to their site. For KSM, the POD will be the Treaty Creek Switching Station. The KSM site will take electrical service via a 30 km long, 287 kV line extension from the Treaty Creek Switching Station, to be located on the NTL adjacent to Highway 37, approximately 18 km south of Bell II. This installation will also be in the vicinity of the Treaty Creek Access Road junction with Highway 37. Metering will also be located at this point. The Treaty Creek Switching Station will form part of the BC Hydro system and will be constructed, owned, and operated by BC Hydro. BC Hydro has completed site selection and preliminary design.

Seabridge previously reimbursed BC Hydro for the cost of installing transmission line dead end structures when the NTL was constructed, as required to facilitate the connection of the of the proposed Treaty Creek Switching Station into the grid.

The power supply facility infrastructure will include BC Hydro System Reinforcement and the Basic Line Extension, which is the circuit breaker and metering at the POD. It is not the transmission line to the KSM site. Construction of the Treaty Creek Switching Station, as per the BC Hydro Facilities Study, will require a direct cash payment from the developer of KSM to cover a large part of the cost of the installation that is classified as the Basic Line Extension. This cost is included in the 2016 PFS capital cost estimate. The remaining station cost is classified as System Reinforcement and is not a KSM development capital cost. However, Tariff Supplement (TS) No. 6 (TS 6) Clause 5 (c) “Offset” requires that the customer provide bonding for up to seven years, such that BC Hydro is assured of receiving enough revenue from the KSM operation to justify the capital expenditure. Security is required in the form specified in TS 6 Clause 13 and in the amount as per Clause 5(b). The foregoing bonding and charges are set out under the tariffs and are not negotiable. BC Hydro will return part of the security each year as the offset, based on power billing, reduces the required bonding. The amount of the bonding is not included under the direct capital cost budget, but is otherwise accounted for in the 2016 PFS economics. An important point to note is that as per the tariffs, the customer must pay the actual final cost of construction, not the amount estimated by BC Hydro in a Facilities Study.

BC Hydro is responsible for obtaining all approvals and permits for the Treaty Creek Switching Station. A formal environmental assessment of the Treaty Creek Switching Station under BC’s environmental assessment process for reviewing major projects is not required.

| 18.10.3 | Transmission Line Extension To KSM

The voltage selection for the proposed 287 kV transmission line extension for KSM was based on the 287 kV transmission voltage selected for the NTL. A review of the technical requirements to serve a load larger than 150 MVA confirmed that stepping the voltage down to a lower level is not technically acceptable nor economic.

The developer/operator of KSM Mine will be responsible for the construction and operation of the transmission line extension, in accordance with the established BC Hydro tariff requirements. Line construction will utilize steel monopoles, such that the line can be generally run in the TCAR right-of-way, beside the road, thus largely eliminating the requirement for a separate access route.

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The 287 kV transmission line from the BC Hydro Treaty Creek Switching Station will cross Highway 37 and the Bell-Irving River, then closely follow the mine access road along the north side of Treaty Creek for approximately 12 km to a deviation point where the line transitions from following the TCAR, to following the South Diversion Cut-off Ditch, up to Substation No. 1 at the Treaty OPC. Steel monopoles are ideal for use where a transmission line is to be constructed next to a road and in areas of high snow fall. To protect against avalanche damage, several structures will be mounted on concrete piers, to raise the pole bases above the avalanche flow.

The environmental assessment for the 30 km section of transmission line from Treaty Creek to Substation No. 1 was included in the EA; therefore, approval is in place. Land tenure has been obtained for the transmission line right-of-way, from the Treaty Creek Switching Station to Substation No. 1 at the Treaty OPC.

The transmission line includes a fibre optic cable connection to the BC Hydro NTL fibre-optic cable system as required by the utility. This fibre connection will also carry the general communications to site for the permanent operations phase.

The transmission line land tenure has been obtained. No additional (specific) permits will be required other than the general mine permitting. The cost of right-of-way clearing is included in the road clearing budget.

| 18.10.4 | System Studies

BC Hydro performs studies to determine the cost, method, and timing of transmission system customer interconnections. Seabridge first commissioned a BC Hydro System Impact Study (SIS) for KSM in 2009 to confirm the technical viability of the interconnection. Subsequently, several updates were commissioned to account for the construction of the NTL and several design changes, including changes to the KSM Site access route from Highway 37.

System load flow studies have been performed by PFS consultants using system analysis software to confirm process plant and mine power system voltage control from no load to full load. System voltage stabilization is based on switched reactors to control light load over voltages due to 287 kV transmission line and 138 kV cable capacitance, and also assumes power transformers have automatic tap changers and that there is automatic control of the process plant synchronous ball mill drive motor excitation systems for instantaneous voltage control, as requested by BC Hydro. Substation No. 1 also includes a ±20 MVA static var compensator, as identified by the BC Hydro SIS as required, to ensure system transient stability.

Service from the Skeena Substation to the KSM Site via the NTL will be delivered over a single-circuit line. BC Hydro service studies indicate very high reliability for single-circuit high-voltage transmission lines, with few outage-hours in a year. Occasional service interruptions and planned maintenance outages can be expected and are considered normal for mining projects.

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| 18.10.5 | Electric Utility Requirements, Tariffs, And Cost of Electric Power

The electric service to the KSM Site (including all terms and conditions such as rates and metering requirements, connection charges, and many aspects of the KSM connecting transmission line) will be in accordance with the latest edition of BC Hydro Electric Tariffs, in particular:

| ● | Rate Schedule 1823 – Transmission Service – Stepped Rate

| ● | Rate Schedule 1901 – Deferral Account Rate Rider

| ● | BC Hydro Electric Tariff Supplement No. 5 (TS5) Agreement for Customers Taking Electricity under 1821 (1821 is now 1823) (TS5 is a template for the Electricity Supply Agreement with the format set as per the tariffs and is not subject to change)

| ● | BC Hydro Electric Tariff Supplement No. 6 (TS6) Agreement for Transmission Service Customers (TS6 is a fill in the blanks template for the Facilities Agreement with the format set as per the tariffs and is not subject to change)

| ● | BC Hydro Electric Tariff Supplement No. 37 (TS37) covering the required NTL “overrun” Contribution

| ● | BC Hydro Electric Tariff Supplement No. 74 (TS74) Customer Baseline Load Determination Guidelines.

BC Hydro Rate Schedule 1823 is a two-tier schedule, nominally with 90% of the Customer Baseline Load charged at economical Tier 1 energy rates, and the last 10%, plus all power above the Customer Baseline Load, charged at costly Tier 2 rates. This system is designed to encourage energy conservation, as consumption reductions due to energy conservation measures are applied against costly Tier 2 power. BC Hydro, under their Power Smart program for demand side load control, offer incentives to transmission customers to reduce energy consumption, and for new customers incentives are given for energy-efficient plant design.

The calculated cost is below regular rates due to a large reduction or elimination of costly Tier 2 energy in accordance with an efficient plant design as accepted by BC Hydro’s “Power Smart” program. Thus, HPGR energy savings have an impact far greater than just the energy savings in the grinding area. A separate report to BC Hydro has confirmed that the use of HPGRs for KSM qualifies for these incentives.

TS6 currently requires potentially large non-refundable 500 kV transmission and generation system reinforcement charges for operations with a Contract Demand of over 150 MVA. This would apply to KSM and the required capital contribution would apply to the entire load, not just the load that exceeds 150 MVA. The Contract Demand (peak load) for KSM is currently estimated to be well above 150 MVA, even considering that energy conservation measures will be implemented. Seabridge currently has an application before BC Hydro for an increase in the Contract Demand from 150 MVA to 200 MVA, with the expectation that the generation reinforcement charges will be set at zero and 500 kV system reinforcement charges will only apply to the additional 50 MVA of Contract Demand, not the entire load, reflecting the state of the current system. However, as this is uncertain, and a determination will not be made until after publication of this PFS, a combustion turbine has been allotted for in the estimates for peaking purposes, as this is far more economical than the application of the currently applicable reinforcement charges (TS6).

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The cost of power for KSM, delivered to the 25 kV bus bars of the Treaty OPC, has been estimated as Cdn$0.062/kWh, including applicable taxes and energy cost savings due to BC Hydro’s Power Smart program. The KSM power cost includes the transmission line losses from the metering point at the Treaty Creek Switching Station, plus Substations No. 1 and No. 2 transformer losses and peaking power cost.

The KSM power cost calculation takes into account reduced rates due to BC Hydro Demand Side Management (DSM) and associated Power Smart initiatives for energy conservation measures designed into new plants (such as using HPGR grinding in lieu of SAG milling). Such measures, as may be certified by BC Hydro, serve to reduce the standard 10% of energy under the two-tier 1823 Rate Schedule that would fall under the costlier Tier 2 category. If HPGR grinding and similar energy conservation measures were not to be implemented, there would not only be greater energy consumption, but the cost of electric power for the entire KSM operation would increase.

Each year on April 1 (the start of their fiscal year), BC Hydro sets new rates that are applied in accordance with the tariffs, subject to BCUC approval.

| 18.10.6 | Treaty Plant Main Substation No. 1

The KSM 287 kV step-down Substation No. 1 will be located at the Treaty OPC and will be constructed and owned by Seabridge in accordance with BC Hydro policy, which is also the most economical solution. This substation is a critical installation for KSM. The substation equipment has been sized based on the latest PFS load list. Redundant transformer capacity was included in the design. The substation will be a GIS switchgear) design, utilizing 138 kV and 287 kV gas insulated circuit breakers and bus bars, allowing a compact design contained in a building adjacent to the Treaty process plant. The circuit breakers will use point-on-wave switching as required by BC Hydro. Connections to transformers will use high-voltage solid dielectric cables.

The substation will include:

| ● | three transformers, each of the three winding type oil filled 75/100/125 MVA, ONAN/ONAF1/ONAF2 step down power transformers, with automatic on-line tap changers

| ● | six 287 kV GIS circuit breakers

| ● | seven 138 kV GIS circuit breakers

| ● | one 287 kV switched reactor for compensation of the incoming 287 kV line, to limit Ferranti effect over voltages

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| ● | two 138 kV switched reactors at the OPC end of the 24 km long, 138 kV cable to compensate for cable capacitance, thus controlling bus voltage

| ● | 25 kV grounding transformers and resistors.

The three transformers included in Substation No.1 will be installed in a concrete vault. They will provide redundancy, allowing one transformer to be out of service. Shipping restrictions to the site were taken into account when sizing the transformers. The 138 kV tertiary windings will be connected to the 24 km long tunnel cables feeding the Mitchell (open pit mine) area Substation No. 2. Space has been allocated for a future fourth power transformer.

Substation No. 1 will also include a line-up of 25 kV metalclad or GIS switchgear, including all 25 kV circuit breakers required for power distribution to the process plant and around the Treaty OPC. The secondary distribution voltage for the Treaty OPC will be 25 kV.

Substation No. 1 does not include harmonic filters. If these are required by harmonic generating plant loads, they would be best located at the process plant near the harmonic sources.

| 18.10.7 | 138 kV Cable

Substation No. 1 will be interconnected with Substation No. 2 by three, 138 kV, single-core, 300 mm2, cross-linked polyethylene (XLPE) solid dielectric power cables suspended from the back (roof) in one of the MTT that will run between the two plant sites. The 300 mm2 (600 kcmils) conductor size quoted is the minimum physical size that vendors typically manufacture at 138 kV (due to the electric field gradient at the conductor). This is a more than adequate capacity to carry any anticipated load, including allowance for the cable charging current. In order to limit induced sheath currents, the cable sheaths will be “cross bonded”, which is the normal design for high-voltage, high-current, single-core cable installations. Adequate (significant) space must be allowed in one of the train tunnels for these cables.

| 18.10.8 | Mitchell Substation No. 2

The 138 to 69 kV - 25 kV Substation No. 2 is also a critical infrastructure for KSM. As an alternative to a standard 138 kV air-insulated outdoor substation, Substation No. 2 is planned to be a GIS installation. This is a very compact design, requiring only a fraction of the space of a conventional air insulated high voltage substation and allows for the total installation to be included in a reinforced concrete building that provides a high degree of protection against geo-hazards such as avalanches. It also eliminates hazardous high-voltage overhead lines in the vicinity of the Mitchell OPC and requires much less plant area. The substation includes:

| ● | two 138 - 69 - 25 kV, 55/73/90 MVA ONAN/ONAF1/ONAF2, oil filled power transformers with automatic on-line tap changers (two, 3 phase units are provided for redundant capacity, with space provided for a third unit to cater to future load growth)

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| ● | six 138 kV GIS circuit breakers and associated bus work

| ● | two switched 138 kV reactors to compensate 138 kV cable capacitance

| ● | six 69 kV GIS circuit breakers connecting to site 69 kV power distribution system to the facilities more remote from the No. 2 Mitchell Substation site

| ● | grounding transformers and resistors

| ● | a line-up of 25 kV metalclad or GIS switchgear for site local power distribution including 25 kV circuit breakers for all local power distribution.

| 18.10.9 | Site Power Distribution

Site power distribution in the mine area from Substation No. 2 will be by 25 kV cables and overhead pole lines locally and by 69 kV overhead pole lines to feed large loads at more distant facilities where modular substations will step the 69 kV down to the local distribution voltage. The relatively long distances and high initial and future pumping loads require 69 kV distribution to transmit the power and limit voltage drop.

| 18.10.10 | Mine Power

Power to the Mitchell open pit itself will be provided by local 25 kV overhead distribution lines. The required pit 25-7.2 kV portable substations (also serving as pit switch-houses), and trailing cables for the 7.2 kV pit mobile electric shovels and drills, are included in the electrical power budget. 7.2 kV to 600 V portable substations are also included for pit dewatering. Similar installations are included in sustaining capital or the Sulphurets and Kerr open pits.

| 18.10.11 | Construction and Standby Power

Modular diesel generator sets will be provided to supply construction power for tunnel driving, camps, temporary water treatment plants, plant construction sites, and other initial construction-related facilities. The capital and operating costs of these facilities plus local distribution including step-down transformers and overhead pole lines have been included in construction indirect costs. Fuel and operating costs for construction power are also accounted for in the construction indirect costs. The power distribution costs for supply and installation of cable and electric panel boards within the various tunnels are included in tunnelling costs.

Any additional costs for moving equipment and fuel to site during the early stages of the KSM construction, either by helicopter or by a glacier road, are included elsewhere in the capital cost estimate and are not in the construction power budget.

The construction generating stations are modular, complete with switchgear, and designed for PLC automatic unattended operation. Environmentally approved double-walled fuel storage tanks and associated piping are included for each power station.

Several of the construction gensets will be retained after initial construction is complete and reconfigured to serve as future standby/emergency generation for the mine, process plant, and accommodation centres. The cost to refurbish construction gensets and reconnect this equipment for standby service in the permanent plant was included in the process plant electrical budget.

The estimates include the purchase rather than rental of construction gensets. The relatively very long KSM construction period will make construction genset rental uneconomic.

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| 18.10.12 | Energy Recovery and Self Generation

There are several opportunities for energy recovery from process flows, as well as power generation from mini-hydro plants, taking advantage of water flows that must otherwise be diverted around the mining operations. As these energy recovery and mini hydro schemes, to a large extent, make use of facilities otherwise required for KSM, they are generally economically attractive and will also reduce the total energy consumption of the KSM operation. The value of the generated power would be at the BC Hydro rate schedule 1823 Tier 2 energy (set at BC Hydro’s marginal cost of generation).

All of the listed energy recovery plants will be located within the KSM mining lease. The energy recovery plants recover energy from process plant flows.

All of the generating plants, similar to small IPP hydroelectric plants, will operate unattended and will be automatically controlled by PLC systems. The locally generated power will be fed into the 25 kV mine distribution power lines.

The generation facilities included in this PFS are summarized in Table 18.8.

Table 18.8 Mini Hydro and Energy Recovery Power Generation

Name | Type | Installed Capacity (kW) | Net Annual Generation (kWh) | Machines
WTP* | Energy Recovery | 9,000 | 23,773,000 | Pelton Turbines
Tailings | Energy Recovery | 1,194 | 7,794,060 | Pumps as Turbines
Mitchell Diversion* | Mini Hydro | 7,500 | 17,638,000 | Pelton Turbine
McTagg Stage 2 Diversion (Years 10 to 15) | Mini Hydro | 8,000 | 32,981,000 | Pelton Turbine
McTagg Stage 3 Diversion* and **(Year 15 onwards) | Mini Hydro | 12,000 | 45,242,000 | Pelton Turbine

| Notes: | *Operation continues after mine closure. **McTagg Stage 3 replaces Stage 2.

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| 18.11 | Treaty OPC and Mine Site Secondary Electrical Power Distribution And Utilization

| 18.11.1 | Mine And Plant Power Consumption

The total mine and process plant annual energy consumption is estimated to be 1,333 GWh based on the Tetra Tech (HPGR) load list for a 130,000 t/d operation valid on average for Years 1 to 5, based on the currently proposed blending of ore. This equates to an average annual load of 152 MW. With a load factor (LF) in the range of 0.85 as typical for a property such as KSM, the peak load (30-minute demand) is estimated as 179 MW. The plant running (normal every day) load is estimated to be 167 MW, again based on norms for this type of mine and milling operation.

The required utility supply will be reduced in the summer and fall by self-generation from energy recovery and mini-hydro plants. During the winter low stream flow conditions, the average self-generation will be almost zero. To prevent the KSM operational demand from exceeding the 150 MW trigger point for generation reinforcement, the proposed peaking combustion turbine, located in at the Treaty OPC, will be operated.

| 18.11.2 | Power Distribution – Treaty Plant Main Substation No. 1

For a discussion of power distribution refer to Clause 18.10.6 Treaty Plant Main Substation No. 1.

Ball Mills

The ball mills as planned were integrated into BC Hydro’s SIS and any changes to these drives will require a new BC Hydro SIS and potentially a new Facilities study.

The Treaty process plant ball mills are major power consumers. Each of the four ball mills (rated 14,000 kW each) will be fed via dedicated 25 kV feeders and step-down transformers to 13.8 kV. The mills will each be equipped with two, 10,000 hp, low-speed “Quadratorque” fixed speed synchronous motors, directly driving mill dual pinions via air clutches, as has been used in the industry for many years.

Step-down to 4.16 kV

The ball mills will be fed at 13.8 kV. Other large fixed speed motors (generally those rated 250 hp and greater) and large variable speed drives (generally those rated over 400 hp) will be fed at 4,160 V. The 4,160 V supply will be derived from 25 kV to 4,160 V outdoor liquid filled step-down transformers. Redundancy will be provided by utilizing sets of two transformers, each feeding a 4,160 V metal clad switchgear line-up with the two line-ups connected by a tire breaker that may be closed if one of the transformers fails or must be taken out of service.

Step-down to 600 V

Motors and other loads below 250 hp will be fed from one of several 600 V systems. Generally, these systems will consist of liquid insulated 25 kV to 600 V step-down transformers, feeding two line-ups of 600 V power distribution centres (with tie breaker), which in turn feed a series of 600 V motor control centres (MCCs). General power and lighting will also be fed from the 600 V system.

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Remote Loads

Remote Treaty OPC loads will be served by 25 kV overhead lines. Examples of remote loads include, fresh water pumping, the TMF return water and seepage pumps and ancillary buildings.

| 18.11.3 | Mitchell Substation No. 2

For a discussion of power distribution refer to Clause 18.10.8, Mitchell Substation No.2.

Power Feed to Pits and Primary Crusher

The Mitchell primary crusher will be fed from the substation by a 25 kV cable. The Mitchell pit overhead power line will be fed from a section cable leading into the substation.

The mining electric shovels and drills will be served at 7.2 kV via portable 25 to 7.2 kV step-down substations fed from the perimeter pit pole line. The estimates include appropriate lengths of trailing cable and couplers. 7.2 kV to 600 V portable step-down substations and trailing cables are also included for pit dewatering.

A 69 kV GIS circuit breaker and cable will feed an overhead pole line supplying remote loads including the truck shop, WTP, Selenium WTP, WSD pumping, explosives facility, permanent camp, and also connecting to the mini hydro and energy recovery power plants. There will be local substations stepping down from 69 kV to the local distribution voltage.

Seabridge Gold Inc. | 18-56 | 219221-01-RPT-001
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| 18.12 | Permanent and Construction Access Roads

The Mine Site and PTMA construction will require a combination of permanent and temporary access roads. Preliminary design work was performed pre-2009 utilizing remote sensing data (LiDAR) and was groundtruthed between 2009 and 2012. Primary access to the mine during construction will be via the CCAR and TCAR, with a winter access road being utilized during the initial construction phase. In total, 9 roads are proposed with varying design criteria that match that anticipated traffic volumes generated at the Mine Site and PTMA during construction and operation. A map outlining the proposed roads is shown in Figure 18.19.

Current proposed permanent access roads include the existing 59 km long resource access route from Highway 37 to the former Eskay Creek Mine and camp facilities. The proposed 35 km long CCAR will commence near the southern limit of this existing road, and extend south then west to the proposed Mine Site.

The Treaty Creek Valley road network provides access to the Treaty OPC, the TMF, and the MTT Saddle Area. It will include a 30 km two-lane access route from Highway 37 to the Treaty OPC, TMF, and east portal of the MTT, and include portions of the TCAR and NTAR.

The current proposed access roads include the:

| ● | Eskay Creek Mine Route

| ● | CCAR

| ● | TCAR

| ● | Lower NTAR (early mine life)

| ● | Upper NTAR (early and mid-mine life)

| ● | Cut-off Ditch Access Road

| ● | TMF Service Roads

| ● | Frank Mackie Winter Access Road (design by Tetra Tech).

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Figure 18.19 Proposed Access Roads Network

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| 18.12.1 | Route Descriptions

The current updated route descriptions, including relocations of the road alignments, are provided in the following sections.

Eskay Creek Mine Access Route

Seabridge plans to use portions of the existing Eskay Creek Mine Access Road, linking Highway 37 to the proposed CCAR.

This road was constructed in 1993 to provide access to Homestake Canada Inc. Eskay Creek Mine. In 2002 the mine was assumed by Barrick Gold until its closure in 2008. The property is now under option with Skeena Resources Ltd. The road commences at Highway 37, south of the Bob Quinn Forest Service Road, and follows the Iskut River Valley west for approximately 38 km to the crossing of Volcano Creek. The road was originally designed as a single lane, 5 m wide gravel road, with a nominal design speed of 60 km/h. Substantial portions are built to a nominal 8 m wide (double-lane standard), providing ample passing opportunities.

An overview evaluation of the road condition and its suitability for Seabridge’s requirements was conducted; findings are summarized in McElhanney (2013).

Coulter Creek Access Road

The CCAR will be constructed as a single lane (6 m surface), radio-assisted road with pullouts.

Heading southwest from near the end of the existing Eskay Creek Mine Access road (approximately 59 km off of Highway 37), this road will follow an existing mine access road for approximately three or more kilometres towards Tom MacKay Lake. It will then descend out of the alpine meadows, along the height of land between Coulter Creek and the Unuk River.

The proposed three-span bridge crossing of the Unuk River will be 88 m in length. The Unuk River is a major crossing and will need to meet the requirements of the Navigable Waters Protection Act. Beyond the Unuk River, the route traverses a short section of low-lying wet and swampy areas and then starts to climb steeply through a series of switchbacks into the Sulphurets Valley and canyon. The alignment extends to the HDS WTP where McElhanney’s road design (and Special Use Permit [SUP] boundary) ends. Beyond this point, the road design is determined by the mine development, and is the responsibility of MMTS.

Treaty Creek Access Road (HWY 37 to KM 17.9)

The TCAR will leave Highway 37 approximately 19 km south of Bell II, and head west. It will be constructed as a two lane (8 m surface) all-season road to the junction of the Treaty Creek and North Treaty Upper road at km 17.9.

Meetings were held between McElhanney, Seabridge, and the provincial Ministry of Transportation and Infrastructure (MOTI) to discuss and establish a set of design criteria for the proposed intersection at the Highway 37/TCAR location. Functional design work has been completed based on the criteria.

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Initially, the TCAR will follow a former forestry access road. A three-span 119 m long bridge is proposed for the crossing of the Bell-Irving River. This is a major river and will need to meet the requirements of the Navigable Waters Protection Act. The proposed road follows the north side of the Treaty Creek Valley. It will generally be located between the flatter riparian zone below and the steeper avalanche-prone terrain on the north slope. The TCAR will continue as a double-lane road further west to a future intersection with NTAR. Heading west from there, the TCAR will transition into a single-lane road leading to the MTT saddle area.

Treaty Creek Access Road (KM 17.9 To Tunnel Saddle Access Portal)

Beyond km 17.9 road intersection and heading west up the Treaty Creek Valley, the TCAR will provide construction period access.

North Treaty Creek Access Roads

There are currently three permanent access road alignments proposed within the North Treaty/Teigen Creek valley. They shall be referred to as the lower NTAR, the upper NTAR, and the Cut-off Ditch Access Road.

Lower NTAR

Earlier access can be obtained by constructing the lower NTAR. This will leave the TCAR at approximately km 16.9 and follow the lower valley. The lower NTAR will be quicker to build. It will result in a slightly shorter haul distance between Highway 37 and the Treaty OPC and TMF, and with generally flatter grades. This road would be used for approximately the first half of the mine life, until such time as it is necessary to construct the southeast tailings dam. The primary purpose of the NTAR is to provide early access to the lower valley for the construction of the tailings dam(s). Eventually, the north section of this road would be buried by the southeast tailings dam.

Upper NTAR

The upper NTAR will leave the TCAR at approximately km 17.9. It will traverse approximately 12 km north from the TCAR to the Treaty OPC and TMF. The road would then parallel the proposed drainage cut-off ditch, which will divert drainage off the west slope of the valley, north to the Teigen Creek Valley.

Cut-off Ditch Access Road

The proposed cut-off ditch access road is to provide construction and maintenance access only. The power transmission line to the process plant will follow this route.

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North Treaty/Teigen TMF Service Roads

Access to the Tailings Management Facility will be provided by approximately 28.4 km of 6 m wide service roads (with pullouts). These roads will provide access to the east side of the North Treaty and Teigen Creek valleys, including a water well. The TMF service roads include the following segments:

| ● | South Teigen Road 11 (5.5 km)

| ● | South Teigen Road 11a (3.15 km)

| ● | South Teigen Road 12 (9.1 km)

| ● | South Teigen Road 12a (0.3 km)

| ● | Upper Dam Road 12b (1.2 km)

| ● | South Teigen Road 15 (3.7 km)

| ● | South Teigen Road 15a (2.7 km)

| ● | Water Well Access Road (2.7 km).

| 18.12.2 | Winter Access Road

A Winter Access Road is proposed from a laydown area near the former Granduc Mine to the KSM Site. The approximate alignment for the proposed Winter Access Road is depicted in Figure 18.20. An evaluation of the route was completed by Tetra Tech (Tetra Tech, 2012).

The suggested route will be approximately 38.4 km long. It appears that as much as 32.8 km of the road will be constructed on the glaciers. Although the topographic data is not very precise, the bulk of the route appears to have grades of 4% to 6%. Steeper grades upwards of 30% exist at the toe of the Berendon Glacier and on the small side glacier that allows access onto the Frank Mackie Glacier from the Berendon Glacier. There are also steep sections with grades of up to 15% near the crest of the Frank Mackie Glacier. The total vertical variation is roughly 1,020 m (3,350 ft) between the Granduc Mine area and the crest of the Frank Mackie Glacier.

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Figure 18.20 Proposed Winter Glacier Access Route

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| 18.12.3 | Road Design Requirements

The KSM Site access roads are classified as resource development roads.

The Eskay Creek Mine Road and CCAR will be maintained for the life of the mine to support the mine development, transport of oversize loads, and to provide alternate emergency access. However, these roads will only be used seasonally, and not used during winter months.

The CCAR will be a single-lane (6 m surface) radio-assisted road with turnouts and widenings to allow the largest vehicles and loads access to the Mine Site. The CCAR would have some sections with sustained maximum grades of 12%. Design speeds vary greatly, in large part controlled by the terrain.

The proposed TCAR to km 17.9, and the connecting upper and lower NTARs, will be required for permanent access to the Treaty OPC and TMF, and to the Mine Site via the MTT tunnels. These will be two-lane roads (8 m finished surface), capable of carrying highway legal axle loading year-round. The roads will provide access for supplies, equipment, and crew transport, and be used for hauling concentrate to Highway 37.

Alignment controls such as maximum 10% sustained grades (11% short pitch), and minimum 100+ m radius horizontal curves are utilized for the higher-traffic volumes anticipated on this route. Appropriate vertical profile crest and sag curve “K” values are applied. Except for a few control sections, the nominal minimum design speeds for these sections of road is 50 km/h, and maximum 60 km/h where feasible.

All bridges will be designed to BC Forest Service L100 loading (90,680 kg gross vehicle weight [GVW]) and minimum 1.5 m clearance above the estimated 100-year flood level (Q100). Select structures must meet additional requirements, as prescribed by the Navigable Waters Act. All bridges, including those on the TCAR, will be single lane.

Major culverts have been designed to pass the estimated 100-year flood level (Q100) with no headwater.

| 18.12.4 | Design Progress

Early Design Work (Pre-2009)

Utilizing LiDAR survey data acquired in summer 2008 and fall 2011, and the resulting digital elevation models developed, the preferred preliminary access routes identified by Seabridge and McElhanney were prepared. The preliminary road alignments were subsequently located in the field using GPS, and marked with flagging. The objective was to locate and map the most appropriate road alignment for each route based on design standards established by the team. Road alignments and cross sections took into consideration the requirements for both construction and operation phases.

The routes were assessed in the field and adjusted as deemed appropriate. Often several preliminary lines were investigated in order to achieve the preferred road location. Selecting the ultimate road locations was an iterative process involving both field and office design. Based on the preliminary layout, terrain information was gathered, along with bridge and major culvert crossing information. The originally flagged centerline provided a base for follow-up environmental and geotechnical assessments.

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Based on the field reconnaissance, design standards, and associated surveys and preliminary assessments/input by other sub-consultants; preliminary road design plans and profiles, conceptual bridge and stream crossing structure designs, and construction cost estimates were prepared. Engineering assessments were conducted in conjunction with available geotechnical and environmental studies of all proposed routes. The field reconnaissance and bridge site surveys confirmed the accuracy of the LiDAR data.

Design Work (2009-2012)

From 2009 through 2012, consultants BGC and Rescan (now ERM) conducted further geotechnical and environmental assessments, respectively, on the proposed, and altered routes. Where appropriate, McElhanney’s QP (Mr. Bob Parolin) accompanied these consultants in the field to make joint determinations with respect to the most appropriate locations for specific sections of road. McElhanney worked with these consultants to optimize the road locations and designs.

All proposed final road locations were marked with survey flagging. Flagging was marked with survey crew and date information (black felt marker), and locations identified by real-time kinematic (RTK)-GPS survey methods. Select field station references are now indicated on the road plan/profile design drawings for cross reference.

Work included gathering of detailed information to be utilized in refining the design(s), including soils, vegetation, potential borrow/waste areas, drainage culvert requirements, and other relevant information. Stream crossing surveys were completed on “smaller” tributaries. Generally, this includes any stream with an estimated 100-year peak flow of 6.0 m3/s or greater. Details for all such structures were completed to satisfy the requirements of the BC Ministry of Forests, Lands and Natural Resource Operations (MFLNRO) for the SUP applications. Preliminary stream crossing structure designs have been completed for all sites requiring bridges or major culverts.

Design Updates (post-2012)

Additional field work was conducted in 2012 along the proposed access routes by McElhanney, BGC and ERM. New information was incorporated to optimize the road and structure designs. Horizontal and vertical alignments were modified to best meet the environmental, geotechnical and archaeological concerns and requirements. During 2012, work was completed to locate potential borrow and waste sites, at appropriate locations to accommodate road grade construction requirements along the access corridors. Provision was also made to identify areas of potential gravel sources for road construction and surfacing materials.

Log landing locations were identified for decking of timber felled during right-of-way clearing operations. Log landings were located, and spaced, as appropriate for logging/skidding operations. Timber maturities/volumes, etc. were considered in establishing proposed landing locations.

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The proposed right-of-way (clearing) boundaries now defined on the design drawings are minimum 30 m wide, expanded to include proposed borrow and waste areas, and log landings as described above. The approved SUP boundary limits extend a minimum of 37.5 m either side of proposed road design centerline (total 75 m width), widened as required to incorporate additional areas as otherwise defined. The intention is that this will provide some flexibility in adjusting the design or construction methodology as may be required due to actual field conditions encountered, without requiring multiple amendment to the SUP during the construction period.

The potential for ML/ARD has been assessed for all access road rights-of-way. Additional assessments were conducted in late 2014 for km 0 to 7 of the CCAR which had been flagged by government agencies as an area of special concern.

The SUPs for road construction associated with the site access construction were granted by the Provincial MFLNRO Road Division office on September 27, 2014. There is a requirement to provide a security, payable to MFLNRO, prior to commencing construction.

Access road surveys, designs and drawings were prepared in conformance with standards provided in the then most current version of the BC government Forest Service Engineering Manual (November 29, 2012). Detailed engineering of specific slope stability measures will be subject to review by the geo-technical engineer(s), immediately in advance of, and during construction activities.

Bridge and major culvert structure site plan surveys, designs and general arrangement drawings have been prepared in accordance with MFLNRO Road Division requirements and current industry standards. General arrangement design drawings have been signed and sealed independently by a professional engineer registered to practice in BC. Detailed structure design details will need to be completed in advance of construction.

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| 18.13 | Logistics

The KSM Site is currently accessible by helicopter only. Helicopter support will be used initially to transport equipment, supplies, and personnel prior to completion of the access pioneering roads, and to support winter construction work ongoing as necessary on the Mine Site since CCAR is a seasonal road (closed winters) and MTT would not be operational until the end of the construction duration.

Once surface routes of TCAR, CCAR and the Winter Access Road are established, construction equipment, materials, personnel and consumables will be imported to various work fronts located across the KSM site through these routes and dependence on helicopter support will reduce significantly.

Copper concentrate will be transported from the KSM site by trucks to a deep-water port facility in Stewart, BC, and then loaded onto oceangoing vessels. Two full service ports exist at Stewart, each with roll-on/roll-off freight handling capacity and either are presently, or would by the time operations begin, be capable of concentrate storage and handling to ship loading.

The port is at the head of the Portland Canal, which is a 150 km fjord that is the northernmost ice-free port in North America. The port is accessible via truck on Highway 37A; however, there is no direct rail service. Concentrates from other northern BC mines are currently shipped from this port. In addition, there is interest from other operations in the region for concentrate handling services at the port.

For the purposes of this PFS, Tetra Tech calculated that the copper concentrates will be shipped in bulk, and that the annual output for the initial 10 years will be approximately 350,000 t copper concentrate (dry tonne).

Molybdenum concentrate will be transported in bags from the KSM site via trucks to the port of Prince Rupert. The bags will be transferred from the trucks to containers and then delivered to Fairview Terminal for ultimate loading onto an oceangoing vessel.

It was assumed that the processed molybdenum will be loaded in 1 t bags for transport purposes, and that the annual output will be approximately 1,155 t molybdenum (LOM).

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| 18.14 | Preliminary Construction Execution Plan

The Construction Execution Plan describes how the KSM Mine could be constructed. It is a plan in the preliminary planning stage and briefly defines the construction elements required to successfully execute construction management for KSM.

| 18.14.1 | Introduction

The KSM Mine requires six years to construct. The construction scope is intended to meet the following key objectives:

| ● | deliver an optimized, safe, and environmentally compliant construction in accordance with the systems and procedures in place

| ● | perform construction activities safely, striving for zero recordable accidents

| ● | the construction is carried out in accordance with the Impact Benefit Agreements (IBA) that are in place with the First and Treaty Nations

| ● | ensure that regulations, license agreements, applicable specifications, and standards are met.

| ● | complete construction within the agreed schedule, not exceeding the budget, and delivering the full scope as described in the construction authorization.

| 18.14.2 | Early Works Plan

An Early Works Plan will be developed to ensure certain key infrastructure (e.g., construction camps) and support services (e.g., catering) are in place early during construction and functioning efficiently for a successful construction program.

The following planning and field construction focus areas must be addressed in the Early Works Plan:

Planning

| ● | permit review and renewal plan

| ● | construction procedures

| ● | staffing, recruiting and labour relations plan, including commitments in accordance with the IBAs that are in place with the First and Treaty Nations

| ● | contracting strategy and plan, including commitments in accordance with the IBAs that are in place with the First and Treaty Nations – vetted and approved by the Owner

| ● | site access plan – pioneer roads, bridges, followed by completed permanent roads

| ● | health, safety and security (HS&S) management plan and manual

| ● | site and camp rules and regulations plan

| ● | environmental and cultural sensitivity awareness training plan

| ● | health and hygiene program

| ● | site safety and security orientation program

| ● | geohazards and avalanche monitoring response plan

| ● | logistics supply and materials management plan for early material requirements, including helicopter support

| ● | employee transportation plan for early construction program; air and ground planning required

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| ● | environmental management plan to manage sediment control, waste, spills, fueling, etc. and wildlife management plan for early construction activities

| ● | community relations plan

| ● | quality management system

| ● | safety and emergency response plans including content related to medical facilities and medical attention, emergency medevac, etc. Final Level IV resource loaded construction schedule

| ● | Final development execution plan

Field Construction

| ● | establish explosive supply storage and controls

| ● | identifying and proving borrow pits

| ● | sourcing road materials and aggregates; setting up crushing and screening facilities

| ● | build pioneering roads along CCAR and TCAR alignments to establish road access to the Mine Site and Treaty OPC

| ● | establish aggregate plant, aggregate wash plant, batch plant installation and supply of cement and aggregates

| ● | install asphalt plant and supply of asphalt

| ● | develop fuel supply and storage locations on site immediately upon achievement of road access build construction camps

| ● | establish temporary construction power – standalone power supply systems (gensets) in containers with fuel systems; build TWTP ponds and muck pads, and install TWTP's where tunnelling is on the critical path (e.g., MTT)

| 18.14.3 | Construction Scope

The construction scope outlined in this section, summarizes the main infrastructure items constructed as permanent facilities or activities required to support permanent constructions within and surrounding the Mine Site, Mitchell OPC, MTT and PTMA:

| ● | Mine Site:

| - | Frank Mackie Glacier Road (Winter Access Road)

| - | upgrades to the existing Eskay Creek Mine Access Road

| - | CCAR (35 km, permanent access road)

| - | Mitchell OPC (primary crushing at a peak of 10,000 t/h)

| - | WSF (WSD crest built to elevation 716 masl) and ancillary facilities

| - | HDS WTP with five large clarifiers and sludge storage to initially process up to 5.4 m3/s

| - | six TWTPs and associated muck piles and treatment ponds

| - | surface water diversion tunnels (MDT, MTDT and MVDT)

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| - | power distribution comprising construction gensets, overhead transmission lines, power distribution, and substation

| - | logging, site clearing and grubbing (overburden, soils stockpiles and large woody debris for future reclamation purposes); rough/finish grading; structural excavations and fills; foundations; steel erection; architectural, mechanical, electrical and instrumentation works

| - | Mine Site ancillary buildings and infrastructure such as camps (permanent and temporary), fuel storage yard, sludge storage building, diversion ditches and bypass pipelines, material handling, temporary truck assembly yard, energy recovery plants

| ● | MTT:

| - | two 22.7 km long tunnels plus ancillary excavations

| - | train system for ore transport at a peak of 10,000 t/h

| - | two 15,000 t ore bins and associated transfer conveyor from the Mitchell OPC

| - | power distribution infrastructure

| - | train maintenance building

| - | train loading/unloading facilities

| ● | PTMA and Saddle Area:

| - | TCAR (33 km, permanent access road)

| - | TMF (North Dam built to 930 masl, and Splitter and Saddle dams built to 935 masl, with a fully lined and drained basin for placement of CIL tailing)

| - | COS and transfer conveyors

| - | process plant built for 130,000 t/d average throughput, including concentrate storage and load out

| - | two TWTPs and associated muck piles and treatment ponds

| - | power distribution comprising construction gensets, overhead transmission lines, power distribution, and substations

| - | logging; site clearing and grubbing (overburden, soils stockpiles and large woody debris for future reclamation purposes); rough/finish grading; structural excavations and fills; foundations; steel erection; architectural, mechanical, electrical, and instrumentation works

| - | PTMA ancillary buildings and infrastructure such as camps (permanent and temporary), cold storage, fuel storage yard, diversion ditches and pipelines, and material handling

| ● | Infrastructure (major, both on- and off-site):

| - | concentrate storage and loading at the Port of Stewart

| - | switching station at TCAR turn off from Highway 37(by BC Hydro)

| - | 287 kV overhead power line from switching station to PTMA

| - | fibre optics along power line right-of-way and tie-ins

| - | Highway 37 marshalling yard and Highway 37 turnoff, including site security infrastructure.

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| 18.14.4 | Construction Schedule

The 2016 PFS construction schedule was compiled in accordance with the AACE® International (AACE®) recommended scheduling guidelines (level of detail is at Level 2), with a Class 4 definition. The construction schedule is estimated to be six years and has been designed to accommodate major seasonal and environmental constraints.

Critical path consists of pioneering roads along the alignments of the principal access arteries to the construction areas (CCAR and TCAR), MTT, and the train system that will connect the Mine Site with the PTMA. Prior to completion of the KSM site access pioneering roads, helicopter support will be utilized to support early construction activities at multiple road construction headings. The strategy is to establish site access pioneer roads as early as possible to reduce heli-support costs exclusively for Mitchell Valley construction activity during winter months. Upon completion of the access roads to full width, major equipment and materials can be transported to site via ground freight.

Major site infrastructure such as the WSF and the TMF could potentially be on the critical path should the MTT tunneling duration be shortened. A preliminary construction schedule has been developed with a start date for the construction program assumed for mid-Year -6. Contractors to begin construction on the CCAR and TCAR construction would assume to mobilize for mid-Year -6.

Mine Site pioneering begins with the development of the site access roads to the major infrastructure pads such as HDS WTP area, WSF, Mitchell OPC, MTT portals, CCAR, batch plant, TWTPs, accommodation complexes and powder storage, initially from where the Frank Mackie Glacier Road ends. Early works material and equipment will mobilize on this Winter Access Road and the major equipment, general construction materials, and heavy earth moving equipment will mobilize via the CCAR. The Treaty OPC will utilize the TCAR to transport all material and equipment for PTMA construction.

The construction duration is estimated at 66 months. The high-level schedule is shown in Figure 18.21.

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Figure 18.21 Construction Schedule Summary (Level 1)

Source: Tetra Tech

| 18.14.5 | Engineering and Procurement

Engineering and procurement activity will be managed by teams of professionals who will report up through the EPCM contractor's directorate. The Engineering Team will provide the required drawings, specifications, and documents to the Procurement Team in order to purchase all equipment and materials for the construction, and to allow field construction of the scope to the design intent. The EPCM contractor’s scope will include process facility and infrastructure engineering, including managing specialty contractors for major dam and tunnel designs. Mine designs will be developed and delivered by the Owner's Team.

The Procurement Team will receive the engineering documentation and obtain multiple quotations that meet engineering specifications and provide a purchase recommendation to EPCM director. After EPCM director approval, the Procurement Team will purchase equipment and materials and arrange all logistics to deliver the items to the construction site ready for installation. The Procurement Team will also be responsible for establishing service contracts for engineering and field construction services.

Both the Engineering and Procurement teams will be including commitments in accordance with the IBAs that are in place with the First and Treaty Nations.

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| 18.14.6 | Construction Management

The Construction Management Team will be responsible for the management of all activities related to the construction management scope that includes all construction activity in the mine, process, and infrastructure areas (on site and off site). Mining activity, environmental monitoring and reporting and community affairs, which will be the accountability of the Owner's Team.

The Construction Management Team will oversee the installation of all materials and equipment according to engineering and manufacturers specifications and build the facilities to satisfy the design intent and be fully operable. The Construction Management Team is also accountable for construction activity and the construction site until hand over to the Owner following dry commissioning.

| 18.14.7 | Construction Supervision and Contractor Management

The objective of all site construction activities is the timely and cost-effective completion of the construction facilities in a safe manner to the design intent and required standards in accordance with schedule. Construction supervision staff, while ensuring that standards are maintained, will provide all oversight management to contractors in achieving this objective.

The Contracts Management Group, which falls under the responsibilities of the site procurement manager, will use an integrated data management system to track contractor invoicing, changes, and requests for information (RFIs). The EPCM contractor will develop a comprehensive set of procedures, in conjunction with and approved by the Owner. These procedures will outline the requirements for the execution of the administrative activities.

| 18.14.8 | Contracting Packaging and Strategy Overview

The preliminary construction strategy includes dividing the construction into contract packages including commitments in accordance with the IBAs that are in place with the First and Treaty Nations. During the contractor expression of interest and pre-qualifications phase and during the advancement of detailed engineering, the contract packages will be combined to reduce the total number of contracts and form a final contracting strategy for the construction.

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| 18.14.9 | Site Organization Structure

The EPCM site organization structure has been developed to provide a balanced combination of senior managers, area managers, engineers, superintendents, and discipline specialists, to provide the Owner and contractors continuous support during the installation period. A high-level organizational chart is provided in Figure 18.22.

Figure 18.22 EPCM Organizational Chart

Source: Tetra Tech

The site organization and staffing plan has been designed by work type (e.g., engineering vs. cost controls) with the geographical constraints of a large construction site incorporated.

Each of the two major construction sites will have a dedicated health and safety manager and multiple health and safety representatives to assist contractors with the daily issues and training requirements.

| 18.14.10 | Environmental and Community Affairs

Environmental and community affairs during construction will be managed exclusively by the Owner's Team to maintain independency from the EPCM Team. Environmental knowledge and community relationships have been developed through historical activity at the construction site and these relationships must continue to be managed appropriately in the context of regulatory permits granted and the societal expectations that have been expressed to the Owner’s Team. These activities will continue to be of paramount importance to the Owner well beyond the construction period, thus are best addressed by the mine owning entity that will have presence throughout the mine life.

The cultural awareness training program will identify and provide an overview of the various Indigenous groups who have an interest in the development, focusing on their rights as it pertains to their traditional use of the natural resources of the area. Contractual obligations negotiated between the Owner and the various groups as components of Benefit Agreements will also be reviewed at a very high level.

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| 18.14.11 | Pre-commissioning/Commissioning

The commissioning period starts in any specific work area after all materials and equipment have been installed to design specification and the EPCM contractor certifies installation complete and hands the area over to the commissioning team. For the purposes of this PFS update, commissioning starts after equipment or material installation for a system or work area is complete and ends when ore starts to be processed to yield a revenue stream (i.e. the battery limit between Year -1 and Year 1). In this PFS the EPCM contractor's scope includes commissioning through dry commissioning when work areas are handed over to the Owner's commissioning team. The Owner's commissioning team executes wet and process commissioning with select operating staff and professionals that are separate in reporting line and accountability from the operations staff. When these phases of commissioning are complete, they will be handed over to Owner's operations staff who will operate the facilities through ramp up in Year 1 and on to normal operation.

| 18.15 | Owner’s Implementation Plan

The KSM Mine will be constructed as outlined in Section 18.14 and in the time frames indicated in the construction schedule in Section 18.14.4. It is the Owner's responsibility to attain, and renew when necessary, all environmental and operating permits allowing site access road development, mine construction, and all mine operations for KSM.

This Owner's implementation plan described herein attempts to provide a preliminary outline to the key responsibilities and actions the Owner's Team will take, including interaction with EPCM contractors during the construction stage and commitments in accordance with the IBAs that are in place with the First and Treaty Nations. It is assumed KSM will be developed as a joint venture (JV) or consortium of two or more companies that will form a partnership to build and operate the KSM Mine. A JV organization would allow KSM's partners to reduce risk and spread capital expense. The “Owner” referenced in this section is synonymous with this JV organization. It is further assumed that the structure developed will assign decision making authority to the majority stakeholder to eliminate bureaucracy and streamline development and mine production decisions. The KSM Mine will therefore have its own operating structure and reporting line through the JV partnership, maintaining its own profit and loss accountability to the JV partners. The Owner's organizational structure will have a KSM president with multiple reporting lines through a six-layer organization. Site based reporting lines to the president comprise construction, mine, and process with on-site administrative functional support as necessary to enable the Owner’s Team’s success. Additionally, off-site business and external relations functions would also report to the KSM president.

A central office in Vancouver is not anticipated. Instead, satellite offices will be located in Terrace, Smithers and Stewart, BC to facilitate support functions sufficiently close to the construction site to provide effective support. The implementation plan described in this section highlights some key tasks required for execution by the Owner's Team over the course of construction. There are two initial critical tasks for the Owner's Team, starting with the identification and hiring of the KSM president, who will initially select a team, who in turn will do the same for their respective teams. This process is expected to be repeated throughout the course of construction until the entire enterprise organization has been built, while directing and supporting construction in various roles.

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The second initial critical Owner's Team task is the engagement of an EPCM contractor early in the development schedule to drive the majority of the scope that resides outside of the Owner's direct responsibility. The type of contractual arrangement between the Owner and EPCM contractor has not been established, as it relies on the strategy that will be developed after forming the JV and may be influenced heavily by the operating style of the JV partners.

The Owner will manage any early engineering work required to prepare design documents that support permit applications or renewals and compliance reports for permits issued by the Province of British Columbia and the Government of Canada. Site road access permits, construction camps approvals, and limited site development permits have been obtained. Additional permits will be required by the KSM Mine to ensure the completion of construction and the initiation of long term operations. During construction, the Owner will be responsible for:

| ● | mine development/construction including pre-stripping

| ● | supervision of mine fleet assembly

| ● | all environmental baseline monitoring, permitting and compliance

| ● | operation of temporary water treatment and sewage treatment plants

| ● | community and governmental relations

| ● | Treaty and First Nation relations

| ● | competitively bidding, adjudication and award of EPCM

| ● | Owner's Team recruitment

| ● | training of operating personnel for the Mitchell pre-mining phases

| ● | medical support

| ● | verification surveying for measurement and payment

| ● | on boarding all G&A staff for both on-site and off-site positions in advance of commissioning to assist as part of the Commissioning Team and to develop process and procedures for each department/function to efficiently support operations

| ● | all personnel required for ultimate mine and plant start-up and operations.

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The Owner will recruit and train technical operations and administrative staff to work in the following locations:

| ● | Treaty OPC Operations and Water Management – process plant and train operations, maintenance, TMF operations, security, and administrative personnel

| ● | Mine Site and Water Management – operations, maintenance, security and warehousing personnel

| ● | Smithers Office – proposed to be developed to service all of KSM's needs for external relations comprising governmental affairs, environmental management, permitting and compliance, public and community relations and communications, First Nations and Treaty Nation relations

| ● | Terrace Office – proposed as a business centre where home office support will be based for these administrative functions: supply chain and logistics, human resources, IT, accounting functions, tax, business analysis, legal and audit. Health, safety and loss prevention may have occasional presence in this office, but will be primarily based on site to support ongoing operations as they develop

| ● | Stewart Port Site – management of deliveries and security for incoming construction equipment/materials and outgoing concentrate shipments.

A conceptual onboarding plan for specific G&A functions will be developed prior to turnover of constructed areas of the site from EPCM to the Owner's Team and is programmed to be well in advance of the turnover. This will allow sufficient time for the development of internal KSM Mine processes and procedures as a means to facilitate a smooth mine start up. The early onboarding plan is intended to cover gaps in service areas that may not have been detected in this early stage of design.

The time sequences for key Owner activities by year are outlined in Table 18.9. Note that this task list assumes that a FS has been completed or is running concurrently with the construction start and that either full or conditional/partial construction funding will be granted by the Owner ahead of the construction start.

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Table 18.9 Owner's Key Activities by Year

Year | Activities
(Year -6) | ● complete FS update by Q2 Year -6 ● renew early works permits, as necessary, to support initial construction ● recruit KSM president and project directors ● recruit and onboard key department leads for process, human resources, business manager, and environment and community relations ● initiate training programs required to meet First and Treaty Nation employment objectives ● establish a business office in Terrace, BC ● implement site environmental monitoring program during construction ● recruit and install at site environmental monitors, TWTP and sewage plant operators, field coordinators to support early works construction mainly for road building and camp construction ● competitively bid, adjudicate, and award EPCM services ● in conjunction with the EPCM contractor, develop a detailed execution plan for the construction leveraging all previous engineering, construction planning, and environmental permitting work. ● establish project governance between Owner and EPCM teams ● release early works contracts for road and bridge construction (critical path) ● obtain permitting for use of Frank Mackie Glacier Road during winter, as required ● initiate fish habitat compensation construction
(Year -5) | ● manage EPCM contractor focussing on detail engineering and long-lead procurement activities off site, and early works construction on site ● augment site Owner's Team adding environmental services, survey (measurement and payment), medical services ● expand Smithers, BC office to accommodate larger office headcount ● continue Mine Site development ● establish road to Sulphurets quarry to support WSD construction
(Year -4) | ● finalize concentrate smelting contract terms ● finalize detailed engineering work for early phase construction activities, initiate detailed design for the remainder of project scope ● augment site Owner's Team in the following areas: process operations, site administration, security, site project management, human resources and environmental services; process operations staff are expected initially to reside in the EPCM contractor's office to guide detailed design and process equipment selection ● continue to manage EPCM contractor focussing on detail engineering and procurement activities off site and early works construction on site; procurement focussing on large equipment purchases required at site in Year -3 and beyond ● continue mine site development ● start delivery Sulphurets rock to WSD for construction
(Year -3) | ● complete final detailed design for all remaining project scope and finalize all equipment purchases. ● initiate enterprise computer systems set up in off-site offices ● continue to manage EPCM contractor whose focus is now shifted mainly to field activity ● initiate business readiness planning leveraging Owner's Team resources, working collaboratively with the EPCM contractor ● augment off-site human resources team ● continue Mine Site development ● continue delivery of Sulphurets rock to WSD for construction

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Year | Activities
(Year -2) | ● continue to manage EPCM contractor whose focus is solely on field activity ● establish satellite office and recruit port staff for facility at Stewart, BC ● expand Terrace, BC office footprint to full size necessary to fully support mine and process operations on site ● recruit all operations staff for HDS WTP to participate in wet commissioning of the facility following dry commissioning and handover by EPCM, initiate plant start up and operations ● continue Mine Site development ● continue delivery of Sulphurets rock to WSD for construction
(Year -1) | ● recruit and on board all remaining positions for the process plant, metallurgical laboratory and TMF; plant operators are highest priority and TMF staff will be on boarded toward year end ● complete recruitment for mine operations team ● perform a significant amount of operator training in preparation for operations start up ● oversight of OEM assembly of major mine fleet equipment ● after completion of WSD, initiate ore mining in Mitchell open pit Phase I ● delivery of first ore to Mitchell OPC ● fully execute wet commissioning following dry commissioning and handover by EPCM of the process plant and all infrastructure ● initiate ore transfer through the MTT controlled by process operations ● introduce first ore to Treaty OPC and begin process plant ramp up
Operations Year 1 | ● complete recruitment and on boarding for remaining G&A and process operations staff during process plant ramp up ● complete oversight of OEM assembly of major mine fleet equipment ● produce first copper concentrate and doré at the Treaty OPC ● initiate sand tailings production at the TMF and train employees on procedures for this long term dam construction effort ● achieve commercial production ● ramp up to full scale mining and primary crusher operations at the Mitchell OPC from Phase I mine development stages at Mitchell and Sulphurets ● shipment of copper concentrate from the port facility in Stewart

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19.0 | Market Studies and Contracts

Seabridge engaged Neil Sheldon and Associates Ltd. (NSA) to provide opinion reports on marketing inputs for the 2016 PFS and review the 2019 copper-gold concentrate market and related concentrate treatment charges, excluding off-site transportation costs. The information and options in this section come from the opinion reports. No smelter contracts are currently in place or being negotiated. All currency amounts used in this section are in US dollars, unless otherwise specified.

19.1 | Copper Concentrate

19.1.1 | Marketability

When considering the marketability of copper concentrates, quality and quantity are determining factors. There is considerable variation in the quality of concentrates and the requirements of various smelters do vary; such variation relates to the technical abilities of the smelter and its overall concentrate feed and blend.

Ideally, smelters prefer to blend their feed with approximately 30% copper and similar amounts of iron and sulphur. In the last several years, however, the grade of some of the major high-grade suppliers has been dropping. At the same time, many new suppliers tend to blend copper-gold concentrates with copper content in the low to mid 20% range. Consequently, the market has seen the blend for most smelters drop to a copper level of 27 to 28%. Apart from the level of copper, iron, and sulphur, other key elements in determining concentrate salability include the levels of gold and silver content, as well as any impurities.

Based on the impurity levels projected by Tetra Tech (using the test results completed to date – see Table 17.3), concentrates are relatively clean. Depending on the market situation at the time of contract negotiations, penalties will likely be minimal, if at all applicable. Some smelters, such as in Japan, South Korea, and Europe, are expected to have more interest in copper concentrates with high gold content.

19.1.2 | Smelting Terms

Copper Concentrate Smelting Market

Copper concentrates account for approximately four-fifths of total newly-mined copper production, with the balance of output coming from solvent extraction and electrowinning copper cathode and other copper-bearing by-products.

Concentrate supply started to increase over 2013 and is expected to continue to increase towards the end of this decade as a result of new developments and announced expansions of operating mines.

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A significant portion of copper concentrate is processed by integrated smelters—captive plants that are vertically integrated with mines through ownership. However, an increasing annual volume of global copper concentrate is treated by custom smelters that generally are not integrated, although there is, in many cases, investment ownership in mines. Custom smelters have increased their overall smelting market share from 30% in 1980 to approximately 50% recently.

Over the last two decades there has been a significant expansion of smelting and refining capacity, particularly in India and China. The Chinese smelting industry has increased imports, as limited domestic mine capacity could not meet demand. This trend has been a key determinant in world concentrate supply/demand balances.

The copper concentrate market has seen significant structural imbalances in the recent years between mine production and smelting capacities. Recently there have been significant increases in smelter treatment charges and refining charges (TCs/RCs). The balance of supply and demand for concentrates is set by the whole of the concentrate output of the mining industry and by the availability of capacity across the smelting industry. The availability of custom concentrates, relative to smelting capacity, should, in theory, be the ultimate determinant of terms for custom treatment of concentrates.

19.1.3 | Copper Concentrates Contracts and Terms

The concentrate market is basically split into two types of contracts. First, there are long-term off-take contracts between mines and smelters that reflect, in general, the annual concentrate supply and demand balance. Second, there is spot or short-term business primarily between mines and traders and, on a much smaller scale, between mines and smelters. By its nature, such business is much more volatile and there is considerable variation in spot TCs/RCs, not only annually, but over each year.

Current and Future Terms

NSA suggests that annual benchmark numbers are beginning to reflect a move towards sustainable long-term numbers. NSA believes that the most likely scenario is that ultimately charges have to move up towards a level that is economical for the smelting industry over the long term. The benchmark numbers for the last several years at the time of the marketing study are shown in Table 19.1

Table 19.1 Benchmark Smelting Terms

| 2016 | 2015 | 2014 | 2013 | 2012
Copper Treatment Charges ($/dmt) | 97.500 | 107.000 | 92.000 | 70.000 | 63.500
Copper Refining Charges ($/lb) | 0.0975 | 0.1070 | 0.0920 | 0.0700 | 0.0635

For comparison, the spot market of April 2016 indicates sales into the Chinese market where the levels of TCs/RCs were between $90/dmt and $95/dmt of concentrate and $0.090/lb and $0.095/lb of copper, respectively.

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Over approximately 20 years (up to 2005), historical TCs/RCs averaged approximately $77/dmt and $0.077/lb (including approximately $0.01 of participation for these purposes, split between the treatment charge and the copper refining charge) at an average price of $0.93/lb of refined copper.

The general view of NSA was to not expect price participation materializing in the near future; however, it should not be ignored. Historically, when price participation first became a factor in concentrate negotiations it was only applicable at a price level higher than the price existing at the time of negotiation.

NSA suggests that the annual benchmark terms realized at the time of the study are likely to be a guide to future levels. With this in mind, and for purposes of 2016 PFS, the assumption should be a copper treatment charge of $100/dmt, with copper refining charges of $0.10/lb of copper.

TCs/RCs are not the only terms that are used in valuing copper concentrates. Payments and deductions are a matter of negotiation and will vary with many factors, including supply and demand, and custom individual markets.

The following terms are an indication of “standard” long-term smelter charges, including suggested TC/RC terms. Delivery is on the basis of Cost, Insurance and Freight – Free Out (CIF-FO) smelter ports (the mine pays all costs up to delivery port and the buyer arranges and pays for cargo discharge).

Payable Metals

| Copper | Pay 96.5% with a minimum deduction of 1 unit (amount deducted has to equate to a minimum of 1% of the agreed concentrate copper assay).
| Silver | If over 30 g/dmt pay 90%.
| Gold | A scale is applicable with some variations of the following:

| ● | less than 1 g/dmt, no payment

| ● | 1 to 3 g/dmt, pay 90%

| ● | 3 to 5 g/dmt, pay 93%

| ● | 5 to 7 g/dmt, pay 95%

| ● | 7 to 10 g/dmt, pay 96.5%

| ● | 10 to 20 g/dmt, pay 97%

| ● | over 20 g/dmt pay 97.5%

| ● | over 30 g/dmt pay 97.75%.

Gold and silver payments may vary between smelter locations. In China, high gold in copper concentrates is not generally desired; relating more to internal pricing issues rather than technical concerns. Technically, the more modern smelting facilities are able to accept payment formulas similar to Japan and South Korea, but for many of the older smelters in North China, this is not the case. In Europe, with grades of over 40 g/dmt of gold content, payment of 97.75% with a minimum deduction of 1 g/dmt is likely to apply.

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Refining Charges

| Copper | $0.10/lb payable copper
| Gold | $6.00 to $8.00/oz payable gold
| Silver | $0.50/oz payable silver

Treatment Charges

| Treatment Charge | $100.00/dmt CIF-FO main smelter port.

Price Participation

Not applicable at present.

Penalties

| Arsenic: | $2.50 to $3.00 per 0.1% over 0.1% up to 0.5% arsenic
| Antimony: | $3.00 to $4.00 per 0.1% over 0.1% antimony
| Lead: | $2.00 to $3.00 per 1% over 0.5% to 1.0% lead
| Zinc: | $2.00 to $3.00 per 1% over 2% to 3% zinc
| Mercury: | $2.00 per each 10 ppm over 10 ppm mercury
| Bismuth: | $3.00 to $5.00 per 0.01% over 0.03 to 0.05% bismuth
| Selenium: | $3.00 to $5.00 per 0.01% over 0.05% selenium
| Tellurium: | $4.00 to $5.00 per 0.01% over 0.02% to 0.03% tellurium
| Fluorine | $1.00 to $2.00 per 100 ppm over 300 ppm fluorine
| Chlorine | $1.00 to $3.00 per 100 ppm over 300 ppm chlorine.

Furthermore, penalties may also vary from smelter to smelter. It should be noted that for the elements where a percentage range is used, this relates to ranges of penalty thresholds that are negotiated. The penalties noted in this section are generally in line with levels applicable over recent years, but there is a tendency towards higher levels.

Based on the anticipated impurity levels derived from the test results by Tetra Tech (as presented in Table 17.3), the concentrates from the KSM concentrate production are relatively clean, and depending on the market situation at the time of contract negotiations, penalties will likely be minimal if at all applicable. As most of the mill feeds will be the blended materials from different deposits and spatial locations, the blend should effectively mitigate penalty elements rising for the ore from some limited locations.

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Other Off-site Costs

Various indirect costs other than smelter charges include:

| ● | Losses; assumed to be 0.1% or less, due to improvements in material handling.

| ● | Insurance; marine insurance is assumed to be in the range of 0.1 to 0.15% of net invoice value of the concentrate.

| ● | Supervision, assaying and umpire costs; the costs for third-party supervision and assaying are assumed to be approximately US$1/dmt.

| ● | Marketing; the cost of marketing varies with concentrate tonnage, location, and number of smelters to be shipped. For 2016 PFS, the estimated marketing cost is in the range of US$5 to US$10/dmt.

| ● | Concentrate transportation; the transportation costs for copper concentrate are based on the following assumptions by Tetra Tech:

| - | trucking: US$38/wmt

| - | port storage and handling: US$14/wmt

| - | ocean transport to Asian port: US$26/wmt.

19.2 | Molybdenite Concentrate

19.2.1 | Smelting Charge

Molybdenum concentrates of either primary production origin, or as a co-product, need to be further processed. This is initially to produce molybdenum oxide by roasting, or by use of autoclaves for upgrading. Quality is an important consideration and certain elements can be deleterious. As a rule of thumb, 50% molybdenum content is considered the minimum. Below that, buyers will begin to be a bit selective and charges will rise somewhat. One guideline, given for each 1% below 50%, would be an increase in charges of $0.05/lb of molybdenum.

Currently, the deduction for roasting is very quality dependent, with high copper content concentrates generally selling at a 10 to 15% discount. Assuming that the copper content of its molybdenum concentrates is reduced to 0.45% or less, the lower end of the range will apply to clean high-grade concentrates. In the years (2005 to 2009), discounts for high-copper molybdenum concentrates have reached 25%.

In summary, it is recommended to use a discount of 12% from the price, with a minimum of $1.00/lb and a maximum of $2.50/lb. This discount would be inclusive of all charges mine to market, such as delivery costs, irrespective of whether or not the concentrate is sold to a trader or a roaster directly.

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On average, the molybdenum concentrate could contain approximately 1,000 to 2,000 ppm rhenium or higher. Given the high rhenium content, some roasters would recognize the rhenium content in the form of lower treatment charge or some payments. However, it is difficult to project the rhenium value at the current market and the level of study. This potential for additional value in the rhenium content represents a future opportunity.

19.3 | Gold and Silver Doré

There are no gold and silver doré marketing studies conducted for this PFS. No smelter contracts are currently in place or being negotiated. The terms used for this study are based on the typical terms currently used in the markets. All currency amounts used in this section are in US dollars, unless otherwise specified. The general payment terms for gold and silver doré are assumed as below:

| ● | Gold: pay 99.8% of content less a refining charge of US$1.00/accountable oz. Doré transportation is assumed to be US$1.00/oz.

| ● | Silver: pay 90.0% of content less a refining charge of US$1.00/accountable oz. Doré transportation is assumed to be US$ 1.00/oz.

Insurance and assay costing is assumed to be in the range of 0.10 to 0.15% of net invoice value of the doré.

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20.0 | Environmental Studies, Permitting, and Social or Community Impact

20.1 | Licensing and Permitting

The KSM mine development plan was subject to the BC Environmental Assessment Act (BCEAA, the Act), the Canadian Environmental Assessment Act- 1992 (CEAA), and Chapter 10 of the Nisga’a Final Agreement (NFA).

As of January 2020, mine development plans for the KSM property have successfully gone through the provincial and federal environmental assessment review processes, and the appropriate certificates/approvals have been obtained. Additionally, permits for early-stage construction activities, continuation of exploration, and certain permit and project approval renewals have also been obtained. Seabridge continues to advance permitting to allow for the construction of the KSM mine, as well as to continue exploration activities. Details of the provincial, federal, and NFA processes and current statuses, as well as the current permitting status of KSM property, are included in this section.

The KSM property underwent a harmonized EA process with the provincial and federal governments, in accordance with the principles of the Canada-BC Agreement on Environmental Assessment Cooperation (Cooperation Agreement 2004). The process included a working group comprising federal and provincial officials, the Nisga’a Lisims Government (NLG), Aboriginal groups, and local government agencies. Representatives of the US federal and Alaska state agencies were extensively involved in the EA review processes, as a matter of courtesy at the insistence of Seabridge, given that the mineral deposits are located on a tributary of the Unuk River, a transboundary river, 30 km upstream of the US/Canada border. Authorizations are not required from any US federal or state regulatory agency for mining developments on the KSM property to proceed into construction and operation.

The following major permits have been obtained for the KSM area:

| ● | BC MEMPR Mines Act (1996) Permits MX-1-571 (KSM), MX-1-763 (PTMA), MX-1-965 (DKEA).

| ● | Mining Leases 1031440 and 1031441 were issued to Seabridge Gold Inc. on 6 October 2014.

| ● | BC Ministry of Forests, Lands, Natural Resource Operations and Rural Development (BC MFLNRORD) Land Act (RSBC 1996c), Licence of Occupation for MTTs Roadway SK904033. Licence of Occupation for Treaty Transmission Line SK908555.

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| ● | BC MFLNRORD, Forest Practices Code of British Columbia Act Special Use Permits for Treaty Creek (S25751) and Coulter Creek Access Roads (S25750).

| ● | BC MFLNRORD, Forest Act Occupant Licence to Cut L49546 (Mine Site), L49608 (CCAR), L49658 (PTMA) and L49612 (TCAR).

| ● | BC ECCS, Environmental Management Act, Authorization for Effluent Discharge PE106814 (TWTP#6), PE108155 (DKEA), and PE106824 (TWTP#4).

| ● | BC ENV, Environmental Management Act, Authorization PA 106826 for air emissions for five KSM camps.

| ● | BC ENV, Environmental Management Act, Authorizations PR 106834 (Mine Site) and PR106835 (PTMA) for authorized landfills.

| ● | BC ENV, Environmental Management Act, Registration for Municipal Wastewater Regulation Authorization PE 106836 for waste discharge to the environment for Camp 9/10 and PE106837 (Camp 6), PE106839 (Camp 5), EP106809 (Mitchell Operating Camp), PE106841 (Camp 4).

| 20.1.1 | Provincial Process

Under the BC EA process and its regulations, certain categories of larger-scale projects must undergo an EA, and an EA Certificate must be obtained before the KSM mine development can proceed. The scope, procedures, and methods used for each assessment are tailored to the specific circumstances of a proposed project. The EA must assess a project’s potential environmental, economic, social, heritage, and health effects.

Under the BCEAA Reviewable Projects Regulation, the proponent of a new mine facility, with a production capacity of greater than 75,000 t/a of mineral ore, must obtain an EA Certificate. The 2016 PFS will have an annual mill throughput of 43,800,000 t/a, which substantially exceeds this threshold.

Seabridge was accepted into the BC Environmental Review process in March 2008, following submission of a Project Description (Rescan 2013). In July 2013, Seabridge submitted an Application/EIS (Rescan 2013) under the BCEAA (2002) in accordance with the approved project Application Information Requirements (AIR) to the BC Environmental Assessment Office (BCEAO). The Application/EIS was approved and EA Certificate #M14-01 for the KSM property was issued on July 29, 2014 for a period of 5 years.

The full Application/EIS can be found on the BCEAO web site here: http://a100.gov.bc.ca/appsdata/epic/html/deploy/epic_project_doc_list_322_r_app.html

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In October 2018, under Section 18(2) of the Act, KSM Mining ULC applied to the BCEAO for an extension of the deadline specified in the Certificate. On March 21, 2019, the BCEAO extended the date specified in the Certificate #M14-01 by which Seabridge must have substantially started the mine development from July 29, 2019 to July 29, 2024.

The KSM property is in full compliance with the EA Certificate #M14-01 annual compliance report submissions and conditions of the Certificate.

| 20.1.2 | Federal Process

The KSM property was subject to the CEAA (1992) because it may require several statutory authorizations listed in the Law List Regulations. KSM was subject to a comprehensive study level of assessment under the CEAA (1992) because the proposed daily ore mill feed of 130,000 t/d exceeds two thresholds set out in the CEAA (1992) Comprehensive Study List Regulations; specifically, the 4,000 t/d threshold for metal mills, and the 600 t/d production threshold for gold mines. Certain dam structures proposed for KSM also exceed the 10,000,000 m3/a threshold for water diversions.

Development of the KSM property was deemed to require a “comprehensive study” in July 2009 and a “notice of commencement of an environment assessment” was submitted to Seabridge. The terms of the scope of assessment was developed and posted by CEAA for public comment in late May 2010. With the CEAA (2010) amendment, the terms of reference were subsequently re-posted for public comment by the CEA Agency in July 2010. The draft KSM (Kerr-Sulphurets-Mitchell) Project Comprehensive Study Report was subsequently issued by the CEA Agency in July 2014.

The KSM Project Comprehensive Study Report, along with filed public comments from the NLG, other Aboriginal groups, and the public, were considered by the Minister of the Environment when making her final EA decision. KSM received federal approval on December 19, 2014.

Nisga’a Final Agreement (NFA)

The NFA is a treaty signed by Nisga’a Nation, the Government of Canada, and the Government of BC in 1999. The NFA came into effect in May 2000 under the federal Constitution Act and the BC Nisga’a Final Agreement Act, and sets out Nisga’a rights over approximately 27,000 km2 of land in the Nass River system and surrounding drainages.

The NFA establishes three categories of lands with different specified Nisga’a interests-the Nisga’a Lands (approximately 2,000 km2), the Nass Wildlife Area ([NWA], more than 16,000 km2), and the Nass Area (approximately 27,000 km2)—the latter incorporating the Nisga’a Lands and the NWA within it. The NFA affords title to Nisga’a Nation within the Nisga’a Lands and defines the rights of Nisga’a Nation to self-government and law making authority in this area. The NFA also specifies Nisga’a Nation rights to access and make use of natural resources in the NWA and the Nass Area.

Seabridge proposes to develop some components of the 2916 PFS footprint within the Nass Area, including the Treaty OPC, the TMF, and the northern portion of the MTT. No components of the KSM infrastructure will physically occupy any portion of Nisga’a Lands or the NWA, both of which are located south of the potentially affected portion of the Nass Area.

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The NFA makes explicit provision for Nisga’a participation in federal or provincial EAs of projects sited anywhere within the outer Nass Area boundary. Seabridge was directed by the federal and provincial governments to ensure that it conducted its EA responsibilities for KSM in compliance with all relevant Nisga’a treaty rights, including those dealing with economic, social, cultural, and environmental interests. Chapter 10 of the NFA (Environmental Protection and Assessment), paragraphs 6 to 10, provide for meaningful Nisga’a participation in the EA through effective coordination, timely notice, provisioning of information and studies to Nisga’a Nation, and a clear focus on assessment of potential adverse project effects on residents of Nisga’a Lands, the Nisga’a Lands themselves, or more generally, on Nisga’a interests as set out in the NFA.

The Government of Canada worked collaboratively with the NLG and the Government of BC to facilitate the assessment of paragraphs 8(e) and 8(f) effects as part of the comprehensive study. Seabridge conducted an economic, social, and cultural impact assessment (ESCIA) on the well-being of Nisga’a citizens (i.e., paragraph 8(f) effects) based on a work plan that was required by the joint AIR. Effects defined under paragraph 8(e) were described in the Application/EIS as part of Seabridge’s analysis of the effects KSM mine development will have on environmental valued components (VCs).

The KSM Project Comprehensive Study Report examined both paragraphs 8(e) and 8(f) effects on Nisga’a citizens, lands, and interests and provides the federal perspective on these effects. This information, along with comments received during the final public consultation and any agreements between Seabridge and the NLG concerning the effects of the KSM mine development, were used to inform the Minister of the Environment’s NFA Project Recommendation. KSM received its NFA approval in December 2014 following receipt of the Federal CEAA Approval on December 19, 2014.

| 20.1.3 | Provincial Permits

The Application/EIS was accompanied by applications for eligible provincial authorizations in accordance with the BCEAA (2002) Concurrent Approvals Regulation (BC Reg. 371/2002).

This set of initial permits is referred to as the “Batch 1 Permits” and included permits for the following mine components:

| ● | KSM Project Mines Act and Environmental Management Act Permit Application for Limited Site Construction (May 2013)

| ● | Special Use permits for the CCAR and TCAR

| ● | KSM Construction Camps

| ● | KSM Project Treaty Transmission Line

| ● | Land Act tenure for the MTT Roadway.

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In November 2015, Seabridge submitted a Mines Act Notice of Work and an Environmental Management Act (2003) permit application to construct the Deep Kerr Exploration Adit (DKEA). Mines Act permit MX-1-965 authorizing construction of DKEA was issued September 20, 2016, and EMA permit PE-108155 was issued on August 24, 2016 authorizing the discharge of mine effluent from the DKEA project to Sulphurets Creek. The DKEA will be located near the existing temporary KSM exploration camp.

As of January 2020, it is estimated that over 100 provincial permits, licences and approvals are still required to fully develop the 2016 PFS.

| 20.1.4 | Federal Permits

The Application/EIS included applications for the following federal permits and authorizations:

| ● | Metal and Diamond Mining Effluent Regulations (MDMER) Schedule 2 Amendment;

| ● | International Rivers Improvements Act Licence

| ● | Navigable Waters authorization.

Further details are provided in the following sections.

Fisheries Act (RSC 1985, c F-14) - MDMER Schedule 2 Amendment Application

Section 36(3) of the Fisheries Act (1985) prohibits the deposition of deleterious substances into fish-bearing waterbodies; however, deposition of tailings and/or waste rock into such waterbodies may be exempted from this regulation if they are included in Schedule 2 of the MDMER of the Fisheries Act (1985). In order to permit the construction of the proposed tailings management facility in North Treaty Creek and South Teigen Creek (both fish-bearing streams), an application to amend Schedule 2 of the MDMER was submitted with the Application/EIS along with a fish habitat compensation plan designed to offset the habitat losses associated with the Schedule 2 amendment (see MMER Compensation Plan – Appendix 15Q of the Application/EIS). The waterbodies contained in the proposed KSM tailings impoundment area were gazetted and added to the Schedule 2 amendment list in July 2017.

Other Fisheries Act Permitting

Section 35 of the Fisheries Act include the fish habitat protection provisions, which prohibit the harmful alteration, disruption, or destruction (HADD) of fish habitat. In addition to the Schedule 2 Amendment to permit deposition of tailings into North Treaty and South Teigen Creeks, KSM will require an authorization under Section 35(2) of the Fisheries Act to permit fish habitat losses associated with other KSM infrastructure.

The KSM mine development plan entered the EA process prior to the 2012 changes to the Fisheries Act, which removed the habitat protection provisions (i.e., harmful alteration, disruption or destruction (HADD) of fish habitat), and limited the reach of the legislation to “commercial, recreational, or Aboriginal” fisheries. Despite these changes in 2012, the approach to fish habitat offsetting and permitting at the Property was not altered, and the habitat protection provisions were restored in the 2019 changes to the Fisheries Act. Applications for fisheries authorizations are anticipated to be submitted in early 2020.

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International River Improvements Act licence

Seabridge submitted an application for a licence under the International River Improvements Act to Environment Canada on February 2015.

The application was prepared in accordance with Sections 6 and 7 of the International River Improvements Regulations (C.R.C., c982). The application is for improvements on the Unuk River, specifically within the Sulphurets Creek Watershed (which is a tributary to the Unuk River).

The licence was issued to Seabridge on October 21, 2016, and the transfer of ownership to KSM Mining ULC completed in 2018. The licenced Improvements include the Water Storage Facility and ancillary water works for 2016 PFS, representing dams, reservoirs and associated water diversion, collection and management structures for the purpose of: diverting fresh (non-contact) water around the mine site to downstream receiving waters and collecting water that has been in contact with disturbed areas from the mine site for control prior to discharge into the receiving waters.

The Licence is valid for 25 years until October 20, 2041.

Navigable Waters Application

Exemptions to the Navigation Protection Act (RSC 1985, c. N-22) were submitted as components of the Application/EIS. Further to a letter received from Transport Canada dated August 1, 2014, Transport Canada subsequently determined that the waterways of the Property were not navigable waters as defined within the Navigation Protection Act, and as a result there was no requirement to obtain an exemption to the prohibition pursuant to Section 24 of the Navigation Protection Act.

| 20.1.5 | BENEFITS AGREEMENT

Tahltan Nation

On June 10, 2019 the Tahltan Nation and Seabridge Gold Inc. announced that the parties reached agreement on the terms of a Co-operation and Benefits Agreement in connection with KSM. Tahltan Nation voted 77.8% in favour of the KSM Project Impact Benefits Agreement (IBA). The IBA provides a thorough and co-operative framework for the parties to continue building the social licence of KSM mine development through commitments to economic benefits and environmental management of the land.

Nisga’a Nation

On June 17, 2014, Seabridge Gold entered into a comprehensive Benefits Agreement with the Nisga’a Nation in respect of Seabridge Gold’s KSM Property, located in northwest British Columbia. The Benefits Agreement established a long-term co-operative relationship between Seabridge and the Nisga’a Nation under which the Nisga’a Nation will support the mine development, participate in economic benefits from KSM and provide ongoing advice. The Agreement includes commitments by Seabridge regarding jobs and contracting opportunities at KSM, education and training, financial payments and a framework for working together on ongoing development matters. This comprehensive agreement also addresses concerns expressed by the Nisga’a Nation around the potential environmental and social impacts of KSM development.

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| 20.2 | Environmental Settings and Studies

The Property is located in the coastal mountains of northwestern BC, approximately 950 km northwest of Vancouver, 65 km northwest of Stewart, and 35 km northeast of the BC-Alaska border.

The following sections summarize the valued components of the biophysical and socio-community aspects of KSM that have been studied extensively and reported in the Application/EIS (Rescan 2013) under the BCEAA (2002) in accordance with the approved project Application Information Requirements (AIR) to the BC Environmental Assessment Office (BCEAO).

| 20.2.1 | Biophysical Setting

Air Quality

The air quality in the area proposed for KSM development and elsewhere in northwestern BC is predominantly unaffected by anthropogenic sources, reflecting the region’s remoteness and the lack of, and localized nature of, sources of anthropogenic air emissions sources.

Geology and Geochemistry

The mineralized zones in the local area and more regionally, tend to be sulphide-rich. Where sulphide minerals such as pyrite are present, oxidation can create acid rock drainage (ARD), unless sufficient quantities of neutralizing minerals are available. In the event that acidic drainage is formed, low pH conditions can lead to higher rates of metal leaching (ML). Baseline surface water and groundwater quality in the vicinity of mineralized zones in the region exhibit relatively low pH and significant metal concentrations, reflecting the presence of sulphide minerals and the natural occurrence of ML/ARD processes.

Physiography

The KSM topography is very rugged. Glaciers are common in high elevations. Most steep slopes consist of exposed bedrock and accumulations of rubbly colluvium. Gentler slopes have a thin mantle of morainal material (glacial till). Thick glacial deposits are generally restricted to the margins of major valley floors and adjacent lower slopes. Avalanches and slope failures are common features at high and intermediate elevations (above 1,500 masl).

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Geohazards

Locally and regionally, geohazards are linked primarily to landslides and snow avalanches. Landslide hazards are abundant throughout the region. They are attributed to several factors, including the presence of unstable surficial soils and weak bedrock, repeated geologically recent glaciations, resulting in over-steepened valley sidewalls, the loss of slope buttress support following glacial recession, abundance of veneers that are shallow to bedrock, and the high precipitation environment.

Unstable lateral morainal till has been deposited on slopes at angles that exceed the angle of repose, resulting in rubbly colluvium accumulating along moderate steep slopes and valley bottoms. The unloading of the valley walls following glacial retreat has led to pressure release cracks and associated local instability on over-steepened slopes resulting in geohazards, such as rock fall, debris avalanches, and slumping of surficial materials. These geohazard processes are endemic to the local area.

Snow avalanche hazards are abundant due to high elevation, substantial snow supply and generally steeper slope gradients, and tend to be associated with terrain that is open and steep.

Hydrology/Surface Water Quantity

Regional and local surface water quantity characteristics were determined from data collected from specially installed hydrometric stations, used in conjunction with a regional analysis prepared for long-term hydrometric data from Water Survey of Canada hydrometric stations.

The monthly distribution of flow tends to be concentrated in the open water season (May to October), with less than 20% of the annual flow occurring from November to April at a majority of the regional stations.

Groundwater Quantity

Groundwater conditions correspond with the mountainous, wet environment that comprises the mine site and the PTMA. Groundwater gradients are high, driven by heavy rainfall and recharge at higher elevations in the mountains. Valley bottoms are discharge zones, with groundwater levels near or above (artesian) ground surface. Discharge zones also exist along valley walls in the mine site, where seeps of acidic water have been observed (with pH readings as low as 2.5).

Surface Water Quality

The hydrological regime affects water quality in two ways:

| ● | increased flows during freshet, glacial melt, and heavy rainfall events dilutes concentrations of major ions and total dissolved solids

| ● | increased sediment load and transport during high-flow periods leads to increased concentrations of TSS and particle-associated metals.

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Streams near the mine site and PTMA have distinct surface water quality. Metal leaching due to naturally occurring ARD is associated with total and dissolved metal concentrations in Mitchell and Sulphurets creeks that are frequently higher than levels established in BC water quality guidelines for the protection of freshwater aquatic life. The high suspended sediment load, low concentrations of bioavailable nutrients and high concentrations of total and dissolved metals identified in Mitchell and Sulphurets creeks are contributing factors to the poor productive capacity of mine site streams. The lower suspended sediment load, increased concentrations of bioavailable nutrients, and lower concentrations of total and dissolved metals identified in the Snowbank, Teigen, Treaty and Bell-Irving watersheds are contributing factors to the greater productive capacity of PTMA streams relative to the mine site.

Groundwater Quality

Groundwater quality at the mine site is heavily influenced by the sulphide ore deposits. Groundwater is acidic near, and within, the mineral deposits, with pH measurements as low as 2.5 in seeps along the valley walls of Mitchell Creek. Concentrations of certain metals are elevated in groundwater throughout the mine site, and are particularly high near and within the mineral deposits. Metals with elevated concentrations include iron, aluminum, copper, chromium, lead, manganese and zinc. Groundwater in the Mitchell Valley is not suitable for human consumption or the sustenance of fresh water aquatic life.

Wildlife Species

Mature forests, wetlands, alpine areas, and riparian forests provide high-value habitat to a diverse wildlife community. Common species or groups that occur in the RSA include ungulates (e.g., moose and mountain goat), omnivores/carnivores (e.g., grizzly bear, black bear and wolves), furbearers (e.g., fisher, marten and wolverine), hoary marmots, bats, birds (forest birds, raptors and waterfowl), and amphibians (e.g., Columbia spotted frog and western toad). Forest harvesting within the RSA has been minimal compared to many other areas in BC, due to the remoteness of the area and the relatively poor productivity of the forests, so that the wildlife habitats found in the majority of the wildlife RSA are essentially undisturbed.

| 20.2.2 | Economic, Social, and Cultural Setting

Governance

There are five levels of governance in the area of northwestern BC where the 2016 PFS will be developed. Municipal, regional, provincial and federal bodies comprise the non-Indigenous forms of governance, while Indigenous communities have their own governing bodies.

The 2016 PFS is situated in the Regional District of Kitimat-Stikine, and Electoral Area A of the Bulkley Nechako Regional District. Local communities include municipalities, Nisga’a villages, Indian reserves, and unincorporated settlements. Municipal governance only exists for the District of Stewart, the City of Terrace, the Village of Hazelton, the District of New Hazelton and the Town of Smithers. The remaining communities that are not administered by Indigenous bodies (Dease Lake, South Hazelton, Bell II, Meziadin Junction and Bob Quinn Lake) are unincorporated and governed by the regional district in which they are situated.

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Economic Setting

Economically, the region where the Property is located has been dependent upon timber and minerals for well over 100 years. The majority of non-Indigenous communities in the region were initially established to serve natural resource activities such as the mine operations near Cassiar, Stewart, Smithers, and Bob Quinn Lake. Historically, the region’s economic and social diversity has been constrained by limited access and infrastructure, lengthy distances, remote and small communities which provide limited labour or services, and long winters. More recently government and private investment has improved road access, port development, and independent power generation, and brought provincial grid power into some parts of the region. The Red Chris and Brucejack mines recently opened, providing long-term direct and indirect employment to local communities and others within and outside of the region.

Forestry, fishing, and mining were the key economic drivers of northwestern BC through the 1950s to the 1980s. Today, the economies of local communities continue to be largely resource-based, and focus on supporting these sectors in the region.

Overall, the economy in northwestern BC is gradually becoming more diversified and now includes hydroelectric power generation. In some communities, employment levels have increased in the public service, sales and service, tourism, transportation, and mineral exploration sectors. Employment sectors in local Indigenous communities now include sales and service, mineral exploration, labour and government administration components. There are recent signs that the population decline may be reversing.

Today, the mining industry continues to provide an important source of employment in the region, supplying an estimated 30% of jobs for communities along Highway 37 in recent years.

Social Setting

Recent economic conditions in the mining sector influence population levels in regional communities. Population losses have stabilized and are now increasing with the opening of Brucejack and Red Chris mines, hydroelectric power operations, port development and a healthy mineral exploration industry.

Indigenous Groups

Several Indigenous groups may be potentially affected by the development of KSM. The PTMA is situated within the Nass Area, as defined by the NFA, which came into effect on May 11, 2000. The Tahltan First Nation (as represented by the Tahltan Central Council) asserts a claim over part of the 2016 PFS footprint. Both the Gitanyow First Nation (notably wilp Wiiltsx-Txawokw) and the Gitxsan Nation (as identified by the Gitxsan Hereditary Chiefs Office), including wilp Skii km Lax Ha, which represented itself separately in the EA process, have identified potentially affected interests within the broader region, notably downstream of the PTMA. The Skii km Lax Ha are claiming an area covering the mine site and PTMA.

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Land Use Setting

The Property is located in an area of northwestern BC known as the “Golden Triangle”, due to its high mineral potential and the occurrence of several gold projects in the region. For the past century, land and resource uses in the region have been largely driven by forestry, mining and mineral exploration, and this is still true today. A limited amount of commercial and non-commercial recreation also occurs in the region, including hunting, trapping, fishing, heli-skiing, hiking and camping.

Land Use Planning Context

The KSM area is subject to the provisions of two land use plans—the Cassiar Iskut-Stikine Land Resource Management Plan (LRMP) and the Nass South Sustainable Resource Management Plan (SRMP), developed in partnership with Indigenous groups, government and non-government interests.

Mineral resource activity, timber harvesting, commercial recreation and tourism, guide outfitting, hunting, fishing, trapping and cultural land uses are all allowable activities within these land use plans.

| 20.3 | Water Management

| 20.3.1 | Overview of Water Management

An extensive system of water management facilities will be constructed and maintained throughout the life of the KSM mine to divert fresh (non-contact) water away from disturbed areas and to collect water that has contacted disturbed areas (contact water) for treatment before release into the environment. Please refer to Section 18.2 of this PFS for details of the updated water management plans for KSM.

An overview of the Water Management Plan for operations is included in Figure 20.1.

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Figure 20.1 KSM Mine Site Water Management Schematic

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| 20.3.2 | Summary of Water Management Plan

Objectives and Targets

The objectives of the Water Management Plan are to provide a basis for management of surface water on site including:

| ● | diverting non-contact water around the mine site and PTMA

| ● | protecting ecologically sensitive areas and resources and avoiding harmful impacts to aquatic life and wildlife habitat

| ● | providing and retaining water for mine operation

| ● | defining required environmental control structures

| ● | collecting and treating contact water from the mine site to meet discharge requirements prior to release to the receiving environment.

The targets intended to optimize the Water Management Plan to achieve surface water objectives include:

| ● | minimizing the production of contact water by implementation of best management practices and water diversion measures

| ● | collecting and treating contact water where required in order to meet applicable water quality standards

| ● | implementing and maintaining an on-site monitoring and control system to regulate surface water quantity and quality.

Legislation and Standards

The Water Management Plan has been developed in accordance with the following legislation:

| ● | BC Mines Act (1996)

| ● | Fisheries Act (RSC 1985, c F-14)

| ● | Canada Water Act (1985, c. C-11)

| ● | BC Water Sustainability Act (2014)

| ● | BC Water Protection Act (1996).

Processing and Tailing Management Area

Water management in the PTMA is focused on the construction of diversion channels to control and divert water in the PTMA catchment area to either South Teigen Creek or North Treaty Creek.

Monitoring

Monitoring programs will enable KSM mine operators to measure the success of its management strategies and to identify where additional mitigation is necessary.

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Several management plans and monitoring programs include components that will assist in ensuring the long-term protection of the aquatic environment downstream of Property.

| 20.4 | Waste Management

| 20.4.1 | Tailing Management Facility Management and Monitoring Plan

Tailing produced after the mined ore has passed through a process of high pressure grinding, flotation, and leaching at the Treaty OPC will be transported by slurry pipeline to the TMF. Conventional flotation tailing will be stored in two cells, the North Cell and the South Cell. A separate tailing stream consisting of sulphide-rich CIL residue tailing, produced as a waste product in the cyanide leaching process, will be transported in a pipeline to a fully lined CIL Lined Pond located in the Centre Cell between the North and South cells, and will operate during the filling of the North and South cells. The TMF is designed to store 2.3 Bt of tailing produced over the 53-year mine life.

The TMF will ultimately consist of three storage cells retained by four compacted cyclone tailing dams: the North Dam, the Splitter Dam, the Saddle Dam, and the Southeast Dam. The tailing dams will be constructed to final heights of 218 m, 194 m, 168 m, and 239 m, respectively. Seepage from the tailing dams will be collected in seepage collection ponds constructed downstream of the tailing dams.

The tailing dams and associated seepage recovery dams proposed for the 2016 PFS fall into the category of major dams. Prior to commencing work, plans, operating and monitoring procedures, developed in concert with an Independent Tailing Review Board, must be submitted for approval by the Chief Inspector of Mines. The Code also requires supporting plans to address ML ARD, Reclamation and Closure, Emergency Preparedness and Response, and closure cost estimates for security bonding. Please refer to Table 22.5 for the cost.

KSM Mining ULC has obtained Batch 1 of permits for the initial 3 years of construction but has yet to develop and submit detailed applications for the Mines Act permit to construct, operate and close the TMF. A set of detailed mine development and reclamation plans will be submitted at a future date as part of the Mine Plan and Reclamation Program Permit application.

Monitoring

A monitoring program will be developed that will include requirements for inspection of dams and water control structures and procedures for instrumentation monitoring during the construction, operation, and closure phases. The results of the monitoring program will be reviewed on a regular basis to measure the success of the management strategies, to compare the recorded data against design criteria, and to identify where design changes or additional mitigation may be necessary.

A schedule for routine inspection and instrumentation monitoring will be developed at the time of mine permitting based on the mine construction and operation schedule. Additional inspections of the dams and water control structures will be undertaken following extreme rainfall events, significant runoff events, or significant earthquake events.

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The TMF will remain in operation after the end of mining operations until such time as the quality of the water stored in the impoundment reaches an acceptable level for discharge and reclamation is completed. The dam and associated facilities will require ongoing monitoring and maintenance to ensure dam safety and to meet regulatory requirements for dam safety.

In the event of temporary mine closure, visual inspection and maintenance of the dams, diversion channels, collection ditches, and spillways will be required.

| 20.4.2 | Best Available Tailings Technology Assessment

Seabridge recently completed an assessment of tailing technologies, tailing facility locations, and management practices. The assessment was an update of the tailing alternative assessment that was completed as part of the Application/EIS and the Schedule 2 Amendment process and subsequently re-completed in 2015–16 to ensure that the appropriate tailing technology for the project had been selected (see Chapter 33 and Appendix 33-B of the Application/EIS; https://ksmproject.com/bat-report/).

| 20.4.3 | Waste Rock Management

The proposed management of waste rock is outlined in the Rock Storage Facilities Management and Monitoring Plan, which can be found in Volume 26 of the Application/EIS (Rescan 2013).

Waste rock from open pit and underground mining operations that is not used for construction purposes will be consigned to the Mitchell RSF located in the Mitchell Creek Valley and to the McTagg RSF located in the McTagg Creek Valley. Waste rock from the Kerr pit will be backfilled into the mined-out Sulphurets pit. All waste rock placed in the RFSs is assumed to be potentially acid generating. The total amount of mine waste rock to be removed during open pit excavations at the Mitchell pit, Sulphurets pit, and Kerr pit is approximately 3 Bt over the LOM.

NPAG mine waste rock removed from the Sulphurets pit during pre-production will be used to construct the basal drain beneath the Mitchell RSF, and it will be used as rockfill material in the construction of the WSD. A rock drain will also be constructed under the McTagg RSF.

The RSFs proposed for 2016 PFS fall into the category of a major dump as defined under Section 10.5.5 of the Code (BC MEMPR 2017).

A set of detailed mine development and reclamation plans will be submitted at a later date as part of the Mine Plan and Reclamation Program Permit application. The following sections outline the general provisions included in these documents in terms of RSF operation and monitoring.

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Closure

For long-term stability and reclamation purposes, the slopes of the Mitchell and McTagg RSFs will be re-contoured at closure. The modifications to the RSF areas will include:

| ● | re-contouring the Mitchell and McTagg RSF slopes below an elevation of 1,100 m (treeline) to an overall slope of 2H: 1V (27°) with a 50 cm soil cover and vegetated

| ● | re-contouring the western face of the Mitchell RSF below an elevation of 840 m

| ● | above the treeline, higher benches are proposed to be left un-vegetated to reflect existing talus slopes present in the Mitchell and McTagg Creek valleys.

Please refer to the Application/EIS (Volume 26) (Rescan 2013) for further details.

Post-closure

The post-closure phase includes complete reclamation of the RSFs and continued treatment of water collected in the WSF until the water quality meets acceptable standards for direct discharge to the environment. At that time, all facilities will be decommissioned and flows downstream of the mine site will be restored to pre-mine conditions.

| 20.4.4 | Domestic and Industrial non-hazardous and hazardous Waste Management

The proposed management of domestic and industrial waste, including hazardous and non-hazardous waste, and dangerous goods is outlined in the Domestic and Industrial Waste Management Plan and Dangerous Goods and Hazardous Materials Management Plan, (Waste Management Plans) which can be found in the Application/EIS (Volume 26) (Rescan 2013).

Landfills

Two landfills will be established, one in the Mitchell OPC and one in Treaty OPC. The landfills will be used to dispose of only solid inert, non-reactive waste such as used conveyor belts, empty dry latex paint cans, grinding balls, air filters, non-recyclable plastics, and incinerator ash. To deter wildlife attraction to the landfill, the landfill will be fenced and only solid inert waste that will not act as a wildlife attractant will be deposited there. The garbage will be periodically covered with not potentially acid generating waste rock or local till to prevent wind loss and to mitigate wildlife attraction.

Hazardous Waste

Hazardous waste will be produced in all phases. It includes materials such as waste oil, laboratory chemicals and solvents, lead-acid batteries, oil filters, and used oily rags and absorbent pads.

The Hazardous Waste Regulation (BC Reg. 63/88) under the Environmental Management Act (2003) defines “hazardous waste” (see Volume 26 of the Application/EIS for additional details). Any updates on hazardous waste management since the Application/EIS can be found in Section 18.0 of this PFS.

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Closure and Decommissioning

Activities during the closure phase will be similar to the activities during the construction phase. A range of materials will become available for salvage, recycling, or disposal with the dismantling and removal of buildings, surface structures, fuel tanks, etc. The Reclamation and Closure Plan will cover the closure, reclamation, and decommissioning of the mine infrastructures in detail.

Waste Management during Closure

Significant amounts of waste will be generated from the dismantling of buildings and process-related materials. The approach for waste management during closure will be to identify feasible salvage and recycling options.

Upon closure, the buildings, facilities, and process equipment will be dismantled and either disposed of at the site landfill (inert non-reactive materials only) or removed from the site for recycling and/or disposal. Any equipment or materials with market value will be removed for capital recovery.

| 20.5 | Air Quality Management including Greenhouse Gases

The proposed management of air quality, including greenhouse gases, is outlined in the Air Quality Management Plan and Greenhouse Gas Management Plan, which can be found in the Application/EIS (Volume 26) (Rescan 2013).

| 20.6 | Environmental Management System

Seabridge has developed a conceptual Environmental Management System (EMS) and associated Environmental Management Plans for KSM. As stated in the KSM Project AIR document approved by the British Columbia Environmental Assessment Office (BC EAO; 2011), Environmental Management Plans are essential for the EMS for any major development.

An EMS is a requirement of a Mines Act Permit for mines in BC and is the high level framework supporting each Environmental Management Plan. Environmental Management Plans are the specific and detailed goals, objectives, and procedures for the protection of worker health and safety, environmental monitoring, and operating procedures that show the regulatory agencies how legislation and regulations will be met at the mine site and the PTMA. Environmental Management Plans are managed collectively under the umbrella of the EMS.

Environmental Management Plans are to be applied during the planning, construction, operation, closure, and post-closure phases of KSM.

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| 20.7 | Closure and Reclamation

Closure and reclamation planning for KSM will contribute to the success of closure and reclamation during mining and at the end of mine life, which will reduce the need to restructure the infrastructure components, limit the amount of material re-handling, and reduce the environmental effects. Mine development and operation will incorporate techniques to minimize surficial disturbance and, where possible, progressively reclaim areas affected during construction and operation. Stabilizing and rehabilitating surfaces will reduce the potential for degradation of the resources due to extended exposure to climatic factors, reducing closure-related capital costs at the cessation of mining activities.

| 20.7.1 | Closure and Reclamation Objectives

The conceptual closure and reclamation plan has three objectives that provide assurance to the Province that the site will be left in a condition that will limit the future liability to the people of BC:

| ● | to provide stable landforms

| ● | to re-establish productive land use

| ● | to protect terrestrial and aquatic resources.

Provision of Stable Landforms

The design of KSM’s permanent mine-related landforms, such as the open pits, the TMF, the WSD that will impound the WSF at the mine site, seepage containment dams, and the RSFs, has been undertaken to ensure long-term stability during mine operations, after mine closure, and after reclamation works are complete.

Re-establishment of Productive Land Use

The pre-development land use and conditions form the basis for setting the end land use and capability objectives. The goal is to return the site to a use consistent with the current land uses. The current land use information has been obtained from the environmental and socio-economic baseline studies (Appendix 23-A of the Application/EIS) (Rescan 2013). These studies were undertaken in consultation with the KSM Project EA Working Group which includes provincial and federal government agencies, Nisga’a Nation, Tahltan Nation, Gitxsan Nation, Skii km Lax Ha, and the Gitanyow First Nation.

The end land use objective will be primarily to provide for wildlife habitat for the described wildlife species, including bears, mountain goats, and moose.

| 20.7.2 | Soil Handling Plan

The general goal of reclamation is to restore, where possible, the equivalent land capability so that end land use objectives can be achieved. Site reclamation planning will include the conservation of soil materials suitable for reclamation purposes in areas disturbed by mining, and these areas will be re-vegetated, where feasible. The landforms resulting from KSM mine development will also be designed, where possible, and reclaimed to accommodate the desired end land use objective, involving development of appropriate and functional ecosystems, as supported by appropriate soil material handling and re-vegetation strategies.

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| 20.7.3 | Closure and Reclamation Planning

Mitchell Pit and Block Cave Mine Closure

In the 2016 PFS, the Mitchell deposit will be mined as an open pit from Year -2 to Year 24, with an ultimate wall height of 1,230 m, and as an underground block cave mine from Year 23 to Year 53. Closure of the Mitchell pit includes backfilling with water to form a pit lake and placing large rocks on the benches to discourage wildlife access to the pit lake. The Mitchell pit cannot be backfilled with water until underground mining is completed.

Block Cave Mine

At closure, all mobile equipment and supplies, petroleum products, explosives and hazardous materials will be removed and transported off-site.

By the time block caving is completed, the area within the pit directly above the block cave footprint will have subsided into a block cave crater as described in Section 16.3.

Mitchell Pit

The Mitchell pit will commence flooding after the Block Cave has been closed and will take five years

The Mitchell pit closure dam will be constructed on the west side of the Mitchell pit to allow for controlled discharge.

Twin 6 km long drainage tunnels (the Mitchell underground drainage tunnels) will be used to dewater the underground works during operation. At closure, each of these tunnels will be sealed with an engineered concrete plug.

Sulphurets Pit Closure

The Sulphurets pit will be backfilled with the waste rock from the Kerr pit. The Kerr waste rock is predicted to have elevated selenium concentrations, so it will be placed in the Sulphurets pit to allow for the management of selenium.

Kerr Pit Closure

The Kerr pit external haul roads will be decommissioned. The cut and fill slopes will be re-graded for stability, where required. The culverts will be removed, and cross ditches will provide drainage. The surface will be ripped to reduce surface erosion. Any available stored topsoil material will be spread on the surface and re-vegetated. The access road will be retained to permit ongoing inspection of the pit.

Sulphurets-Mitchell Conveyor Tunnel (SMCT) and Associated Overland Conveyors Closure

At closure, the 3 km long SMCT will be dismantled and all mobile equipment and supplies will be removed from the tunnel. The non-salvageable electrical cables and conveyor will be left in the tunnel. The south portal, located adjacent to the Sulphurets pit, and a north portal, located west of Mitchell pit, will both be sealed with engineered concrete plugs.

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Iron Cap Block Cave Mine Closure

At closure, all mobile equipment and supplies, petroleum products, explosives and hazardous materials will be removed and transported off-site. The major infrastructure, such as crushers, rock breakers, conveyors, electrical cable and piping will be left in the mine.

Mitchell OPC

Closure

At the completion of mining, the Mitchell OPC will be decommissioned. Equipment will be removed from the site. The electrical substation will remain. All other structures will be dismantled and removed. Foundations will be broken up, and the concrete rubble will be buried on site. Any soils that are contaminated with fuel will be excavated and treated at a landfarm to remediate the soil.

Reclamation

The Mitchell OPC ground surface will be ripped and covered with crushed rock and up to 50 cm of topsoil prior to re-vegetation with native species

Mitchell Diversion Tunnels Closure

These diversion tunnels will also be used to generate hydroelectric power from the Upper Sulphurets Power Plant during operations. The diversion tunnels will continue to be used for hydro-electric power generation during closure, except when the water is being diverted to the Mitchell pit. The electricity generated will be used to operate the HDS WTP.

McTagg Diversion Tunnels Closure

The McTagg Power Plant will be located east of the Gingras Creek bridge and will be maintained indefinitely to generate electricity. The electricity will be used to operate the HDS WTP, or sold for use in the provincial electricity grid.

Water Storage Facility and Water Treatment Plant

The WSF will remain in service after mine closure to continue collecting contact water that requires treatment.

The HDS WTP and support infrastructure will remain in operation during the closure and post-closure phases. The plant will operate primarily in the spring, summer, and fall months, and minimally in the winter. The lime material will be transported to the site and consumed during these warmer periods. At closure and post-closure, the filter cake (sludge) will be hauled by truck during the summer to the top of the RSFs and placed in an engineered landfill.

An ion exchange Selenium WTP located near the WSD will remain in service after mine closure.

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At closure and post-closure, the tunnels will be required to provide ongoing access to the mine site because the CCAR, which serves the mine site during operation, will be decommissioned. As the HDS WTP will continue to operate post-closure, lime will be required and will be transported from the PTMA through the tunnels to the mine site. All supplies for monitoring, maintenance, and the operation of the HDS WTP will be transported through the tunnels. The train logistics system will be reconfigured and simplified to handle this much reduced traffic requirement.

Treaty Process Plant, Carbon-in-Leach Plant, and Other Structures Closure

Closure

Several structures at the PTMA that will be closed at the completion of mining including the Treaty Process Plant, the CIL Plant, the Treaty OPC waste management facilities, the Treaty OPC Batch Plant, the crusher building, and several other structures such as the warehouse and lab. All of these structures contain equipment that must be removed at closure.

Once all of the equipment has been removed, the buildings will be dismantled and the materials will be moved off-site for recycling or disposal. Any contaminated soil will be collected and placed in the landfarm. Any materials that can be incinerated will be incinerated on site. The concrete foundations of the various buildings will be broken up, buried, and used for road maintenance or as armouring in TMF reclamation. All equipment and debris will be removed from around the structures.

Reclamation

The footprint areas and the area surrounding the buildings will be deep-ripped to reduce compaction and to improve surface drainage. Approximately 30 cm of soil will be spread over the area as the surface material to improve conditions for vegetation establishment and growth of native plants.

Coarse and Fine Ore Stockpiles Closure

Closure

All of the ore in the coarse and fine stockpiles will be processed. The footprint areas will be cleaned and ripped to reduce compaction and to increase downward drainage.

Reclamation

The footprint areas will be covered with a lime mixture and topsoil, and then vegetated with native plants as described for the RSFs.

Treaty Ore Preparation Complex Batch Plant Stockpile Closure

The Treaty OPC may contain stockpiled materials left over from operations consisting of sands and gravels. Any remaining materials will be used for road maintenance or as rip-rap along the beach edges in the TMF.

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Structures Required Post-closure

Some structures in the PTMA will be required for on going monitoring of mine post-closure. These include:

| ● | the office complex

| ● | the ambulance building

| ● | substation 1

| ● | the operating camp incinerator

| ● | the administration building

| ● | the operating camp (reduced in size from the operation phase).

As described above, the MTT will remain open because the HDS WTP will continue to operate post-closure, and reagents, including lime and personnel operating the HDS WTP, will be transported through the tunnels. A smaller camp will be required to accommodate personnel.

Tailing Management Facility Closure

Following operations, the TMF will be reclaimed to provide for wildlife and wetland habitat. The dams and beaches of the TMF will be reclaimed in stages, with the North Cell being reclaimed during operation, the South Cell during closure, and the Centre Cell during post-closure. The TMF facility will be reclaimed in accordance with the described reclamation and closure plan as follows.

The TMF North Cell will be closed approximately 5 years after tailing deposition into it ceases following expansion of the beach area and the time for water quality to improve for discharge.

The South Cell will be closed in a similar manner to the North Cell.

The Centre Cell is the last cell to be closed, which will be about 5 years following mine closure. The CIL tailing in the CIL lined pond will be covered with approximately 1 m of non-reactive tailing and submerged under 5 m of water.

Access Roads

Treaty Creek Access Road Closure

During closure and post-closure, the TCAR will provide the only remaining road access to the Property. All materials and personnel will be transported via the TCAR, which will extend to the Treaty OPC and the MTT portals, which will all remain open to allow for access to the mine site post-closure.

Coulter Creek Access Road Closure

The CCAR will be decommissioned post-closure. The bridges will be dismantled, and materials that are combustible will be burned. Concrete will be broken and used as rip-rap along the creeks, if required, to reduce potential surface erosion that could occur during the dismantling of the bridges.

Culverts will be removed to restore natural drainage patterns. Cross-ditching will provide drainage across roads and will reduce the potential for surface erosion. The surface of the road and any compacted areas will be ripped, where required, to promote surface drainage and to reduce runoff and potential road bed failure.

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Quarry and Borrow Sources Closure

The borrow areas will be cleared and grubbed. The quarry and borrow areas will be re-sloped and re-contoured to ensure escape routes for wildlife and to restore natural landscape.

Mini Hydro Plants and Energy Recovery

Energy recovery and mini-hydro plants are included in theKSM mine development plan. These plants generate electrical power by making use of facilities already included in the 2016 PFS, resulting in significant net energy savings. The power plants will be used to provide electricity for the mine closure requirements, or the electricity may be sold back to BC Hydro under the Standing Offer Program.

Water Treatment Plant Energy Recovery

Water pumped from the water storage pond to the HDS WTP will generate electric power. A small impulse Turgo-type turbine will be used. The output may be fed into the plant power distribution system at the HDS WTP. This facility will continue to operate after mine closure.

McTagg Diversion Hydro – McTagg Power Plant

The McTagg Power Plant will generate energy at the downstream outlet of the MTDT. The McTagg Power Plant will be constructed in Year 10, once the diversion tunnel inlets of the MTDT are raised in Phase 2. It will consist of two Pelton turbines and will feed power into the plant distribution system at the HDS WTP. This facility will continue to operate after mine closure.

Construction and Operation Camps

Closure

There will be 2 operating camps (the Mitchell Operating Camp and the Treaty Operating Camp) and 10 construction camps.

The camps will generally include portable trailers, an incinerator, materials and equipment storage areas, a helicopter pad, a helicopter fuelling area, fuel storage, a septic field, water/sewage treatment, and diesel generators. The portables will be set up so that they can be dismantled and used at the different sites, as required.

All construction and operation camp sites will be reclaimed to a slope compatible with the surrounding natural topography. The high traffic areas will be ripped in two directions to increase surface drainage and to allow for deeper root penetration.

Reclamation

Prior to construction, topsoil will have been salvaged from the camp site areas and stockpiled along the edges of the camps. At closure, this soil will be spread over the disturbed areas. These areas will then be re-vegetated with the native grasses, shrubs, and tree seedlings that were described for the RSFs.

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| 21.0 | Capital and Operating Cost Estimates

| 21.1 | Initial Capital Costs

An initial capital of US$5.005 billion was estimated for the 2016 PFS, based on capital cost estimates developed by the following consultants:

| ● | MMTS: open pit mining, mine roads, ore trains, and infrastructure and water diversion tunnels

| ● | Lambert Civil Consulting Ltd. (LCC): costing of water management structures and the TMF, designed by KCB

| ● | EBC Inc. (EBC): costing of the WSD, designed by KCB

| ● | Tetra Tech: process plant and associated infrastructure, including plant site preparation, water treatment plant, construction camps, winter access road and review of KCC and EBC cost estimates

| ● | Brazier: permanent power supply, fire detection, mini hydro plant, and energy recovery systems

| ● | ERM: environmental

| ● | BGC: landslide management, avalanche management, and pit depressurization

| ● | McElhanney: main access roads (TCAR, CCAR)

| ● | Seabridge: Owner’s costs.

All currencies in this section are expressed in US dollars, unless otherwise stated. Costs have been converted using a fixed currency exchange rate of US$0.80 to Cdn$1.00. Metal prices are based on the three-year trailing average prices from July 31, 2016 back to August 1, 2013.

The expected accuracy range of the capital cost estimate is +25/-10%.

The costs stated in Table 21.1 include only initial capital, which is defined as all costs to build the facilities that mine, transport, and process ore to produce first concentrate and doré. Costs incurred during ramp-up of the mine and process plant in Year 1, through commercial production, are included in the operating costs in Section 21.3.

This estimate was prepared with a base date of Q2 2016. The estimate does not include any escalation past this date. Budget quotations were obtained for major equipment; vendors provided equipment prices, delivery lead times, spare allowances, and freight costs to a designated marshalling yard in northern BC, with some exceptions for delivery points to different BC locales. The quotations used in this estimate were obtained in Q1 and Q2 2016, and are budgetary and non-binding.

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For non-major equipment (i.e. equipment less than US$100,000), costing is based on in-house data, quotes from previous studies. No cost escalation is included. A comparison of 2016 quotes vs. 2012 quotes for the same vendor revealed no significant trend of escalation; therefore, no escalation was applied.

All equipment and material costs include Incoterms FCA. Other costs such as spares, taxes, duties, freight, and packaging are covered separately in the estimate as indirect costs.

The initial capital cost summary and its cost breakdown structure (CBS), which is based on the work breakdown structure (WBS) for the PFS, are presented in Table 21.1.

Table 21.1 Initial Capital Cost Summary

Major Area No. | Major Area Description | Cost (US$ M)
1 – Direct Costs
1.1 | Mine Site | 1,218
1.2 | Process | 1,336
1.3 | TMF | 441
1.4 | Environmental | 15
1.5 | On-site Infrastructure | 23
1.6 | Off-site Infrastructure | 120
1.7 | Permanent Electrical Power Supply and Energy Recovery | 159
Total Direct Costs | 3,312
2 – Indirect Costs
2.91 | Construction Indirect Costs | 449
2.92 | Spares | 34
2.93 | Initial Fills | 20
2.94 | Freight and Logistics | 99
2.95 | Commissioning and Start-up | 6
2.96 | EPCM | 231
2.97 | Vendor’s Assistance | 23
Total Indirect Costs | 862
3 – Owner’s Costs
3.98 | Owner’s Costs | 160
4 – Contingency
4.99 | Contingency | 671
2016 PFS Capital Cost Total | 5,005

Notes: Costs have been rounded to the nearest million dollars.

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| 21.1.1 | Exclusions

The following items are not included in the capital cost estimate:

| ● | force majeure

| ● | schedule delays, such as those caused by:

| q | major scope changes

| q | unidentified ground conditions

| q | labour disputes

| q | environmental permitting activities

| q | abnormally adverse weather conditions

| ● | receipt of information beyond the control of the EPCM contractors

| ● | salvage value for assets only used during construction

| ● | cost of financing (including interests incurred during construction)

| ● | sales taxes (PST, GST and HST)

| ● | royalties or permitting costs, except as expressly defined

| ● | schedule acceleration costs

| ● | working capital

| ● | cost of this study and future feasibility study

| ● | sunk costs.

Labour Rates

A standard labour rate has been applied to various areas of the PFS. The standard construction labour rate used is US$76.00/h (Cdn$95.00/h) and is considered fully burdened. The base labour rate of US$33.10 (Cdn$41.38) was calculated from a combination of union rates published by Mine site for BC (2015), independent contractor quotes, Christian Labour Association of Canada, and recent construction projects in BC. The loaded labour rate build-up uses the following percentages:

| 21.1.2 | Direct Costs

Mine Site

The Mine Site capital costs are US$1.2 billion and presented in Table 21.2.

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Table 21.2 Mine Site Capital Costs

Area No. | Area Description | Cost (US$ M)
1.1.01 | Open Pit Mining | 436
1.1.02 | WSF | 285
1.1.03 | SWM | 222
1.1.04 | Water Treatment | 183
1.1.05 | Ancillary Buildings | 4
1.1.06 | Site Services and Utilities | 2
1.1.07 | Power Supply and Distribution | 4
1.1.08 | Camps | 58
1.1.13 | Geohazards | 24
Mine Site Capital Costs Total | 1,218

Note: Costs have been rounded to the nearest million dollars.

Open Pit Mining

Open pit capital costs were derived from a combination of supplier quotes and historical data collected by MMTS. Pre-production operating costs of the mining fleet and earth works for pioneering and construction quarries are included in the open pit mining capital.

Water Storage Facility

The WSF capital cost was estimated at US$285.6 million. The main dam structure is the largest cost at US$122.2 million.

Process

The battery limits for process are the Mitchell OPC (primary crusher), the MTT (including Saddle), trains, and the Treaty OPC (crushing, grinding, flotation and leaching equipment; the concentrator and other Treaty OPC buildings). TMF costs are separate from process costs. Process capital costs were estimated at US$1.336 billion and TMF capital costs were estimated at US$440.7 million. Details of process and OPC capital costs are presented in Table 21.3.

Table 21.3 Process-Treaty OPC Capital Cost Estimate

Area No. | Area Description | Cost (US$ M)
1.2.01 | Primary Crushing | 52
1.2.02 | Ore Delivery Tunnel | 365
1.2.05 | MTT Material Transportation | 222
1.2.06 | TWTP No. 4 Water Treatment – Saddle | 4
1.2.07 | TWTP No. 8 Water Treatment – Treaty | 3
1.2.08 | Treaty OPC – Coarse Ore Stockpile | 83
1.2.09 | Treaty OPC – Secondary Crushing | 59
| table continues…

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Area No. | Area Description | Cost (US$ M)
1.2.10 | Treaty OPC – Fine Ore Stockpile | 47
1.2.11 | Treaty OPC – Tertiary Crushing - HPGR | 70
1.2.12 | Process Building | 74
1.2.13 | Primary Grinding | 75
1.2.14 | Copper Flotation | 43
1.2.15 | Pyrite Flotation | 7
1.2.16 | Pyrite Concentrate Regrinding | 23
1.2.17 | Cyanide Leaching | 42
1.2.18 | Gold/Silver Refinery | 10
1.2.19 | Copper Concentrate Handling | 5
1.2.20 | Molybdenum Floatation Circuit | 4
1.2.21 | Molybdenum Concentrate Handling (including Leaching) | 4
1.2.22 | Cyanide Recovery and Destruction | 18
1.2.23 | Reagent Area | 5
1.2.24 | Plant Control System | 4
1.2.25 | Site Services and Utilities | 12
1.2.27 | Treaty OPC – Temporary Laydown Area | 1
1.2.30 | Treaty OPC – Ancillary Buildings | 26
1.2.31 | Process Plant Utilities | 7
1.2.32 | Treaty OPC – Mobile Equipment | 9
1.2.33 | Power Supply and Distribution | 6
1.2.34 | Treaty OPC – Roads | 1
1.2.35 | Treaty Operations and Construction Camps | 55
Process-Treaty OPC Capital Cost Total | 1,336

Note: Costs have been rounded to the nearest million dollars.

| 21.1.3 | Indirect Costs

Indirect costs for the PFS were estimated at US$862.2 million, as shown in Table 21.4.

Table 21.4 Indirect Capital Costs

Area No. | Area Description | Cost (US$ M)
2.91 | Construction Indirect Costs | 449
2.92 | Spares | 34
2.93 | Initial Fills | 20
2.94 | Freight and Logistics | 99
2.95 | Commissioning and Start-up | 6
2.96 | EPCM | 231
2.97 | Vendor’s Assistance | 23
Indirect Capital Costs Total | 862

Note: Costs have been rounded to the nearest million dollars.

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| 21.1.4 | Owner’s Costs

An estimate of US$160.2 million is included in the capital cost estimate for Owner’s costs. This cost had been calculated by Seabridge from first principals based on personnel requirements and onboarding of Owner's staff for supporting both the latter part of the commissioning effort for the construction and onboarding to fill all operations and G&A departments. In addition to labour costs, Owner's costs include off-site office staffing, off-site office facilities, travel, off-site office general expenses, recruiting and training expenses, consulting, insurance, general field expenses, and mineral lease/claims costs. Environmental department in the Owner’s costs include environmental monitoring programs, community relations, communication and public relations, wetland compensation, and permitting costs. Limited duty expenses are also included for certain large capital equipment imported from outside North America.

| 21.1.5 | Contingency

A contingency allowance is included to cover additional costs that could occur as a result of more detailed design, unexpected site conditions, or unusual cost escalation. This estimate adequately covers minor changes to the current scope expected during the next phase of this PFS. The estimated contingency cost is US$671.0 million. The contingency estimate was developed on a line-item basis to account for the specific design details and information available for each area, rather than a single value applied to the sum of all direct, indirect, and Owner’s costs. The values applied range from 5 to 25%.

| 21.2 | Sustaining Capital Costs

The sustaining capital costs are all capital costs required from Year 1 of operations to sustain the mining operation for the LOM. Details of the total sustaining cost of US$5.503 billion, required for the LOM, is presented in Table 21.5.

Table 21.5 Sustaining Capital Costs

Major Area No. | Major Area Description | Cost (US$ M)
1.1 | Mine Site | 3,933
1.2 | Process | 355
1.3 | TMF | 502
1.5 | On-site Infrastructure | 0
1.6 | Off-site Infrastructure | 11
1.7 | Permanent Electrical Power Supply and Energy Recovery | 196
2.91 | Construction Indirects | 57
2.92 | Spares | 33
2.94 | Freight and Logistics | 66
2.96 | EPCM | 22
2.97 | Vendor's Assistance | 35
4.99 | Contingency | 293
Sustaining Capital Cost Total | 5,503

Note: Costs have been rounded to the nearest million dollars.

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| 21.2.1 | Mine Site

The sustaining capital for the Mine Site is US$3.9 billion presented in Table 21.6. This number covers the direct capital costs for the LOM and includes all open pit and underground mining operations, as well as the Selenium WTP and the geohazards direct capital cost.

Table 21.6 Mine Site Sustaining Capital Costs

Area No. | Area Description | Cost (US$ M)
1.1.01 | Open Pit | 821
1.1.03 | SWM | 262
1.1.04 | Water Treatment | 113
1.1.05 | Ancillary Buildings | 44
1.1.09 | Mitchell Block Caving | 1,848
1.1.10 | Iron Cap Block Caving | 780
1.1.11 | Kerr Pit Power Supply | 5
1.1.12 | Sulphurets Pit Power Supply | 2
1.1.13 | Geohazards | 58
Mine Site Sustaining Capital Cost Total | 3,933

Note: Costs have been rounded to the nearest million dollars.

| 21.2.2 | Open Pit Mining

Sustaining capital is based on both fleet expansions and unit replacements over the LOM. Major fleet expansions were planned for Years 1, 6, and 10 as the mining rate and haul distances increase. Capital replacement costs for mobile equipment were calculated based on the expected life of the equipment, the cost of the unit, and the utilization for that equipment.

| 21.2.3 | Underground Mining (Block Caves)

The capital costs presented in this section are for both Iron Cap and Mitchell block cave mines, and are a combination of direct and indirect costs, and applied contingencies. The direct cost estimates have been obtained from supplier quotes for the major mobile equipment and capital infrastructure, such as conveyors, crushers, and ventilation infrastructure. Mine development, mobile equipment costs, and certain fixed and ventilation costs are the responsibility of Golder. The costs for the conveyors, secondary rock breakers (if applicable), crushers, related ancillary underground installations, and indirect costs on infrastructure are the responsibility of Tetra Tech. The underground electrical system cost estimate is the responsibility of Brazier.

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Mitchell Block Caving

The Mitchell block cave mine capital cost estimate includes the purchase and installation of all equipment and the excavation of all the underground excavations. The total capital costs are US$1.8 billion, which also includes a unit capital cost of US$4.50/t.

Iron Cap Block Caving

The Iron Cap block cave mine capital cost estimate includes the purchase and installation of all equipment and the excavation of all the associated underground workings. The total capital costs were estimated to be US$872 million, which also includes a unit capital cost of US$3.87/t.

| 21.2.4 | Mine Site Water Treatment

HDS WTP

The sustaining water treatment cost for the Mine Site is US$113.1 million.

The water treatment capacity at the HDS WTP will increase with the addition of two more clarifiers to cater for the increased water flow that is expected.

When the HDS WTP is commissioned in Year -1, the initial maximum throughput capacity will be 5.35 m3/s.

In Year 5, the plant capacity will increase from the 5.35 m3/s initial capacity to the 7.5 m3/s final capacity. Two additional circuits will be constructed and operated in parallel to the existing five circuits.

Selenium WTP

In Year 5, a 500 L/s Selenium WTP, located adjacent to the WSF near the toe of the Mitchell/McTagg RSFs, will be constructed and become operational to treat seepage from the Sulphurets pit backfill (Kerr waste rock), seepage from the RSFs, and water pumped from the WSF.

| 21.2.5 | Process

Process sustaining capital costs include the Sulphurets pit primary crusher and ore delivery tunnel installed in Year 1, and the Kerr pit primary crusher and rope conveyor installed in Year 23. New primary crushers will be installed for the Mitchell underground mine in Year 22 and for the Iron Cap underground mine in Year 31.

During operation, some process related facilities will be constructed after commissioning of the Process Plant. A 3.0 km SMCT connecting the Mitchell and Sulphurets sites, will be constructed in Year 1. An overland conveyor will be installed inside SMCT for ore conveyance.

The ore from the Kerr deposit, together with the ore from other deposits, will be introduced to the Process Plant starting from Year 24. The ore will be further conveyed through the SMCT to the Sulphurets/Kerr coarse ore stockpile at the Mine Site. The waste rock from the rope conveyor will be backfilled into the Sulphurets pit.

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The Iron Cap ore and the lower Mitchell ore will be mined by block caving and crushed on their own sites to 80% passing 150 mm or finer. The crushed ores will be conveyed to the train transport system for delivery to the coarse ore stockpile at the Treaty OPC.

The estimate also includes a replacement allowance for major process equipment.

The process sustaining capital cost was estimated at US$355.4 million.

| 21.2.6 | Northwest Transmission Line Contribution

The 344 km long, 287 kV NTL runs from the Skeena substation near Terrace, BC, to a new substation near Bob Quinn Lake. This new transmission line was commissioned in the summer of 2014 and currently serves the AltaGas Forrest Kerr Hydroelectric Facility and the Red Chris Mine. A tap from this transmission line will service the KSM operation.

Due to an overrun in the construction cost of the NTL, BC Hydro Tariff Supplement TS37, as approved by the BCUC, was put in place requiring NTL customers to share in the overrun cost. In accordance with TS37, based on a contract (peak) demand of 200 MVA, the required contribution will be just over US$167.7million (Cdn$209.6 million). This amount is separate from system reinforcement and is a required cash contribution. The tariff is not due until the start of commercial production, and BC Hydro offers the option of spreading the payments out over five years, with an applicable finance charge, that is applied in the 2016 PFS.

| 21.2.7 | McTagg Diversion Tunnel Mini Hydro Generation Station

The MTDT mini hydro generation station installation is included in sustaining capital. The current tunnel designs show the initial available hydraulic head to be too low to allow the station to be effective if installed during initial construction. No power will be generated from the Phase 1 MTDT. A penstock tunnel and penstock will be constructed in Phase 2. To enhance hydro power generation during low-flow periods, weirs will be established in Gingras Creek just upstream of the Phase 2 and Phase 3 portals to capture base flow from Gingras Creek. The base flow will be routed to the penstock where it will be combined with the McTagg base flows to increase hydro power generation. Flows in excess of the capacity of the McTagg Power Plant will bypass the de-sanding works and exit into Gingras Creek via the portals. The base flow from the Phase 2 and Phase 3 diversions will pass through a 1 m steel penstock pipe set in a plug at the junction of the penstock tunnel with the diversion tunnel. This pipe will run out to the surface in the penstock tunnel, exiting at a portal on the slope above the McTagg Power Plant. A buried steel penstock will then run down the slope to the McTagg Power Plant. The MTDT mini-hydro generation cost was estimated at US$28.0 million.

| 21.2.8 | Tailing Management Facility

The sustaining capital for the TMF is US$501.4 million.

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Initial capital comprises initial TMF perimeter diversions; the North Dam, Splitter Dam, and Saddle Dam; and associated seepage collection dams that form the North and Centre cells. This includes basin preparation for the starter basins for the North Cell and Centre Cell, preparing and lining the starter basins, and provision of liner drainage.

Dam raising for the North and Centre cells is accounted for in both sustaining and operating expenses. Borrow, haul, and placement of till core and expansion of basin preparation, drains and liner are considered sustaining expenses that begin in Year 1. The processing and placement of cycloned sand is considered an annual operating expense for all stages through the LOM.

The South Cell includes additional sustaining capital items occurring between Year 23 and the LOM. These include development of additional perimeter diversions for the South Cell and the Southeast Starter Dam and associated seepage dam. The expansion of the Centre Cell continues as sustaining capital until the end of mine life and the cyclone sand raising continues as an operating expense.

Ancillary expenses such as seepage pumping and monitoring are included as operating expenses throughout the LOM.

| 21.2.9 | Other Sustaining Capital Costs

Other sustaining capital costs were estimated at US$713.7 million, mostly consist of permanent power supply, energy recovery plants, indirect sustaining capital costs and contingency.

| 21.3 | Operating Costs

The average operating cost was estimated at US$12.03/t milled at the nominal process rate of 130,000 t/d, or US$12.33/t for the LOM average. The operating cost estimate does not include the energy recovery credits (approximately US$0.12/t milled LOM) from mini-hydro power stations and the cost relegated to PST (approximately US$0.15/t milled LOM).

The mining operating costs are LOM average unit costs calculated by dividing the total LOM operating costs by LOM milled tonnages. The costs exclude mine pre-production costs.

The cost distribution for each area is shown in Figure 21.1. All costs are expressed in US dollars, unless otherwise specified.

The operating cost estimates in this section are based on budget prices obtained in Q1/Q2 2016 and/or from databases of the consulting firms involved in preparing the operating cost estimates.

When required, certain costs in this report have been converted using a fixed currency exchange rate of Cdn$1.00 to US$0.80. The expected accuracy range of the operating cost estimate is +25/-10%. A summary of the operating costs is presented in Table 21.7.

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Table 21.7 Operating Cost Summary

| At the Nominal Feed Rate of 130,000 t/d* | LOM Average (US$/t milled)
(US$ M/a) | (US$/t milled)
Mine
Mining Costs – Mill Feed | 190.2 | 4.59 | 4.59
Open Pit – Mill Feed | - | 4.40 | 4.40
Block Caving – Mill Feed | - | 4.99 | 4.99
Mill
Process | 251.1 | 5.29 | 5.34
G&A and Site Service
G&A | 43.3 | 0.91 | 1.03
Site Service | 18.9 | 0.40 | 0.44
TMF and SWM
TMF Dam Management | 6.1 | 0.13 | 0.13
Selenium Water Treatment | 9.4 | 0.20 | 0.21
HDS Water Treatment | 22.0 | 0.46 | 0.53
Mine Site Water Pumping | 2.5 | 0.05 | 0.06
Total Operating Cost | 543.5 | 12.03 | 12.33

Notes: *The nominal feed rate estimate excludes mine operating costs and is based on a mill feed rate of 130,000 t/d; the costs do not reflect higher unit costs late in the mine life when the mill feed rates are lower. Costs have been rounded to the nearest hundred thousands of dollars.

Figure 21.1 Operating Cost Distribution

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The operating costs are defined as the direct operating costs including mining, processing, tailing storage, water treatment, and G&A. The hydropower credits from the recovered hydro-energy during mining operations are not accounted for in the operating cost estimate, but is included in the PFS financial analysis. Sustaining capital costs including all capital expenditures after process plant first production are excluded from the operating cost estimate.

Mining Personnel

Salaries for the supervisory and administrative job categories, and all hourly employee labour rates, are based on an industry-focused and location-specific labour survey conducted by Seabridge. Burdens are included in the salaries and labour rates. The payments include base salaries/labour rates, holiday and vacation pay, government prescriptive benefits (e.g., Canadian Pension Plan, workman compensation insurance, etc.), discretionary employer sponsored benefits, and tool allowance costs.

Labour factors in man-hours/equipment operating hours were estimated for operations and maintenance labour for each of the equipment types. Labour costs were calculated by multiplying the labour factor by the equipment operating hours, and labour costs were allocated to the equipment where labour had been assigned. The total hours required for each job type on all the equipment were added, and any additional labour required to complete a crew was assigned to an unallocated labour category.

Process Personnel

Salary/wage rates for management, technical support and operation are based on an industry-focused and location-specific labour survey conducted by Seabridge. Burdens are included in the salaries and labour rates. The payments include base salaries/labour rates, holiday and vacation pay, government prescriptive benefits (e.g., Canadian Pension Plan, workman compensation insurance, etc.), discretionary employer sponsored benefits, and tool allowance costs.

| 21.3.1 | Open Pit Mine Operating Costs

Open pit mine operating costs, including operating and maintenance salaried staff and hourly labour, equipment major component and running repairs, fuel, power, and all other consumable goods, were derived from a combination of supplier quotes and historical data collected by MMTS. The quantities of consumables required were determined for each specific open pit mining activity from vendor input and in-house experience. Labour factors for operations and maintenance of the open pit mining equipment was also estimated based on vendor input and MMTS experience. These inputs were used to build up open pit mine operating costs from first principles.

Freight costs for all consumable goods and fuel are included in the estimate as part of the budgetary quotations.

Major component replacement for larger pieces of mobile equipment were calculated based on the expected life of the major component, the cost of the component, and the fleet size for that equipment. This puts large component repair costs into future years, giving a more representative LOM cash flow.

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The cost of minor parts and running repairs were estimated as an hourly operating cost for the mining equipment.

GME is a category for open pit mine operations, mine maintenance, and technical services departmental overhead costs. It consists of costs for all salaried supervisory and technical staff, a consumable and rental allowance, crane rentals, and software and fleet management systems’ licensing and maintenance. This category is a fixed cost, and does not vary by production or fleet size, with the exception of ramp-ups to full staffing.

The distribution of unit cost by mining area is shown in Figure 21.2.

Figure 21.2 LOM Average Unit Operating Cost for Open Pit Mining (US$/t Material Mined)

| 21.3.2 | Underground Mining Operating Costs

The underground mining operating cost estimates are based on first principles calculations and productivity modelling. The average daily production for each mine was used in the productivity model. The operating cost estimates include direct and indirect costs, which were estimated in a similar manner to direct and indirect capital costs. The indirect labour estimate was a combination of a factor based on the amount of mobile equipment operating for supervisor and maintenance staff, and experience at large, complex mines for technical staff. There are no contingencies added to the operating costs. Golder is responsible for the block cave operating cost estimates.

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Mitchell Block Cave Operating Costs

The average Mitchell block cave mine operating cost was estimated to be approximately US$4.88/t, which includes the cost of equipment and labour required to move material from the drawpoint to the surface conveyor portal, and the fixed costs to operate the LHDs, secondary breakers, crushers and conveyors,

Iron Cap Block Cave Operating Costs

The average Iron Cap block cave mine operating cost was estimated to be US$5.22/t, which includes the cost of equipment and labour required to move material from the drawpoint to the MTT train tunnel and the fixed costs to run the mine. This includes operating the LHDs, crushers, conveyors, mine services, and the labour required to plan and execute the mining plan. The dewatering cost is included in the fixed operating costs because the majority of the dewatering uses gravity and the cost per tonne is relatively low compared to Mitchell.

| 21.3.3 | Process Operating Costs

Summary

The LOM average annual process operating costs for the different mineralizations were estimated as:

| ● | Mitchell and Iron Cap mineralization: US$249 million (US$5.25/t milled)

| ● | Kerr mineralization: US$250 million (US$5.27/t milled)

| ● | Sulphurets mineralization: US$273 million (US$5.75/t milled).

The process operating costs for these mineralizations are based on a process rate of 130,000 t/d and 94% plant availability. The estimated average operating cost for the Sulphurets ore is higher than the ores from the other deposits, mainly due to harder mineralization for the Sulphurets ore, as compared to the other deposits. Due to the variations in operating costs for the different deposit ores, the average operating costs were estimated based on the ratio of the different ore tonnages processed and their individual operating costs.

The estimated process operating costs are summarized in Table 21.8 and Table 21.9, which include:

| ● | personnel requirements, including supervision, operation and maintenance; salary/wage levels based on an industry-focused and location-specific labour survey conducted by Seabridge, the payments include base salaries/labour rates and various burdens

| ● | liner and grinding media consumption estimated from the Bond ball mill work index and abrasion index equations; maintenance supplies are based on approximately 6% of major equipment capital costs

| ● | reagents based on metallurgical test results

| ● | other operation consumables including laboratory, filtering cloth, and service vehicles consumables

| ● | power consumption for the process plant

| ● | no taxes or import duties are included in the process operating cost estimate, unless specified.

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Table 21.8 Summary of Process Operating Costs by Deposit

Area | Mitchell and Iron Cap | Sulphurets | Kerr
Personnel | Annual Cost (US$) | Unit Cost (US$/t milled) | Personnel | Annual Cost (US$) | Unit Cost (US$/t milled) | Personnel | Annual Cost (US$) | Unit Cost (US$/t milled)
Human Power
Operating Staff | 39 | 3,900,000 | 0.08 | 39 | 3,900,000 | 0.08 | 39 | 3,900,000 | 0.08
Operating Labour | 150 | 11,500,000 | 0.24 | 158 | 12,100,000 | 0.26 | 158 | 12,100,000 | 0.26
Maintenance | 84 | 6,900,000 | 0.15 | 84 | 6,900,000 | 0.15 | 84 | 6,900,000 | 0.15
Subtotal Human Power | 273 | 22,300,000 | 0.47 | 281 | 22,900,000 | 0.49 | 281 | 22,900,000 | 0.49
Major Consumables and Supplies
Major Consumables
Metal Consumables |  | 55,100,000 | 1.16 |  | 68,200,000 | 1.44 |  | 53,900,000 | 1.14
Reagent Consumables |  | 91,500,000 | 1.93 |  | 91,500,000 | 1.93 |  | 91,500,000 | 1.93
Supplies
Maintenance Supplies |  | 24,400,000 | 0.51 |  | 25,300,000 | 0.53 |  | 25,800,000 | 0.55
Operating Supplies |  | 2,700,000 | 0.06 |  | 2,700,000 | 0.06 |  | 2,700,000 | 0.06
Subtotal Consumable and Supplies | 173,700,000 | 3.66 |  | 187,700,000 | 3.96 |  | 173,900,000 | 3.67
Power Supply |  | 52,900,000 | 1.11 |  | 62,200,000 | 1.31 |  | 53,200,000 | 1.12
Subtotal Power |  | 52,900,000 | 1.11 |  | 62,200,000 | 1.31 |  | 53,200,000 | 1.12
Process Operating Cost Total | 273 | 248,900,000 | 5.24 | 281 | 272,800,000 | 5.75 | 281 | 250,000,000 | 5.27

Note: Costs have been rounded to the nearest hundreds of thousands of dollars. Unit costs have been rounded to the nearest hundredth. Totals may not add up due to rounding.

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Table 21.9 Operating Costs per Area of Operation by Deposit

Area | Mitchell and Iron Cap | Sulphurets | Kerr
Personnel | Annual Cost (US$) | Unit Cost (US$/t milled) | Personnel | Annual Cost (US$) | Unit Cost (US$/t milled) | Personnel | Annual Cost (US$) | Unit Cost (US$/t milled)
Crushing, Grinding and Copper Flotation Plant | 146 | 161,400,000 | 3.40 | 154 | 184,600,000 | 3.89 | 154 | 162,700,000 | 3.43
Tunnel Transport | 44 | 11,500,000 | 0.24 | 44 | 11,500,000 | 0.24 | 44 | 11,500,000 | 0.24
Molybdenum Flotation Plant | 8 | 5,900,000 | 0.12 | 8 | 5,900,000 | 0.12 | 8 | 5,900,000 | 0.12
Leach Plant | 43 | 33,700,000 | 0.71 | 43 | 34,500,000 | 0.73 | 43 | 33,600,000 | 0.71
Cyanide Solution/Residue Handling | 8 | 29,100,000 | 0.61 | 8 | 29,100,000 | 0.61 | 8 | 29,100,000 | 0.61
CIL Water Treatment | 4 | 1,500,000 | 0.03 | 4 | 1,500,000 | 0.03 | 4 | 1,500,000 | 0.03
Tailing Pumping/Reclaim Water | 20 | 5,800,000 | 0.12 | 20 | 5,800,000 | 0.12 | 20 | 5,800,000 | 0.12
Total | 273 | 248,900,000 | 5.25 | 281 | 272,900,000 | 5.75 | 281 | 250,100,000 | 5.27

Note: Costs have been rounded to the nearest hundreds of thousands of dollars. Unit costs have been rounded to the nearest hundredth. Totals may not add up due to rounding.

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| 21.3.4 | TMF Dam Management Operating Costs

On average, the operating costs for ongoing tailings dam construction by cycloning sand, including seepage water pumping costs, were estimated to be approximately US$6.1 million/a, or US$0.13/t milled.

| 21.3.5 | Mine Site Water Management Costs

Overall SWM at the Mine Site includes HDS site water treatment, selenium removal treatment, and site water pumping. The average annual cost for Mine Site SWM was estimated to be approximately US$34.0 million/a, or US$0.72/t milled, at a mill feed rate of 130,000 t/d. The estimated average operating cost for the HDS WTP is approximately US$0.46/t milled, at a mill feed rate of 130,000 t/d, or US$0.31/m3 water, treated at an average flow rate of approximately 70 Mm3/a. The maintenance manpower will come from the overall mine site maintenance team. The major cost for HDS water treatment is reagent consumption at US$0.33/t milled. Power consumption was estimated to be approximately 40.3 GWh/a.

The estimated LOM average operating cost for the Selenium WTP is approximately US$0.20/t milled at a mill feed rate of 130,000 t/d, or US$1.02/m3 water, treated at an average flowrate of approximately 9.3 Mm3/a. The maintenance manpower will come from the overall site services maintenance team. The major cost for selenium water treatment is US$0.15/t milled for reagent consumption and maintenance. Power consumption was estimated to be approximately 11.8 GWh/a.

| 21.3.6 | General and Administrative

G&A costs are costs that do not relate directly to mining or processing operating costs. These costs include:

| ● | personnel: executive management, staffing in accounting, supply chain and logistics, human resources, external affairs functions, and other G&A departments

| ● | expenses : including insurance, off site offices, administrative supplies, medical services, legal services, human resource related expenses, travelling, community and environmental programs, accommodation/camp costs, air/bus crew transportation, regional and property taxes, and external assay/testing.

The G&A costs were estimated at approximately US$43.3 million/a, or US$0.91/t milled at a nominal mill feed rate of 130,000 t/d, including approximately US$0.28/t for personnel and US$0.63/t for general expenses. The major costs are accommodation and crew air transportation, estimated at about US$15.6 million/a.

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| 21.3.7 | Site Services

The overall site service cost was estimated at US$0.40/t milled or approximately US$18.9 million/a. The estimate is based on requirements for this remote site in northern BC and on in-house experience. The estimate includes the following:

| ● | personnel: general site service human power

| ● | site mobile equipment and light vehicle operations

| ● | potable water and waste management

| ● | general maintenance including yards, roads, fences, and building maintenance

| ● | off site operation expense

| ● | building heating

| ● | power supply

| ● | avalanche control.

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| 22.0 | Economic Analysis

| 22.1 | Introduction

Tetra Tech prepared an economic evaluation of the 2016 PFS based on a pre-tax financial model. For the 53-year LOM and 2.199 billion tonne Mineral Reserve, the following pre-tax financial parameters were calculated using the 2016 Base Case metal prices:

| ● | 10.4% IRR

| ● | 6.0-year payback on US$5.005 billion initial capital

| ● | US$3.263 billion NPV at a 5% discount rate.

Seabridge engaged Lilburn in Denver, Colorado to prepare the tax component of the model for the post-tax economic evaluation for this 2016 PFS with the inclusion of applicable income and mining taxes, and they engaged PwC in Toronto, Ontario to review this work. PwC is an Ontario limited liability partnership, which is a member firm of PricewaterhouseCoopers International Limited, each member firm of which is a separate legal entity.

The following post-tax financial results were calculated:

| ● | 8.0% IRR

| ● | 6.8-year payback on US$5.005 billion initial capital

| ● | US$1.539 billion NPV at a 5% discount rate.

The 2016 Base Case results apply the following key inputs:

| ● | gold – US$1,230/oz

| ● | copper – US$2.75/lb

| ● | silver – US$17.75/oz

| ● | molybdenum – US$8.49/lb

| ● | exchange rate – Cdn$1.00 to US$0.80.

Sensitivity analyses, along with multiple additional metal price scenarios, were developed to evaluate the 2016 PFS economics.

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| 22.2 | Forward-looking Statements

The results of the economic analysis represent forward-looking information that is subject to a number of known and unknown risks, uncertainties and other factors that may cause actual results to differ materially from those presented here.

Forward-looking statements in the 2016 PFS include, but are not limited to, statements with respect to future metal prices and concentrate sales contracts, assumed currency exchange rates, the estimation of Mineral Reserves and Mineral Resources, the realization of Mineral Reserve estimates including the achievement of the dilution and recovery assumptions, the timing and amount of estimated future production, costs of production, capital expenditures, costs and timing of the development of ore zones, permitting time lines, requirements for additional capital, government regulation of mining operations, environmental risks, unanticipated reclamation expenses and title disputes.

Additional risk can come from actual results of reclamation activities; conclusions of economic evaluations; changes in parameters as mine and process plans continue to be refined, possible variations in ore reserves, grade or recovery rates; geotechnical considerations during mining; failure of plant, equipment or processes to operate as anticipated; shipping delays and regulations; accidents, labour disputes and other risks of the mining industry; and delays in obtaining governmental approvals.

| 22.3 | Pre-Tax Model

Metal revenues projected in the 2016 PFS cash flow models are based on the average metal values indicated in Table 22.1.

Table 22.1 Metal Production from the KSM Mine

| Years 1 to 7 | LOM
Total Tonnes to Mill (million) | 322.75 | 2,199
Annual Average Tonnes to Mill (million) | 46.11 | 41.48
Average Grades
Gold (g/t) | 0.82 | 0.55
Copper (%) | 0.24 | 0.21
Silver (g/t) | 2.8 | 2.6
Molybdenum (ppm) | 48 | 43
Total Production
Gold ('000 oz) | 6,529 | 28,597
Copper (million lb) | 1,435 | 8,270
Silver (million oz) | 18.22 | 114.67
Molybdenum ('000 lb) | 11,154 | 62,080
Average Annual Production
Gold ('000 oz) | 933 | 540
Copper (million lb) | 204.9 | 156.1
Silver ('000 oz) | 2,603 | 2,164
Molybdenum ('000 lb) | 1,593 | 1,171

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| 22.3.1 | Financial Evaluations: Npv and Irr

The production schedule has been incorporated into the 100% equity pre-tax financial model to develop annual recovered metal production from the relationships of tonnage processed, head grades, and recoveries.

Metal revenues, principally gold and copper, were calculated based on each scenario's prices. Operating cost for mining, processing, site services, G&A, tailing storage and handling and water treatment, energy recovery areas as well as off-site charges (smelting, refining, transportation, and royalties) were deducted from the revenues to derive annual operating cash flow.

Initial and sustaining capital costs as well as closure and reclamation costs have been incorporated on an annual basis over the mine life and deducted from the operating cash flow to determine the net cash flow before taxes. Initial capital expenditures include costs accumulated prior to first production of concentrate, including all pre-production mining costs. Sustaining capital includes expenditures for mining and processing additions, replacement of equipment, and TMF expansions.

Initial and sustaining capital costs applied in the economic analysis are US$5.005 billion and US$5.503 billion, respectively. LOM Provincial Service Tax (PST) applicable to initial and sustaining capital is estimated to be US$134 million.

Financial evaluations account for physical reclamation costs at various times in the LOM, for the development of a fund to address water treatment costs post reclamation and for special use securities associated with permanent access roads.

Working capital is estimated at two months of receivables and one month of payables and varies from year to year. The working capital is recovered at the end of the mine life.

Pre-production construction period is estimated to be six years. NPV and IRR reported in this section are estimated at the start of this six-year period.

The pre-tax undiscounted annual cash flows are illustrated in Figure 22.1.

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Figure 22.1 Pre-tax Undiscounted Annual and Cumulative Cash Flow

Source: Tetra Tech, 2016

| 22.3.2 | Metal Price Scenarios

The 2016 Base Case uses the three-year average metal prices as of July 2016 and a US$/Cdn$ exchange rate of 0.80. In addition to the 2016 Base Case, three metal price/exchange rate scenarios were also developed: the first uses the metal prices and exchange rate used in mine optimization and design (2016 Design Case); the second uses the spot metal prices and closing exchange rate on July 1, 2016 (2016 Spot Case); and the third uses higher metal prices to indicate upside potential (2016 Alternate Case). The input parameters and pre-tax results of all scenarios are shown in Table 22.2.

Table 22.2 Summary of the Pre-tax Economic Evaluations

| Unit | 2016 Base Case | 2016 Design Case | 2016 Spot Case | 2016 Alternate Case
Gold | US$/oz | 1,230.00 | 1,200.00 | 1,350.00 | 1,500.00
Copper | US$/lb | 2.75 | 2.70 | 2.20 | 3.00
Silver | US$/oz | 17.75 | 17.50 | 20.00 | 25.00
Molybdenum | US$/lb | 8.49 | 9.70 | 7.00 | 10.00
Exchange Rate | US$:Cdn$ | 0.80 | 0.83 | 0.77 | 0.80
Undiscounted NCF | US$ million | 15,933 | 13,727 | 16,101 | 26,319
NPV (at 3%) | US$ million | 6,217 | 5,128 | 6,461 | 11,138
NPV (at 5%) | US$ million | 3,263 | 2,510 | 3,507 | 6,541
NPV (at 8%) | US$ million | 960 | 475 | 1,175 | 2,928
|  |  |  | table continues…

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| Unit | 2016 Base Case | 2016 Design Case | 2016 Spot Case | 2016 Alternate Case
IRR | % | 10.4 | 9.2 | 11.1 | 14.6
Payback | years | 6.0 | 6.5 | 5.6 | 4.1
Cash Cost/oz Au | US$/oz | 277 | 311 | 404 | 183
Total Cost/oz Au | US$/oz | 673 | 720 | 787 | 580

Note: Net cash flow (NCF)

| 22.4 | Sensitivity Analysis

Tetra Tech investigated the sensitivity of NPV, IRR and payback period to the key variables. Using the 2016 Base Case as a reference, each of key variables was changed between -30% and +30% in 10% increments while holding the other variables constant.

Sensitivity analyses were carried out on the following key variables:

| ● | gold, copper, silver, and molybdenum metal prices

| ● | exchange rate

| ● | capital costs

| ● | operating costs.

The analyses are presented graphically as financial outcomes in terms of pre-tax NPV, IRR, and payback period. The NPV is most sensitive to gold price and exchange rate, followed by operating costs, copper price and capital costs. The IRR is most sensitive to exchange rate, capital costs and gold price, followed by operating costs and copper price. In general, sensitivity to metal price is roughly equivalent to sensitivity to metal grade. The payback period is most sensitive to gold price and exchange rate, followed by capital costs, copper price and operating costs. Since the majority of costs are in Canadian currency and the economic analysis is developed in United States of America currency, a significant increase in the exchange rate by 30% will result in a significant increase in the costs when converted to US dollars and this leads to a sharp increase in the payback period. Also, when gold price decreases by 30%, the revenue side decreases significantly and this results in sharp increase in the payback period. Financial outcomes are relatively insensitive to silver and molybdenum prices. The NPV, IRR, and payback sensitivities are shown in Figure 22.2, Figure 22.3, and Figure 22.4.

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Figure 22.2 Sensitivity Analysis of Pre-tax NPV at a 5% Discount Rate

Source: Tetra Tech, 2016

Figure 22.3 Sensitivity Analysis of Pre-tax_IRR

Source: Tetra Tech, 2016

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Figure 22.4 Sensitivity Analysis of Pre-tax Payback Period

Source: Tetra Tech, 2016

| 22.5 | Post-tax Financial Evaluations

Seabridge engaged Lilburn in Denver, Colorado to prepare the tax component of the model for the post-tax economic evaluation for this 2016 PFS with the inclusion of applicable income and mining taxes. Seabridge also engaged PwC in Toronto, Ontario to review the tax component of the model prepared by Lilburn.

The following general tax regime was recognized as applicable in Q3 2016:

| 22.5.1 | Canadian Federal and BC Provincial Income Tax Regime

The federal and BC provincial income taxes are calculated using the currently enacted corporate rates of 15% for federal and 11% for BC.

For both federal and provincial income tax purposes, capital expenditures are accumulated in tax pools that can be deducted against mine income at different prescribed rates, depending on the type of capital expenditures.

Historically, pre-production mining expenditures are accumulated in the Canadian Exploration Expense (CEE) pool. The CEE pool is generally amortized at 100%, to the extent of taxable income from the mine. A phase-out rule was enacted in 2013 to phase out the treatment of pre-production mine development expenses as CEE to Canadian Development Expense (CDE) from 2015 to 2017. Effective 2017, all the pre-production mine development expenses are treated as CDE.

In addition to pre-production mine development expenses (that are reclassified as CDE due to the phase-out rule), Canadian resource property acquisition costs and the costs of mine shafts, main haulage ways, and other underground workings are considered CDE and are accumulated in the CDE pool. The KSM Financial Model for the 2016 PFS treats all such expenses as CDE.

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Fixed assets acquired for the mine are accumulated in an undepreciated capital cost pool (Class 41) and are generally amortized at 25% on a declining balance basis. Certain fixed assets (acquired after March 20, 2013 and before 2021) may qualify to be accumulated in a Class 41.1 pool that can be amortized at an accelerated rate of up to 100%. However, as a substantive portion of the fixed assets are expected to be acquired post-2020 (after the phase out of the Class 41.1 pool), the KSM Financial Model assumes that the accelerated depreciation will not be available.

The KSM Mine is expected to incur costs related to the NTL as described in Section 21.2.6. These sustaining capital costs are expected to be incurred on a property that is not owned by the mine, the costs associated with the NTL should be treated as eligible capital expenditures for income tax purposes. Effective January 1, 2017, eligible capital expenditures are treated as a Class 14.1 asset with an amortization rate of 5%. The KSM Financial Model treats all the costs associated with the NTL overrun as a Class 14.1 asset.

| 22.5.2 | Bc Mineral Tax Regime

The BC Mineral Tax regime is a two tier tax regime, with a 2% tax and a 13% tax.

The 2% tax is assessed on "net current proceeds", which is defined as gross revenue from the mine less mine operating expenditures. Hedging income and losses, royalties and financing costs are excluded from operating expenditures. The 2% tax is accumulated in a Cumulative Tax Credit Account (CTCA) and is fully creditable against the 13% tax.

All capital expenditures, both mine development costs and fixed asset purchases, and mine operating expenditures are accumulated in the Cumulative Expenditures Account, which is amortized at 100% against the 13% tax.

The 13% tax is assessed on "net revenue", which is defined as gross revenue from the mine less any accumulated Cumulative Expenditures Account balance to the extent of the gross revenue from the mine for the year.

A "new mine allowance" is available in respect of new mine or capital costs incurred in connection with expansion of an existing mine commencing production with reasonable commercial quantities. Generally, this allowance provides that 133% of capital expenditures incurred prior to commencement of production may be used to offset net revenue for BC mining tax purposes. Under current legislation, the provision for the new mine allowance expired on January 1, 2020.

A notional interest of 125% of the prevailing federal bank rate is calculated annually on any unused Cumulative Expenditures Account and CTCA and is added to the respective balances.

BC Mineral Tax is deductible for federal and provincial income tax purposes.

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| 22.5.3 | Taxes and Post-tax Financial Results

At the 2016 Base Case long-term metal prices and exchange rate used for this study, total estimated taxes payable on KSM profits are US$5.951 billion over the 53-year LOM. The total estimated taxes payable by the mine in all scenarios provided are shown in Table 22.3.

Table 22.3 Component of the Various Taxes for all Scenarios

| Unit | 2016 Base Case | 2016 Design Case | 2016 Spot Case | 2016 Alternate Case
Gold | US$/oz | 1230.00 | 1200.00 | 1350.00 | 1500.00
Copper | US$/lb | 2.75 | 2.70 | 2.20 | 3.00
Silver | US$/oz | 17.75 | 17.50 | 20.00 | 25.00
Molybdenum | US$/lb | 8.49 | 9.70 | 7.00 | 10.00
Exchange Rate | US$:Cdn$ | 0.80 | 0.83 | 0.77 | 0.80
Corporate Tax (Federal) | US$ million | 2,229 | 1,953 | 2,242 | 3,562
Corporate Tax (Provincial) | US$ million | 1,635 | 1,433 | 1,644 | 2,612
BC Mineral Tax | US$ million | 2,087 | 1,804 | 2,106 | 3,424
Total Taxes* | US$ million | 5,951 | 5,190 | 5,993 | 9,598

*Totals may not add up due to rounding.

Post-tax financial results are summarized in Table 22.4.

Table 22.4 Summary of Post-tax Financial Results

| Unit | 2016 Base Case | 2016 Design Case | 2016 Spot Case | 2016 Alternate Case
Gold | US$/oz | 1230.00 | 1200.00 | 1350.00 | 1500.00
Copper | US$/lb | 2.75 | 2.70 | 2.20 | 3.00
Silver | US$/oz | 17.75 | 17.50 | 20.00 | 25.00
Molybdenum | US$/lb | 8.49 | 9.70 | 7.00 | 10.00
Exchange Rate | US$:Cdn$ | 0.80 | 0.83 | 0.77 | 0.80
Undiscounted NCF | US$ million | 9,983 | 8,537 | 10,109 | 16,721
NPV (at 3%) | US$ million | 3,513 | 2,789 | 3,691 | 6,696
NPV (at 5%) | US$ million | 1,539 | 1,028 | 1,718 | 3,663
NPV (at 8%) | US$ million | -2 | -343 | 161 | 1,282
IRR | % | 8.0 | 7.0 | 8.5 | 11.4
Payback | years | 6.8 | 7.4 | 6.4 | 4.9

Table 22.5 summarizes the 2016 PFS annual cash flow for the pre-production period, Years 1 to 7, and the LOM, providing mine and mill production, revenue projections, operating costs and capital costs, and undiscounted cash flows both before and after taxes.

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Table 22.5 2016 PFS Annual Cash Flow for Pre-production Period, Years 1 to 7 and LOM

| Unit | Pre-prod. Period | Production Periods
Years -6 to -1 | Year 1 | Year 2 | Year 3 | Years 4 | Years 5 | Year 6 | Year 7 | LOM
Mine and Mill Production
Waste Mined | Mt | 90 | 141 | 136 | 40 | 84 | 129 | 130 | 111 | 3,003
Mill Feed Processed | Mt | - | 38 | 47 | 47 | 47 | 47 | 47 | 47 | 2,198
Grade
Gold | g/t | - | 0.72 | 0.82 | 0.96 | 0.98 | 0.71 | 0.72 | 0.82 | 0.55
Copper | % | - | 0.25 | 0.28 | 0.30 | 0.27 | 0.18 | 0.18 | 0.21 | 0.21
Silver | g/t | - | 2.5 | 2.0 | 2.3 | 4.5 | 2.4 | 2.2 | 3.4 | 2.6
Molybdenum | ppm | - | 30 | 47 | 61 | 20 | 56 | 65 | 56 | 43
Metal Recovered
Gold | Moz | - | 0.7 | 0.9 | 1.1 | 1.2 | 0.8 | 0.8 | 1.0 | 28.6
Copper | Mlb | - | 175.3 | 251.6 | 273.2 | 243.2 | 155.0 | 154.6 | 181.6 | 8,270.4
Silver | Moz | - | 1.9 | 1.9 | 2.1 | 4.9 | 2.0 | 1.9 | 3.4 | 114.7
Molybdenum | Mlb | - | 0.6 | 1.5 | 2.6 | 0.0 | 2.0 | 2.4 | 2.0 | 62.1
Gross Revenue
Gold Revenues | US$ million | - | 810 | 1,167 | 1,385 | 1,448 | 1,015 | 1,019 | 1,187 | 35,174
Copper | US$ million | - | 482 | 692 | 751 | 669 | 426 | 425 | 500 | 22,744
Silver | US$ million | - | 35 | 34 | 38 | 87 | 36 | 34 | 60 | 2,035
Molybdenum | US$ million | - | 5 | 13 | 22 | - | 17 | 20 | 17 | 527
Gross Revenue | US$ million | - | 1,332 | 1,906 | 2,196 | 2,204 | 1,495 | 1,498 | 1,763 | 60,480
Total On-site and Off-site Operating Costs | US$ million | - | 662 | 774 | 739 | 745 | 692 | 753 | 775 | 33,216
Operating Cash Flow | US$ million | - | 670 | 1,132 | 1,457 | 1,459 | 804 | 746 | 987 | 27,264
Total Capital Costs | US$ million | 5,005 | 623 | 133 | 98 | 127 | 34 | 88 | 75 | 10,508
PST | US$ million | 64 | 6 | 1 | 1 | 2 | 2 | 1 | 0 | 134
Reclamation- Water Treatment Fund and SUP Costs | US$ million | 8 | 52 | - | - | - | - | - | - | 688
Pre-tax Undiscounted NCF | US$ million | (- 5,077) | (- 11) | 998 | 1,358 | 1,331 | 767 | 657 | 912 | 15,933
Corporate Tax (Provincial) | US$ million | - | - | - | - | 92 | 38 | 51 | 88 | 1,635
Corporate Tax (Federal) | US$ million | - | - | - | - | 126 | 51 | 69 | 120 | 2,229
BC Mineral Tax | US$ million | - | 14 | 24 | 31 | 31 | 17 | 17 | 22 | 2,087
Total Taxes | US$ million | - | 14 | 24 | 31 | 249 | 105 | 137 | 230 | 5,951
Post-tax Undiscounted NCF | US$ million | (- 5,077) | (- 25) | 974 | 1,327 | 1,082 | 662 | 521 | 682 | 9,983

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| 22.6 | Royalties

No royalties are included in the KSM financial models. KSM is subject to a royalty of 1% of the NSR payable to Newmont,, capped at US$3.6 million, with a predetermined buyout option. The full amount of the buyout option is paid in Year 1 in the financial model.

| 22.7 | Smelter Terms

The copper concentrate smelter terms and molybdenum concentrate smelter charges that have been applied in the economic analysis are presented in Section 19.1 and 19.2 of this report, respectively.

Gold and silver doré will generally include payment terms as follows:

| ● | Gold: pay 99.8% of content less a refining charge of US$1.00/accountable oz.

| ● | Silver: pay 90.0% of content less a refining charge of US$1.00/accountable oz.

| 22.8 | Miscellaneous Costs and Charges

An assumption is being made that the customer base will be in Asia. The copper concentrate will be shipped in bulk via port facilities in Stewart, BC, and the molybdenum concentrate will be packed in concentrate bags and shipped inside sea containers via port facilities in Prince Rupert, BC. Transportation costs for the copper and molybdenum concentrate are listed below:

| ● | Copper concentrate:

| o | trucking: US$38.06/wmt

| o | port storage and handling: US$14.40/wmt

| o | ocean transport to Asian port: US$26.00/wmt

| o | moisture content: 9%.

| ● | Molybdenum concentrate:

| o | trucking: US$73.20/wmt

| o | port storage and handling: US$12.20/wmt

| o | ocean transport to Asian port: US$88.93/wmt

| o | moisture content: 5%.

Gold and silver doré transportation cost is assumed to be US$1.00/oz.

An insurance rate of 0.125% was applied to the provisional invoice value of the concentrates and doré to cover land-based and ocean transport from the mine site to the smelter.

A US$8.50/dmt charge was applied for marketing and services provided by the Owner’s representative. Duties would include attendance during vessel unloading at the smelter port, supervising the taking of samples for assaying, and determining moisture content.

An overall weight loss of 0.1% was applied in the economic analysis.

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| 23.0 | Adjacent Properties

In 2010, Pretium purchased the Snowfield and Brucejack mineral resource properties from Silver Standard Resources, Inc. In February 2011, Pretium announced an updated estimate of Mineral Resources for their Snowfield project, which abuts against the east side of Seabridge's Mitchell deposit. Table 23.1 summarizes the publicly disclosed resources of the Snowfield project, which were tabulated using a 0.30 g/t gold equivalent cut-off grade (Pretium 2011).

The QP responsible for this section of this Technical Report has not verified the Mineral Resources that were disclosed by Pretium for their Snowfield deposit. While there appears to be similarities between the Mitchell and Snowfield deposits, the Brucejack mineralization reported by Pretium is not necessarily indicative of mineralization found at the nearby Kerr, Sulphurets, Mitchell, or Iron Cap zones.

Pretium has disclosed the updated Mineral Resources and Mineral Reserves with an effective date of January 1st, 2019 for their Brucejack project (Pretium Annual Information Form for the year end December 31, 2019). The Brucejack is a high-grade underground gold and silver deposit that is located approximately 5 km east of Seabridge's Kerr deposit. The mine is currently being operated at 3,800 tpd with an expected mine life of 14 years.

The QP responsible for this section of this Technical Report has not verified the publicly disclosed Mineral Resources and Mineral Reserves associated with the Brucejack project. Furthermore, the QP does not believe that the Brucejack mineralization is necessarily indicative of the mineralization associated with the various mineralized zones at the Property.

Table 23.1 Pretium Snowfield Mineral Resources Using a 0.30 g/t Cut-off

Resource Category | Tonnes (Mt) | Au (g/t) | Ag (g/t) | Cu (%) | Mo (ppm) | Re (ppm) | Au (000 oz) | Ag (000 oz) | Cu (Blb) | Mo (Mlb) | Re (Moz)
Measured | 189.8 | 0.82 | 1.69 | 0.09 | 97.4 | 0.57 | 4,983 | 10,332 | 0.38 | 40.8 | 3.5
Indicated | 1,180.3 | 0.55 | 1.73 | 0.10 | 83.6 | 0.50 | 20,934 | 65,444 | 2.60 | 217.5 | 19.0
Measured + Indicated | 1,370.1 | 0.59 | 1.72 | 0.10 | 85.5 | 0.51 | 25,917 | 75,776 | 2.98 | 258.3 | 22.5
Inferred | 833.2 | 0.34 | 1.90 | 0.06 | 69.5 | 0.43 | 9,029 | 50,964 | 1.10 | 127.7 | 11.5

Source: Pretium website (http://www.pretivm.com)

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| 24.0 | Other Relevant Data and Information

The PEA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the results of the PEA will be realized. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability.

The PEA has been undertaken to evaluate a different approach to developing the Property by incorporating high value block cave mines at Iron Cap and Deep Kerr deposits, with reduced open pit production and associated surface disturbance.

Iron Cap and Deep Kerr resources have expanded since the 2016 PEA.

Mining of resources from open pit and underground are limited in the 2020 PEA plan to a capacity of 2.4 Bt, which is less than the potential economic limits of mining for this 2020 PEA. In addition, mine planning optimization within the larger 2020 PEA resources has resulted in increased mill feed grades and an improved production schedule compared to the 2016 PEA. The mineral deposits remain open at the end of the 2020 PEA mine plan presented in this report.

This 2020 PEA is a scoping level of study based on the Mineral Resources stated in the KSM Mineral Resource estimate (effective date December 31, 2019), including Inferred Mineral Resources in the 2020 PEA design and projected economics.

The 2020 PEA envisages a combined open pit/underground block cave mining operation that is planned to operate for 44 years.

The proposed process plant for the 2020 PEA design has an average process rate of 170,000 t/d. The 2020 PEA is based on a throughput 30% greater than stated in the approved EA as an initial case for evaluating improved business cases that maintain reasonable and achievable mining and processing rates. Future mine to mill and process throughput studies will optimize the mining and processing rates for further business case improvement.

Mitchell open pit is the sole source of mill feed during initial capital payback. The mill feed then transitions to Iron Cap and Deep Kerr underground mines, supplemented by a Kerr and Sulphurets open pit during the transition. There is no Mitchell underground production in the 2020 PEA.

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The flotation plant produces a copper/gold/silver concentrate for transport by truck to a nearby seaport at Stewart for shipment to overseas smelters. Metallurgical testing indicates that KSM can produce a clean concentrate at an average copper grade of 25% with a high gold and silver content. Separate gold-silver doré is also produced at the KSM processing facility from gold rich pyrite products rejected from the gold/copper flotation. No molybdenum concentrate is produced in the 2020 PEA.

The 2020 PEA mine site general arrangement is shown in Figure 24.1

Figure 24.1 Mine Site General Arrangement (2020 PEA)

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| 24.16 | Mining Methods

| 24.16.1 | Net Smelter Return Block Model

The NSR per tonne for each block in the block model used for the PEA mine planning is determined using the Mineral Resources discussed in Section 14.0. It is net of offsite concentrate treatment and smelter charges and inclusive of onsite mill recovery. It is used as a cut-off item for break-even ore/waste selection, as well as for grade bins used to optimize cash flow in the open pit production scheduling. .

NSR is estimated using Net Smelter Price (NSP) and process recoveries. Metal Prices and major smelter terms used to estimate NSP are specified in Table 24.1 and Table 24.2. Process recoveries have been determined using the metallurgical projections described in Section 13.0, with the exclusion of CIL recovery from mill feed derived from Iron Cap and Deep Kerr zones.

Table 24.1 Metal Prices for 2020 PEA NSR Calculation

| Metal Price (US$)
Cu | 2.70/lb
Au | 1200/oz
Ag | 17.50/oz
Mo | 9.70/lb
Exchange Rate (US$:Cdn$) | 0.83

Table 24.2 Major Smelter Terms Used in the 2020 PEA NSR Calculation

| Amount | Unit
Copper Concentrate
Smelting | 100 | US$/dmt
Cu Deduction | 1 | %
Au Payable | 97.75 | %
Ag Payable | 90 | %
Cu Refining | 0.10 | US$/lb
Au Refining | 7.00 | US$/oz
Ag Refining | 0.50 | US$/oz
Transport Costs | 98 | Cdn$/wmt
Moly Concentrate
Roasting | 2.00 | US$/lb
Other Off-site Costs | 5,298 | Cdn$/wmt
Gold Doré
Au Payable | 99.8 | %
Ag Payable | 90 | %
Au Refining + Transport | 2.00 | US$/oz

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| 24.16.2 | Open Pit Mining Method

The objective of the 2020 PEA open pit mining study is to develop an open pit mine plan that:

| · | maximizes the overall profitability of the mill feed

| · | maximizes 2020 PEA NPV

| · | accelerates initial capital payback

| · | reduce the mine footprint and site impacts.

Potential underground (UG) production has been established for Iron Cap and Deep Kerr prior to optimizing of the economic pit limits. Open Pit (OP) limits have been selected by

OP mill feed (mt) | = TMF capacity (mt) - UG mill feed (mt)
| = ~2,400 mt - ~1,600 mt
| = ~800 mt

Pit Optimization

2020 PEA pit optimization is based on the 3DBMs as described in Section 14.0.

Pit Optimization Method

The economic pit limit is selected after evaluating Lerchs-Grossman (LG) pit cases. The assessment is carried out by generating sets of LG pit shells by varying revenue assumptions. Ultimate pit limits are typically determined by estimating the pit size where an incremental increase in pit size does not significantly increase the pit resource and the economic return starts to decrease.

For the 2020 PEA, the available mill feed from typical economic pit limits would significantly exceed the available tailing storage space due to the inclusion of large block cave underground mines for Iron Cap and Deep Kerr. Pit limits have therefore been selected where the total open pit mill feed of ~800 Mt is achieved using the higher economic margin pit phases ($/t net cash flow) from Mitchell, Sulphurets, and Kerr zones. Note that the physical sequencing constraints of these selected phases have been considered during LOM production scheduling.

LG Pit Assumptions

BGC has provided open pit slope design parameters for the three proposed 2020 PEA open pits: Mitchell, Sulphurets and Kerr. Pit slope design parameters are based on geotechnical site investigations, available local and regional geological data, and well-established geotechnical design methods.

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Inputs to the updated LG pit limit assessment shown in Table 24.3 reflect the following assumptions:

| · | Kerr and Sulphurets mill feed will be hauled to the Mitchell primary crusher

| · | Kerr and Sulphurets waste rock will be hauled to the Mitchell RSF

| · | Kerr mill feed processing will not include CIL.

Table 24.3 LG Pit Limit Primary Assumptions

Assumption | Mitchell (Cdn$/t) | Sulphurets (Cdn$/t) | Kerr (Cdn$/t)
Mining Cost | 1.90 | 2.50 | 3.00
Process, G&A, Site Services, Water Treatment | 9.75 | 10.00 | 7.60
Pit Slope Angle | Variable; Based on previous studies
Base Case Metal Prices | See Table 24.1

LG Pit Limits

The 2020 PEA LG pits have been generated by factoring the net revenue basis for each pit shell ranging from 33% to 88% of the 100% case and NSP has been used to vary the net revenue. Note that the tonnes and grades for each LG pit shell are based on the same cutoff grade as the 100% case, therefore the above factoring cannot be considered a price sensitivity. Figure 24.2 shows the resulting mill feed sensitivity to Net Revenue Pit case. From these pit shells, optimal open pit quantities are selected up to the 800 Mt TSF limit allocated for open pit sources as specified above.

Figure 24.2 Sensitivity of 2020 PEA Mill Feed Tonnes to Pit Case

Source: MMTS (2020)

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Combined Open Pit and Underground Production Sequencing

The optimal cashflow for this 2020 PEA study will result from mining higher Net Cash Flow Cdn$/tonne (NCF) mill feed sources earlier in the mining schedule, however, the time and cost required for development of the alternate mining areas also need to be considered. After selection of the initial mining areas, additional areas have been added from higher NCF to lessor, until the full 2.4 Bt TSF capacity limit for this study is reached for both open pit and underground mining.

Figure 24.3 shows the selected incremental LG pit unit NCF per tonne, compared to the overall NCF per tonne for the proposed Iron Cap and Deep Kerr block cave mines. This indicates mining should start production early in the mining schedule with the higher NCF Iron Cap block cave and Mitchell Starter Pits. The Mitchell Starter Pit is the preferred first mill feed source with a lower development cost and a shorter development timeline. As an open pit, it can also meet the full daily mill feed production rate. The next Mitchell pit phase follows while Iron Cap block cave development progresses. This Mitchell pit phase also has the capability to meet the full mill feed rate which will then be reduced as Iron Cap block cave reaches its full throughput capacity. Mitchell open pit mining followed by Kerr open pit then Sulphurets produce in tandem with Iron Cap and Deep Kerr block caves to meet full throughput capacity.

The pit limits selected as indicated in Figure 24.2 and sequenced as indicated in Figure 24.3 are based on where the incremental pit NCF is similar to the lowest block cave overall NCF of Cdn$16/t (Deep Kerr) as summarized below:

| · | Mitchell Pit Limit: Pit Case 12 (58%)

| · | Sulphurets Pit Limit: Pit Case 11 (55%)

| · | Kerr Pit Limit: Pit Case 9 (50%)

where the % value indicates the factor used on the Net Revenue as discussed above.

An initial capital payback optimization analysis has identified the Mitchell Pit Case 48 as a preferred payback/starter pit. This analysis indicates that combined with a premium cutoff grade of Cdn$20/t applied to the Mitchell starter pit, payback of the initial capital cost could be achieved in less than 5 years.

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Figure 24.3 Sensitivity of Unit Net Cash Flow (NCF) to Pit Case

Source: MMTS (2020)

Slope Sensitivity

A pit slope sensitivity analysis has been completed for the Mitchell area by varying all pit slopes angles as follows:

| · | Base Case (no change)

| · | +/- 3o to all sectors

| · | +/- 6o to all sectors

The cross-section in Figure 24.4 shows the resulting LG pits that produce the same mill feed as the Mitchell Payback pit (Base). The impact of different slope angles was tested where angles were varied +/- 6 degrees, and the results indicate that economic outcomes are not sensitive to Mitchell pit slope angles.

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Figure 24.4 Mitchell Slope Sensitivity Section View Looking West

Source: MMTS (2020)

Pit Phase Design

The open pit development is designed as a conventional truck-shovel operation with 363 t trucks and 56 m3 and 40 m3 shovels. For the Mitchell pit, the mine design includes three nested pit phases to balance stripping requirements while satisfying the process plant requirements. The first Mitchell phase includes quarry material for WSD. Sulphurets and Kerr are each mined as one phase.

The design parameters include:

| · | ramp width of 38.20 m

| · | maximum road grades of 8%

| · | bench heights of 15 m

| · | berm interval of 30 m

| · | variable slope angles by sector

A minimum mining width between pit phases is allocated to maintain a suitable mining platform for efficient mining operations. This is established based on equipment size and operating characteristics. For this study, the minimum mining width generally conforms to 50 m, which provides sufficient room for 2-sided truck loading, but where phases merge, it is sometimes less.

Mitchell Pit

A geochemical and geotechnical evaluation of the P8 monzonite zone that occurs in the north side of the Mitchell pit has confirmed that this waste rock is suitable construction material, particularly as rock fills for the WSD and RSF drains (KCB, 2019b). A quarry (Phase M1) containing 32.6 Mt of P8 Monzonite has been designed keeping the majority of the required construction rock inside the Mitchell pit limit and is mined during the pre-production period.

The second Mitchell pit phase M2 is considered the ‘Starter Pit’. Pit phase M3 mines north to the final Mitchell pit limit. Figure 24.5 shows the 2020 PEA Mitchell pit phases.

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Figure 24.5 Mitchell Pit Design – Plan View

Source: MMTS (2020)

Phase design by MMTS for the Mitchell pit has been carried out in an iterative process in collaboration with BGC. The iterations have included pit slope stability analysis by BGC to ensure that stability criteria are met using the most current geotechnical modeling.

Sulphurets Pits

The mine plan for the Sulphurets area includes a single pit (S1). A plan view of the Sulphurets pit is shown in Figure 24.6.

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Figure 24.6 Sulphurets Pit Design – Plan View

Source: MMTS (2020)

Kerr Pit

The mine plan for the Kerr area includes a single pit (K1). A plan view of the Kerr pit phases is shown in Figure 24.7.

Figure 24.7 Kerr Pit Design – Plan View

Source: MMTS (2020)

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Pit Resource Estimate

The pit delineated resources summarized in Table 24.4 are a subset of the mineral resources and quantities that are included in Section 14.0. The pit delineated resources use a variable NSR cut-off grade between Cdn$15/t and Cdn$25/t for Mitchell, and a constant NSR cut-off grade of Cdn$17/t for Kerr and Sulphurets.

Table 24.4 Pit Delineated Resource

Pit | Mill Feed (Mt) | Au (g/t) | Cu (%) | Ag (g/t) | Waste (Mt) | Strip Ratio
Mitchell Pit Phase M1 | - |  |  |  | 33 | n/a
Mitchell Pit Phase M2 | 467 | 0.81 | 0.21 | 3.4 | 565 | 1.2
Mitchell Pit Phase M3 | 145 | 0.63 | 0.16 | 3.4 | 238 | 1.6
Mitchell Pit Total | 612 | 0.77 | 0.20 | 3.4 | 836 | 1.4
Sulphurets Pit | 105 | 0.73 | 0.31 | 1.0 | 233 | 2.2
Kerr Pit | 89 | 0.31 | 0.59 | 1.5 | 252 | 2.8
Total | 807 | 0.71 | 0.25 | 2.9 | 1,321 | 1.6

Waste Rock Facilities

All open pit waste is placed onto the Mitchell RSF. Several different construction methods will be used for waste placement: top-down, bottom-up, and wraparounds. Top-down platform heights have been limited to a maximum height of 300 m. Bottom-up lifts are 30 m to 50 m high, or less if geotechnically required. Wraparounds are built onto the face of an existing RSF, creating a series of terraces to facilitate intermediate haul roads and lower the overall slope angle of high dumps, which will reduce the end of mine closure costs. Figure 24.8 shows the Mitchell RSF.

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Figure 24.8 Rock Storage Facility Design

Source: MMTS (2020)

Prior to mine development, overburden is salvaged from the footprint area where the material is considered as soil, suitable for reclamation purposes. Soils are generally not salvaged on slopes steeper than 26° due to practical limitations on equipment access and operator safety. Soils salvaged from the RSF footprints are stockpiled in the Ted Morris Valley.

Mine waste rock placed in the valley bottoms will be dumped in low height lifts in a manner to confine and consolidate weaker foundation material. If required, loose tills and clays will be removed. Once the foundation is prepared, a basal drain will be placed using suitable rock (e.g., P8 monzonite).

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Production Scheduling (Open Pit)

Development and pre-production

Development and pre-production activities include:

| · | expose sufficient mill feed for start-up

| · | establish mining areas that will support the equipment required to achieve mill feed production, and annual mill feed requirements on a sustainable basis

| · | provide material required for construction in the mine area.

In early development, a contractor fleet will mine Mitchell pit M1. Material from M1 pit will be used to construct the WSD.

During pre-production, Mitchell pit M2 will be pre-stripped using dozers and truck-shovel tramming, where Pre-stripped waste rock will be dozed and short-hauled into the Mitchell valley. Mitchell M2 will be mined to 1110 m elevation on the north side and 1080 m elevation on the south side. The Mitchell mining area at the end of pre-production is shown in Figure 24.9. The fill material (brown) will be completely rehandled to the RSF in first 3 years of Mitchell open pit operations.

Figure 24.9 Mitchell Pit, End of Pre-Production

Source: MMTS (2020)

Open Pit Production Schedule

A series of schedule optimization studies have identified that the 2020 PEA NPV is maximized with the Iron Cap and Deep Kerr underground mines delivering mill feed in Year 6 and Year 12, respectively.

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Open pit production will be carried out to meet the 170 ktpd mill feed requirement incorporating the underground production schedules. Mitchell open pit will be the only operating mine area during initial capital payback. Mining will then transition to Iron Cap and Deep Kerr underground mines, supplemented by Kerr and Sulphurets open pits during the transition. Mill feed will be derived entirely from Iron Cap and Deep Kerr underground mines from Year 17 to Year 28 when Iron Cap production begins to steadily decrease. The M3 Mitchell pit will be mined from Year 26 to 34 and the remaining Deep Kerr underground will continue providing mill feed at 80,000 tpd until depleted by the end of the life of mine (LOM) in Year 44. (See Table 24.20 for the combined open pit and underground production schedule)

Open Pit Major Equipment Selection

The mine load and haul fleet will use diesel hydraulic shovels (85 t bucket class) matched with the 363 t truck. These unit sizes are proven in operating mines around the world. With the relatively short operating life of the open pits, hydraulic shovels have been selected rather than rope shovels.

Suitable drill sizes (311 mm hole size) are chosen to match this size of truck/shovel fleet.

Schedule Criteria

In order to optimize the 2020 PEA NPV, a variable cutoff grade strategy based on NSR block values has been used. Scheduling has been carried out with the MineSight MS-SO optimizer with the primary program objective to maximize the NPV. The schedule optimization is guided by estimated operating and capital costs, process recoveries, and metal prices. Key production schedule assumptions are shown in Table 24.5.

Table 24.5 Open Pit Production Schedule Assumptions

Assumption | Value
Operating Days Per Year | 360
Hours Per Day | 21
Daily Mill Throughput | 170,000 t/d
Haul Truck Speed Limit | 50 km/h
Haul Truck Operator Efficiency | 90%
Haul Truck Operating Efficiency | 83%
Dump and Manoeuvre Time | 1.5 min
Shovel Loading Time | 2.94 min
Shovel Operator Efficiency | 84%
Shovel Operating Efficiency | 83%

The 2020 PEA production schedule requires that Kerr mill feed be kept separate from Mitchell/Sulphurets mill feed. Mill feed from all open pits will be dumped into the Mitchell primary crusher and conveyed to coarse ore bins linked to the train loader system. The coarse ore bins will enable train sets to be loaded with all one ore type so that at the Treaty end they can be distinctly directed to an appropriate stockpile.

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At the mine, the multiple mill feed sources (pits and underground) are only required to segregate to two mill feed types (CIL and none-CIL) to match the two mill feed type stockpiling at Treaty. With sophisticated material management and tracking two coarse ore bins at the Mitchell OPC, the mine could handle this requirement. But for periods where the material balance could overwhelm the bin handling the larger tonnage fraction, an allowance for a third bin has been included when Deep Kerr block cave comes on line. The following system design elements will need to be included in future studies:

| · | each mined source can direct their loads to the designated primary crusher/bin

| · | the trains can be dispatched as appropriate to keep the two bins from over filing

| · | the off-loaded trainloads will be directed to the appropriate stockpile at Treaty

| · | the mill will campaign according to their stockpile inventory.

Pit Production Schedule

The open pit mill feed grades averages are:

| · | 0.89 g/t gold, 0.23% copper, and 3.8 g/t silver during first 5 years of operation.

| · | 0.71 g/t gold, 0.25% copper, and 2.8 g/t silver during LOM.

Open pit copper and gold mill feed grades are shown in Figure 24.10 with production schedule shown in Figure 24.11.

The increase in copper grade and reduction in gold grade between Years 6 and 10 are attributable to Kerr open pit mining.

Figure 24.10 Open Pit Mill Feed Grades

Source: MMTS (2020)

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Figure 24.11 Open Pit Production Schedule

Source: MMTS (2020)

Open Pit Mine Operations

KSM open pit mining operations will employ bulk mining methods and large capacity equipment.

Organization

Mine operations will be organized into three areas: direct mining, mine maintenance, and General Mine Expenses (GME).

In this study, direct mining and mine maintenance are planned as an owner-operated fleet with two 12-hour shifts on a continuous operation. The equipment ownership and labour will be directly under operations. Mining of the M1 Quarry in Years -4 to -3 will be contract mined.

Drilling

Areas will be prepared on the bench floor blast patterns in the in-situ rock. Drill ramps will be dozed on original mountain side surfaces, between benches.

Blasthole drills will be fitted with GPS navigation and drill control systems to optimize drilling.

Diesel hydraulic and electric rotary drills (311 mm bit size) will be used for production drilling, both in mill feed and waste.

Diesel hydraulic percussive drills with a hole size of 6.5 inches (165 mm) will be used for controlled blasting techniques on high wall rows, pioneering drilling during pre-production, and development of initial upper benches.

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Blasting

Powder Factor

An appropriate powder factor has been used to provide adequate fragmentation and digging conditions for the shovels. Similar large open pit designs in the previous studies use a powder factor of 0.32 kg/t for competent rock. Based on a blasting study completed by an explosive supplier, a power factor of 0.35 kg/t is being used for the 2020 PEA.

Explosives

The mine will employ a blasting crew, but the supply and onsite manufacturing of blasting materials will be contracted out. All infrastructure required for the blasting supply contractor will be provided by the operations.

The 2020 PEA mine plan assumes that half of the holes will use a 70/30 emulsion/ammonium nitrate-fuel oil (ANFO) mix explosive (“wet hole” product) and half of the holes will use a 35/65 emulsion/ANFO mix (“dry hole” product).

Blasting accessories will be stored in magazines adjacent to the mining areas suitably located to meet federal and provincial regulations and to avoid potential geohazards.

Explosives Loading

Loading of the explosives will be done with bulk explosives loading trucks provided by the explosives supplier. The trucks should be equipped with GPS guidance and should be configured to receive digital loading instructions for each hole transmitted from the engineering office. The GPS guidance will be a necessity to be compatible with stakeless drilling.

Blast holes will be stemmed to avoid fly-rock and excessive air blasts. Crushed rock will be provided for stemming material and will be dumped adjacent to the blast pattern. A loader with a side dump bucket is included in the mine fleet to tram and dump the crushed rock into the hole.

Blasting Operations

The blasting crew will comprise mine employees and will be on day shift only. The blasting crew will coordinate drilling and blasting activities to target a minimum of two weeks of blasted material inventory is maintained for each shovel. Blasters will require hand-held GPS to identify holes for pattern tie-in as blast patterns will not be staked. A detonation system consists of electric cap initiation, detonating cord, surface delay connectors, non-electric single-delay caps, and boosters will be used. Blasting assumptions are summarized in Table 24.6.

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Table 24.6 Blasting Assumptions

Blasting Pattern – Mill Feed and Waste | Specifications
Spacing | 8.5 m
Burden | 8.5 m
Hole Size | 311 mm (12¼”)
Explosive In-Hole Density | 1.25 g/cc
Explosive Average Downhole Loading | 95.0 kg/m
Bench Height | 15 m
Collar | 6 m
Loaded Column | 11 m
Sub-drill | 2 m
Charge per Hole | 1,046 kg/hole
Rock SG | 2.77 t/m3
Yield per Hole | 3,002 t/hole
Powder Factor | 0.35 kg/t

Loading

Diesel hydraulic shovels (42 m3) have been selected as the primary digging units. The diesel hydraulic shovels are selected for flexibility and mobility. Minimum bench widths of 50 m are designed to ensure sufficient operating width for double-sided loading of trucks at the shovels.

Hauling

Mill feed and waste will be hauled by 363 t off-highway haul trucks.

Pit Maintenance

Pit maintenance services include haul road maintenance, open pit mine dewatering, relocating equipment, snow removal; as well as distributing operating supplies and fuel from the MTT train freight siding to appropriate location at the mine site.

Haul road maintenance is paramount to low haulage costs; dozer and grader hours have been allocated to maintain the haul road network throughout the LOM production schedule.

A fleet of ancillary service vehicles will be allocated to install and service the in-pit sump pumps and the high wall horizontal drains. This includes connecting these pumps to the pit dewatering pipeline system. This crew will also service and supply mobile light plants.

A fleet of service equipment will be allocated for summer season construction and used in winter for snow clearing and spreading crushed rock for traction control. This includes scrapers and loaders. The snow fleet will be manned by mine operations staff in normal winter conditions with operators taken from reduced activities such as dust control and summer field programs. During severe storms, personnel who operate the standby snow fleet will be drawn from truck and shovel operations as the long-haul fleets shut down. This will ensure priority fleets remain operating. In-depth snow management studies for the surface mining activities have been undertaken and are integrated into this mining plan.

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Rock crushers for road grading material, winter traction control, and blast hole stemming are included.

General Mine Expense

The GME area accounts for the supervision, safety, environment, and training for the direct mining activities as well as technical support from mine engineering and geology functions. Open pit mine operation supervision will extend down to the shift foreman level and trainers.

GME costs also include engineering consulting on an ongoing basis for specialty items such as geotechnical, geo-hydrology expertise, and third-party reviews in the open pit mine area.

Open Pit Mine Equipment

Major Equipment

The peak major mining equipment requirement is summarized in Table 24.7. The current production schedule requires a peak loading fleet with 7 shovels and a peak haulage fleet of 62 trucks.

Table 24.7 Peak Major Equipment Requirements

Area/Type | Description | Units
Drilling
Drill – Diesel Hydraulic – 311 mm | Primary Drill | 7
Loading
Hydraulic Shovel – 40 m3 | Loading Mill Feed & Waste | 7
Hauling
Haul Truck – 363 t | Hauling Mill Feed & Waste | 62

Support Equipment

The peak mine support equipment fleet requirements are summarized in Table 24.8.

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Table 24.8 Peak Mine Support Equipment Fleet

Fleet | Function | Units
Hole Stemmer – 3 t | Blast Hole Stemmer | 2
Track Dozer – 630 kW | Shovel Support | 7
Rubber Tired Dozer – 350 kW | Pit Clean Up | 4
Fuel/Lube Truck | Shovel and Drill Fueling and Lube | 3
Wheel Loader Multipurpose – 14 t | Pit Clean Up | 3
Water Truck – 20,000 gal | Haul Roads Water Truck | 3
Track Dozer – 630 kW | Dump Maintenance | 5
Motor Grader – 400 kW | Road Grading | 5
Tire Manipulator | Tire Changes | 2

Ancillary Equipment

The mine ancillary equipment fleet is listed in Table 24.9.

Table 24.9 Open Pit Ancillary Equipment Fleet

Fleet | Function | Units
Track Dozer – 630 kW | Pit Support | 3
Float Tractor/Trailer – 189 t | Float Tractor and Trailer | 2
Hydraulic Excavator – 6 t | Utility Excavator | 4
Sump Pump - 1,400 gal/min | Pit Sump Dewatering | 6
Light Plant | Lighting Plant | 10
250 t Crane | Utility Crane | 2
Crew Cab | Supervision and Crew Transportation | 20
Ambulance | Ambulance | 1
Hydraulic Excavator – 4 t | Utility Excavator | 6
Mine Rescue Truck | Rescue Truck | 1
Crew Bus | Crew Bus | 6
Maintenance Truck – 1 t | Maintenance Truck | 5
Fire Truck | Fire Truck | 1
Screening & Crushing Plant - 12” max. | Road Crush and Stemming | 2
Picker Truck | Maintenance + Overhauls | 2
Scraper – 37 t | Crush Haul for Winter Roads etc. | 5
Crane 40 t Extendable | Utility Crane | 2
Wheel Loader – 14 t | Crusher (Road Crush) Loader | 1
Snowcat | Winter Off Road Crew Transport | 6
130 t Crane | Utility Crane | 2
Forklift – 30 t | Forklift | 2
Forklift – 10 t | Forklift | 2
Service Truck | Service Truck | 5
Welding Truck | Welding Truck | 4

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Snow Fleet

All of the following snow fleet equipment is chosen to start operating during pre-production and continue to the end of mine life, unless otherwise noted:

| · | five 37 t scrapers are included in the fleet. The scrapers will haul and spread crushed rock for traction control and remove snow from the haul roads and mine working areas as necessary. The scrapers will also be used on occasion for small earthmoving jobs and reclamation works outside of the snow season

| · | one wheel-loader with an approximately 14 t bucket to clear snow from the plant area and truck shop, as well as ancillary routes within the mine. The wheel loader is also used to load the cone crusher at the crushing and screening plant

| · | six snowcats to transport operators to equipment in a location that is inaccessible to the crew bus or vans during heavy snowfall periods.

Open Pit Dewatering Systems

The dewatering equipment will include the following:

| · | horizontal drain holes in bench faces

| · | sloped pit floors as required

| · | in-pit sumps

| · | vertical dewatering wells

| · | a dewatering tunnel behind the north high wall (NPWDA).

Pit water will be collected and pumped to the WSF.

Open Pit Ancillary Facilities

Shops and Offices

In addition to providing an area for maintenance bays, tire shops, and a wash bay, the maintenance shop will also house:

| · | a welding bay

| · | an electrical shop

| · | an ambulance

| · | a first-aid room

| · | a first-aid office

| · | a machine shop area

| · | a mine dry

| · | a warehouse

| · | offices for administration, mine supervision, and engineering/geology staff

| · | a lunchroom and foreman’s office.

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Initial contractor pre-stripping will use temporary maintenance and operations facilities. The permanent facilities will be in place for the Owner’s fleet in Year 1.

The recommended shop sizing for the open pit operations includes eight service bays, one welding bay, and three wash bays. The mine maintenance facility will also include a machine shop area, tool storage area, mine muster area, warehouse, and office complex. A separate tire bay facility will be required with an exterior heated pad to accommodate at least two trucks and a tire manipulator.

PIT SLOPE DESIGN

The 2020 PEA includes open pit slope designs for the Kerr, Sulphurets, and Mitchell zones. The slope design recommendations for the north and south walls of the Mitchell zone have been updated by BGC based on the 2019 Geotechnical Model (BGC, 2020) with the following:

| · | field and laboratory testing work completed in 2009 and 2010, supplemented by further programs in 2011, 2012, and 2018

| · | changes in the geological model since 2010

| · | reduced pit slope heights in the initial mining phases.

The 2020 PEA slope design recommendations for the Mitchell zone are supported by data from:

| · | 31 geotechnical drill holes representing approximately 13,300 m of core logging data

| · | borehole televiewer surveys in the geotechnical holes and some supplementary data from exploration holes

| · | point load testing, uniaxial compressive strength testing, triaxial testing, Brazilian tensile testing, and direct shear testing of natural discontinuities

| · | outcrop mapping and photogrammetric mapping of natural slopes

| · | a network of 24 vibrating wire piezometers providing measurements of pore water pressures.

The rocks of the Mitchell zone are generally strong and of good quality; the exceptions being some altered and weathered zones at shallow depths and the advanced argillic altered rocks of the Snowfield Landslide. The Snowfield Landslide does not have material impact on the Mitchell 2020 PEA pit slopes. The geotechnical model has been updated to include nine geotechnical units representing distinct rock or soil units. The properties of these units have been estimated from the available data.

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The geological structures of Mitchell zone affecting the slope design recommendations are:

| · | regional scale thrust and normal faults. These faults have been mapped and intersected by geotechnical and exploration drilling. Three-dimensional interpretations are available in the geological model

| · | foliation within the sedimentary/volcanic Hazelton and Stuhini groups rocks. The foliation is most common on the south wall of the valley. Foliation is less common on the north wall of the valley above the Mitchell Thrust Fault due to the presence of the P8 intrusive rocks.

The Mitchell zone was divided into seven structural domains representing zones of similar discontinuity patterns. The recommended inter-ramp slope angles vary from 36° to 53° based on overall wall height, wall azimuth, rock mass quality, and geological structures.

The 2020 PEA slope design for the Mitchell pit are as follows:

| · | geotechnical berms, included in the pit slopes at 150 m intervals, were increased from 20 m to are 25 m wide. The increase width provides more mine operations flexibility

| · | the inter-ramp angles for the initial phases of the south wall were increased from 36° to 44°; based on the updates to the geotechnical model and the reduced slope heights of the initial mine phases

| · | the inter-ramp angles for the north wall were reduced from 53° to 48° above the Mitchell Thrust Fault; based on the updated geological structure model and the planned slope heights

| · | the quantity of wells and horizontal drains are reduced, reflecting the reduction in the number of mine phases and pit face area and reduced slope depressurization requirements in some slope sectors. Inflow and quantity estimates for the 2020 PEA open pits are based on extrapolation from prior KSM numerical modelling results and the updated mine plan schedules.

The assumed mine operating procedures are:

| · | slopes will be excavated using controlled blasting

| · | a multi-layer monitoring system consisting of robotic total stations, prisms, radars, and additional state of practice equipment will be in-place

| · | a team of qualified geotechnical engineers will be employed at the mine.

| 24.16.3 | Iron Cap Mining Methods

Mining Method Selection

Based on mining costs, deposit grade, geometry, and depth, block caving was selected as the preferred mining method. Iron Cap has been designed for using battery-electric loaders and electric trains for material handling and employs both tele-operation and automation for these units. This enables a reduction to the number of primary ventilation intake and exhaust drifts due to less ventilation demand and cost savings from lower diesel fuel, labour and equipment maintenance costs.

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Geotechnical And Caving Geomechanics Considerations

The characterization of the Iron Cap rock mass focuses on the rock in and around the extraction and undercut levels of the proposed block cave mine (870 m elevation) and on the mineralized rock above this that will be caved.

Characterization of the rock was based on core photographs and data collected for exploration drill holes, detailed geotechnical data collected for drilling programs carried out by BGC in 2010 (BGC, 2011), and an interpreted geological model provided to Golder by Seabridge.

Three geotechnical holes were drilled in the Iron Cap deposit in 2010 and no additional geotechnical holes have been drilled since then. A qualitative geotechnical assessment of the core from a number of exploration holes drilled in 2017 was undertaken by Golder personnel in September 2017. This indicated that the geotechnical quality of the rock mass was good to very good, and the quality was very uniform in the area that is being proposed to be mined by block caving. The overall conclusion was that the geotechnical assessment undertaken by Golder for previous studies (Golder, 2012) remains valid for this 2020 PEA.

The caveability assessments made using Laubscher’s and Mathews’ methods for prior studies indicated that the size of the footprint required to initiate and propagate caving is less than approximately 200 m. This minimum footprint size is significantly smaller than the size of the footprint of the deposit that can potentially be mined economically by caving. This fact, together with the generally large-size continuous nature of the deposit indicates that the Iron Cap deposit is amenable to block caving.

The cave mining will draw down the mineralized rock and a significant depression will develop on surface above the production footprint in the form of a crater. The crater typically develops on surface above and slightly laterally beyond the footprint of the production horizon of the cave mining. The top section of the crater is a relatively steep escarpment (60° to 70°) that is marginally stable but comprised of nominally in place dilated rock. Deeper within the crater is failed broken rock that has progressively sloughed from the rim of the crater. This rock rills down to the bottom of the crater at angle of repose.

The Iron Cap cave will undercut parts of the overlying ice sheet. The ice will break up as part of the surface disturbance created by the cave and progressively melt as it is drawn down. The total volume of water that enters the cave from the melting ice is relatively small. Further studies will be required in the future to assess the need to drain free water that may be present immediately beneath the ice sheet prior to the start of caving.

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Cave Modelling and Production Scenarios

The NSR block model discussed in Section 24.16.1 was input to Footprint Finder (FF) and used to search for the elevation and profile of the footprint with the highest dollar value. The primary FF input parameters are presented in Table 24.10.

Table 24.10 Iron Cap Footprint Finder Input Parameters (CAD)

Footprint Finder Parameter | Value | Comments
Discount Rate | 5% | Provided by Seabridge
Maximum Height of Draw | 750 m | Industry experience
Minimum Height of Draw | 195 m | From caving geomechanics assessments
Vertical Incremental Capital Cost | Cdn$180,000 per m | Golder benchmark from previous studies
Lateral Incremental Capital Cost | Cdn$2,100 per m2 | Golder benchmark from previous studies
Preliminary Operating Cost (Mine, Mill and G&A) | Cdn$11.50 per tonne | $3.80 for mining (Golder) and $7.70 for processing and general and administration (Seabridge)
Maximum Draw Rate | 80 m per year | From industry experience
Drawpoint Opening Rate | 120 per year | Construction of 10 drawpoints per month – industry experience
Production Ramp up Rate | Avg. 5 Mt per year | Industry experience (3 to 6 Mt per year)
Maximum Production Rate | 90,000 tpd | Achievable rate based on the 80 m/yr. draw rate, 120 drawpoints constructed per year and available footprint area

A series of FF runs were performed at varying NSR shut-off values greater than the preliminary operating cost estimate of Cdn$11.50 per tonne. This cost is comprised of Cdn$3.80 mining cost, and Cdn$7.70 processing and general and administrative (G&A) costs. These costs represent the initial estimates that were made to undertake the FF assessments. As indicated in Table 24.13 and discussed in more detail in Section 24.21.2, the mining operating cost was subsequently refined to Cdn$3.88. As well the process and G&A operating cost was subsequently refined to Cdn$7.30. These differences in operating costs are minor and not considered to represent a material change.

The shut-offs used in the FF assessments ranged from Cdn$16/t to Cdn$24/t. A footprint at elevation 870 m produced the highest dollar value and provided for gravity drainage of inflow water from direct precipitation, surface runoff, and operational water. It was based on an NSR shut-off of Cdn$20/t, with a maximum HOD of 750 m. This generated sufficient tonnes to feed the mill in combination with other mill feed sources over a sustained period and at a production rate of 90 kt/d. A second lower cave lift was assessed but the footprint was not sufficiently large to warrant further study. Figure 24.12 shows the final footprint profile at 870 m elevation. The warmer colours shown in this figure represent higher column values.

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Figure 24.12 Iron Cap Footprint for Lift at 870 m Elevation

| Note: | (Golder 2020) Grid system is based on UTM coordinates. Warmer colours indicate higher column NSR values and cooler colours indicate lower column NSR values.

Dilution

Dilution is defined as any material that has a value less than the operating cost that is mucked at the drawpoints. This includes waste blocks of zero grade and waste density that were conservatively assumed to be available to be drawn down representing failure of the overlying slopes and sidewalls of the crater that report to the drawpoints. Only a small percentage of zero grade blocks contribute to dilution.

The mixing of the economic and sub-economic blocks (including zero grade blocks) as they are drawn down was assessed using the standard algorithms in FF that have been extensively tested and back-analyzed at many mines, and have become a standard approach in estimating drawpoint production and grade schedules for block cave mines.

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Design Criteria and Layout

The block cave design for the Iron Cap footprint is at the 870 m elevation. The shape of footprint is an irregular oval, with an area of 570,000 m2, a length of approximately 1 km, and a width at the widest section of approximately 700 m. As indicated in Table 24.13, the total production is 743 million tonnes. The production tonnes and grade are comprised of mineralized material included in the Mineral Resources reported in Section 14, mineralized material not included in the Mineral Resource statement, and minor tonnages of zero grade material.

The heights of draw through which the production tonnes are extracted range from 250 m to 750 m, where the lower values reflect the western side of the footprint and the higher values reflect the eastern side of the footprint. Most of this geometry is dictated by the position of the Sulphurets Thrust Fault which is the structure which constrains economic mineralization below it.

The development direction for the Iron Cap cave starts from the middle of the eastern lobe and propagates away from this middle to the footprint edge; and also generally from west to east, such that draw bells on the western half of the footprint are developed first and draw mill feed columns whose heights are generally lower than the average height across the footprint.

Personnel, material, and supplies will access the underground through the primary access drift adjacent to the Mitchell OPC. For this design, the MTT has been slightly re-aligned towards the south adjacent to the Iron Cap deposit to avoid intersecting with the very weak STF and the adjacent mylonitic zone. Immediately adjacent to the MTT is a rail spur to allow the haulage trains to be loaded under two surge bins that are fed by the Iron Cap conveyor. The conveyor is fed by two gyratory crushers located just outside the northern end of the footprint.

The haulage level maintenance shops, underground offices and warehouse facilities are located at the north end of the footprint.

Two fresh air portals and two exhaust portals are planned on the north slope of the Mitchell Valley as seen in Figure 24.13. A third fresh air intake adit will be collared in the area of the Mitchell OPC to the southwest of the footprint. It is required to provide a flow-through ventilation circuit during the pre-production development phase before the other drifts can be excavated south from the footprint to breakthrough on the north slope. (Developing the tunnels northward from the north valley slope into the mountain was considered unsafe due to the high geohazard and avalanche risks on this slope). The third adit will remain an active intake and access adit throughout the mine life.

The fresh air adits will connect to perimeter drifts constructed around the entire mine footprint to provide fresh air to the mine workings. Exhaust raises on the extraction level will connect down to an exhaust level that then connects to the two adits that exhaust air to the north slope of the valley. The exhaust portals are downwind of the fresh air portals to avoid intake of air contamination and are separated approximately 340 m from one another.

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The Iron Cap design has multiple drifts that can act as emergency egress. The primary emergency egress will be the train tunnel or fresh air adit of the MTT, whichever is accessible. If both are inaccessible, it will be possible to exit the mine through one of the fresh air drifts that connect to the north slope of the Mitchell Valley.

Figure 24.13 Plan View of Iron Cap Mine with Contour

Source: Golder (2020)

Figure 24.14 shows the typical level arrangement for the Iron Cap mine. The mine development occurs on the following seven levels:

| · | the pre-conditioning level is planned to provide access for in situ fracturing of the rock mass prior to caving. From the grid of drifts on this level, a series of boreholes will be drilled, and hydraulic fracturing will be used to generate fractures in the rock mass that it is proposed to cave

| · | the undercut level is comprised of a series of parallel drifts to perform drilling and blasting required to develop void and initiate caving. The level is located 40 m below the pre-conditioning level

| · | the extraction level is comprised of parallel drifts spaced 18 m apart where the drawpoints and dumping points are located. Battery-electric LHD equipment will extract the caved material from the drawpoints and haul it to the nearest ore pass dump point. This extraction level at Elevation 870 m is 20 m below the undercut level

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| · | the fresh air level is 20 m below the extraction level and conveys fresh air to the other levels via vertical ventilation raises

| · | the exhaust ventilation level is 5 m below the fresh air level and this level exhausts air by vertical ventilation raises from the extraction level

| · | the haulage level is 25 m below the exhaust level. It receives the material dumped by LHD’s into ore passes on the extraction level. Each ore pass from the extraction level is fitted with a stationary rock breaker to size the material that gravity feeds down to chutes that load rail cars. Electric locomotives haul the loaded rail cars to one of two gyratory crushers located just outside the northern end of the footprint

| · | the water drainage level is located 25 m below the haulage level and it collects drainage and inflow water and directs it to the 1,314 m long Dewatering Tunnel that gravity drains to the North Pit Wall Dewatering Adit (NPWDA).

Figure 24.14 Typical Level Arrangement for Iron Cap

Source: Golder (2020)

An El Teniente-type drawpoint layout is planned on the extraction level. The routing for the LHD haulage to the ore passes is shown in Figure 24.15. The design incorporates extraction drifts that are 4.5 m by 4.5 m in cross-section and spaced 30 m apart, with drawbell drifts that are 4.5 m by 4.5 m in cross-section, aligned at a 60° angle to the extraction drifts, spaced 18 m apart. The drawbells are created by drilling and blasting and then two drawpoints per drawbell are constructed. These drawpoints are supported with heavy support in the brows in order to withstand the attrition and erosion of the mineralized material as it is mucked. The resulting drawpoint spacing is 18 m by 15 m which is appropriate for the expected fragmentation.

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Figure 24.15 Drawpoint Configuration for Iron Cap

Source: Golder (2020)

Overall, the mine plan is based on there being 7 different levels and approximately 166 km of drifts and raises. Table 24.11 summarizes the various types of drift development and the sizes of these excavations.

Table 24.11 Main Drift Types and Dimensions

Type | Dimensions
Fresh/Return Air Drifts, Ore Handling Access | 4.0 m W x 4.0 m H
Drawpoint and Extraction Drifts | 4.5 m W x 4.5 m H
Pre-conditioning, Undercut, Dewatering Drifts | 4.5 m W x 4.5 m H
Access Tunnels, Conveyor Drifts, Dewatering Drifts, Perimeter Drifts | 5.0 m W x 5.0 m H
Shops, Conveyor Access | 5.5 m W x 5.5 m H
Haulage Levels, Return Air Drifts | 6.5 m W x 6.5 m H
Main Fresh/Return Air Drifts | 7.0 m W x 7.0 m H

Vertical development will be constructed primarily using raisebore machines and raisebore-and-slash techniques. Table 24.12 summarizes the various types of vertical development and the sizes of these excavations.

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Table 24.12 Vertical Development Dimensions

Type | Diameter (m) | Type
Temporary Ore Bin | 5 | Bored Raise
Ore Passes and Haulage Level Ventilation Raises | 4 | Bored Raise
Footprint, Undercut, and Pre-conditioning Level Ventilation Raises | 2.5 | Bored Raise
Crusher Ore Bin | 6 | Bored Raise and Slashed
MTT Ore Bin | 10 | Bored Raise and Slashed

Ventilation

The quantity of required airflow was determined using a VentSim model of the mine design and was based on the requirement for 0.06 m3/s of ventilating air for each kilowatt of power of diesel-powered equipment and 0.025 m3/s per kilowatt of electrically powered equipment. Mine equipment and engine utilization factors were also considered in determining total ventilation needs.

The peak airflow requirement was estimated to be 1,075 m3/s, which is sufficient to dilute all noxious gases, particulate matter, and heat produced by the mining equipment and the operational activities on each mining level. This equates to a rate of 0.011 m3/s per tonne of material mined per day and is approximately one-half of a typical block cave mine using only diesel-powered equipment. This reduced ventilation demand results in fewer primary intake and exhaust drifts and less main fan capacity and power demand.

There are three main intakes and two main exhausts that are 7 m by 7 m in cross-section. There are 3 axial fans at each of 3 intake drifts with a peak power demand of 1,865 kW. Raises that are 2.5 m in diameter connect the extraction level to the exhaust drifts.

Heating of mine air in the winter months is included in the design and cost estimates, and this will be accomplished using mine heaters located at each of the two main fan installations. A mine air heating system of 15 MW capacity is required to heat 1,075 m3/s of air from an average minimum temperature of -5.7°C to 2°C for an average of six months per year (November through April). The propane requirement is 5.1 million liters per year at peak production. Figure 24.16 is a schematic of the ventilation system showing major airflow directions.

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Figure 24.16 Ventilation Schematic for Iron Cap Mine (Section Looking West)

Source: Golder (2020)

Mine Dewatering

The maximum mine water inflow estimate for Iron Cap is 6.0 m3/s. This is based on factoring the more detailed assessment that was undertaken from first principles. The inflow water includes groundwater and surface water inflows (including ice melt from ice that is drawn down with the caved rock when the ice field is undercut and then melts), and water that is introduced to the mine for operations. The inflow from melting ice will have a relatively insignificant impact on total water inflows. The inflows to Iron Cap are conveyed by gravity via a system of sumps and drainholes, from the various levels, to a 1.3 km drift located on the southwest corner of the footprint that connects to the North Pit Wall Dewatering Adit (NPWDA). The NPWDA flows into the Mitchell Valley Depressurization Tunnel (MVDT) which then flows to the WSF.

All drifts have been graded so that water flows towards collection sumps which will feed into the Dewatering Tunnel. Sump pumps will be used locally as required. The water management system has a nominal capacity of 6 m3/s (95,000 gpm).

Production Schedule

The proposed production schedule as developed using FF software is shown in Figure 24.17. Note that the years shown apply to the combined open pit and underground multi-mine plan of the KSM mine production. Iron Cap pre-production will take 4 years with the first cave production being in Year 6.

The ramp-up period of 6 years is governed by the draw point opening rate of 120 per year and maximum draw rate of 80 m per year (220 mm per day). Both are consistent with industry experience. The sustained peak production rate estimated using FF is 90 kt/d. There is also a 6-year ramp-down period as drawpoints are depleted.

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Figure 24.17 Iron Cap Mine Production Plan

Source: Golder (2020)

Only Indicated and Inferred Mineral Resources are included in the production plan. All unclassified materials are treated as waste by setting the grades to zero. Table 24.13 shows the distribution of production tonnes and grade in the 2020 PEA mine plan.

Table 24.13 Iron Cap Production Tonnes And Grade in 2020 PEA Mine Plan

| Measured | Indicated | Inferred | Total*
Tonnes (millions) | - | 57.7 | 685.0 | 742.7
NSR* (Cdn$/t) | - | 32.83 | 36.70 | 36.36
Au (g/tonne) | - | 0.62 | 0.58 | 0.58
Cu (%) | - | 0.28 | 0.36 | 0.35
Ag (g/tonne) | - | 3.1 | 3.0 | 3.0

* Note: Includes 18.5 Mt of mineralized dilution (the portion of Indicated and Inferred material that is below the final mining operating cost of Cdn$3.88 discussed in Section 24.21.2 plus process and G&A operating cost of Cdn$7.30 [Cdn$0 < NSR < Cdn$11.18]) and includes unclassified at zero grade.

The mining sequence supporting the production plan is shown in Figure 24.18. The cave is initiated in the region of the footprint that has the highest NSR values to maximize early value. The mining then propagates radially from this point and continues down the southeastern limb of the footprint.

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Figure 24.18 Proposed Drawpoint Development Sequence for the Iron Cap Mine

Note: Legend is drawpoint number Source: Golder (2020)

Development Schedule

The pre-production development phase of the Iron Cap mine is estimated to be 4 years, which includes mine access, some initial footprint development and construction of major mine infrastructure, such as the underground crusher, material handling system, shops, dewatering system, and primary ventilation. The development schedule was generated using Deswik software based on an advance rate of 4 m/day and a maximum of 3 active headings in the first two years, 6 active headings in Year 3, and 9 active headings in Year 4.

The overall development strategy is to provide access to the crusher location as soon as possible while prioritizing the installation of the materials handling system, haulage system and shop areas. Following this, extraction drifts, drawpoints, and undercut and pre-conditioning levels will be advanced to allow for cave initiation. Several ventilation airways will also be advanced southward towards the Mitchel Valley to establish the flow-through ventilation circuit in advance of production.

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Blasting and Explosives

It is assumed that explosives will be stored in a surface magazine in the Mitchell Valley to service the Iron Cap mine. The explosives will be delivered to the underground mine via the access tunnel as required. Small underground magazines will be established to provide daily supplies of explosives, and explosives will be distributed on a shift basis for five main purposes:

| · | preparation for the development of mineralized areas

| · | draw bell construction

| · | undercutting

| · | development in barren material

| · | secondary blasting of boulders and hang ups.

Mining Equipment

Primary and support equipment requirements for the proposed Iron Cap mine, at full production in KSM operation Year 14, are summarized in Table 24.14.

Table 24.14 Mobile Equipment Requirements for Iron Cap (KSM operation Year 14)

Development Equipment | Unit | Production Equipment | Unit | Support Equipment | Unit
Jumbo | 2 | LHD (BEV) | 18 | Grader | 2
Haul Truck (BEV) | 2 | Emulsion Loader | 1 | Lube Truck | 3
LHD (BEV) | 1 | Drill Rig | 2 | Electricians Truck | 3
Bolter | 1 | Shotcrete Sprayer | 1 | Telehandler | 4
ANFO Charger | 1 | Transmixer | 1 | Mobile Crane | 2
Scissor Lift | 1 | Bolter | 1 | Mine Ambulance | 1
Shotcrete Sprayer | 1 | Water Cannon | 4 | Fire Truck | 1
Transmixer | 1 | Block Holer | 8 | Mine Rescue Truck | 1
Contract Raisebore | 1 | Raisebore | 1 | Fan/Pipe Handler | 1
Rehab Equipment | Unit | Scissor Lift | 1 | Backhoe | 3
Jumbo | 1 | Boom Truck | 1 | Water Truck | 1
ANFO Charger | 1 | Locomotive (Electric) | 5 | Sludge Truck | 1
Shotcrete Sprayer | 1 | Shunting Locomotive (Electric) | 1 | Cable Bolter | 3
Transmixer | 1 | Rail Car | 72 | Big Personnel Carrier | 6
Bolter | 1 | Mobile Rockbreaker | 3 | Small Personnel Carrier | 16

Note: BEV is battery-electric vehicle

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Mine Workforce

The estimated workforce for Iron Cap is 454 persons at the peak of the pre-production construction period with an average of 489 persons during the peak production period.

| 24.16.4 | Deep Kerr Mining Methods

Mining Method Selection

Based on mining costs, deposit grade, geometry and depth, block caving was selected as the preferred mining method. Deep Kerr has been designed for using diesel powered loaders and haulage trucks for material handling.

Geotechnical And Caving Geomechanics Considerations

Dedicated geotechnical drilling or studies have not yet been conducted for the Deep Kerr resource area. Current geological exploration drilling indicates similar rock conditions to the Mitchell underground mine, and for the purposes of the 2020 PEA it is assumed the rock characteristics will be similar.

The caveability assessments made using Laubscher’s and Mathews’ methods indicate that the size (diameter) of the footprint required to initiate and propagate caving is less than 220 m. This minimum footprint size is significantly smaller than the size of the footprint of the deposit that can potentially be mined economically by caving. This fact, together with the general large-sized, continuous nature of the deposit, suggests that the Deep Kerr deposit is amenable to cave mining.

The cave mining will draw down the mineralized rock and a significant depression will develop on surface above the production footprint in the form of a crater. The crater typically develops on surface above and slightly laterally beyond the footprint of the production horizon of the cave mining. The top section of the crater is a relatively steep escarpment (60° to 70°) that is marginally stable but comprised of nominally in place dilated rock. Deeper within the crater is failed broken rock that has progressively sloughed from the rim of the crater. This rock rills down to the bottom of the crater at angle of repose.

GEOVIA’s PCBC™ Footprint Finder software was used to identify the elevation and shape of the highest value footprints. A maximum column height (height of draw, HOD) of 750 m was used to identify and locate the footprint for the upper lift, and 500 m was used for a subsequent lower lift. Mines such as the Cadia East Mine in Australia, which is currently in production based on columns up to 1,000 m, have developed technical and operational methods for accommodating such high column heights.

Life of mine underground infrastructure such as conveyors, vent drifts, ramps, and vent shafts are not planned within or near any areas potentially disturbed by caving.

Cave Modelling and Production Scenarios

The NSR block model discussed in Section 24.16.1 was input into the FF software and used to search for the elevation and profile of the footprint with the highest dollar value. Only Indicated and Inferred Mineral Resource classes were included in the assessment.

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The proposed Kerr Pit was also included. The primary FF input parameters are presented in Table 24.15.

Table 24.15 Deep Kerr Footprint Finder Input Parameters (CAD)

Footprint Finder Parameter | Value | Comments
Discount Rate | 5% | Provided by Seabridge
Maximum Height of Draw | 750 m | 750 m used to define footprint and for production schedule, based on industry experience
Minimum Height of Draw | 195 m | From experience
Minimum Allowable Footprint Width | 200 m | Based on geotechnical/caveability assessments
Vertical Incremental Capital Cost | Cdn$180,000 per m | Golder benchmark from previous studies
Lateral Incremental Capital Cost | Cdn$2,100 per m2 | Golder benchmark from previous studies
Preliminary Operating Cost (Mine, Mill and G&A) | Cdn$13.60 per t | Cdn$6.00 for mining (Golder) and Cdn$7.60 for processing and general and administration (Seabridge)
Maximum Draw Rate | 80 m per year | From industry experience
Drawpoint Opening Rate | 120 per year | Construction of 10 drawpoints per month – industry experience
Production Ramp up Rate (PRC) | Avg. 5 Mt per year | Industry experience (3 to 6 Mt per year)
Maximum Production Rate | 80,000 tpd | Achievable rate based on the 80 m per year draw rate, 120 drawpoints constructed per year and available footprint area

The FF analysis was performed using a preliminary operating cost shut-off of Cdn$13.60/t (Cdn$6.00 mining and Cdn$7.60 processing and G&A), and for a combined 2-lift operation. These operating costs represent the initial estimates that were made to undertake the FF assessments. As indicated in Table 24.18 and discussed in more detail in Section 24.21.2, the mining operating cost was subsequently refined to Cdn$6.21. As well the process and G&A operating cost was subsequently refined to Cdn$7.20. These differences in operating costs are minor and not considered to represent a material change.

The maximum dollar value was obtained for production levels at elevations of 580 m and 85 m. These levels are 45 m lower than the 625 m and 130 m elevations incorporated into the 2016 PEA that allowed for free draining of water inflows from the upper lift. This is a significant advantage since it removes the need for a large pump bank on the first lift. The maximum dollar values were not significantly different for lifts at the slightly higher elevations, less than 2%, so it was decided to maintain the 625 m top elevation and the 130 m elevation for the second lift, 495 m below.

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Using the production level elevations of 625 m and 130 m, various column shutoff values based on NSR block values were then used to determine and optimize the footprint shapes. A shutoff value of Cdn$18/t was used for the final selected two-lift case. This generated sufficient tonnes to feed the mill in combination with other mill feed sources over a sustained period and at a production rate of 80 kt/d. The footprints for these two levels are shown in Figure 24.19 and Figure 24.20. The warmer colours shown in these figures represent higher column values. The final footprint boundaries maintain a minimum footprint width of 200 m to initiate and propagate caving (grid spacing on plan is 200 m).

Figure 24.19 Plan View of Deep Kerr 625 m Lift 1 Footprint.

Golder (2020) Grid system is based on UTM coordinates. Warmer colours indicate higher column NSR values and cooler colours indicate lower column NSR values.

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Figure 24.20 Plan View of Deep Kerr 130 m Lift 2 Footprint.

Golder (2020) Grid system is based on UTM coordinates. Warmer colours indicate higher column NSR values and cooler colours indicate lower column NSR values.

Dilution

Dilution is defined as any material that has a value less than the operating cost that is mucked at the drawpoints. This conservatively includes blocks of zero grade and waste density that were assumed to be available to be drawn down representing failure of the overlying slopes and sidewalls of the crater that fail onto the caved rock. Only a small percentage of zero grade blocks contribute to dilution.

The mixing of the economic and sub-economic blocks (including zero grade blocks) as they are drawn down was assessed using the standard algorithms in FF that have been extensively tested and back-analyzed at many mines, and have become a standard approach in estimating drawpoint production and grade schedules.

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Design Criteria and Layout

The block cave mine design is based on two separate lifts with a combined area of 775,000 m2. Lift 1 is at the 625 m elevation and Lift 2 is at the 130 m elevation. As indicated in Table 24.18, the production from Lift 1 is 525 million tonnes and the production from Lift 2 is 300 million tonnes. Lift 1 tonnes are based on a starting topography that has the Kerr pit exploited. The production tonnes and grade are comprised of mineralized material included in the Mineral Resources reported in Section 14, mineralized material not included in the Mineral Resource statement, and minor tonnages of zero grade material.

Lifts 1 and 2 are both accessed through a series of shared portals and ventilation adits along the northwestern base of the mountain, facing Sulphurets Creek. All lifts connect to a 7 km common inclined UG conveyor system which passes under the Sulphurets Creek and connects directly to an ore bin above the MTT material handling system which supports all mines at the KSM site. Mining on each lift assumes panel caving with a post undercutting method, with load-haul-dump (LHD) equipment for extraction, and a material handling system of muck passes, chutes, truck haulage, two crushers and conveyors.

The ventilation system design assumes five main intake adits and four main exhaust adits. Lift 2 is connected to the common ventilation adit system via bored shafts and decline ramps. Two of the main intake adits also serve as the primary access for the mine and are connected to the decline ramp system for the subsequent lower lift.

Surface camps, shops, warehouses, offices, electrical substation, water distribution, and other surface infrastructures are located just outside the portal area just below 600 m elevation. These facilities will be used in conjunction with similar facilities established for the overall KSM area to support the other mining operations, i.e., Mitchell, Iron Cap, Sulphurets, etc. Figure 24.21 shows a plan view of the mine layout.

The Deep Kerr design has multiple drifts that can act as an emergency egress. The primary emergency egress will be one of the fresh air adits. If these adits are inaccessible, it will be possible to exit the mine through the service adit. Egress from Lift 2 will be up the ramp or via the service ramp parallel to the conveyor decline.

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Figure 24.21 Deep Kerr Mine Plan View

Note: Spacing between Lifts 1 and 2 is 495 m. The distance from the portals to the footprint of Lift 1 is approximately 2 km. Source: Golder (2020)

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Lift 1 has an area of 446,000 m2. The footprint shape is an irregular oval shape spanning 1,290 m in length with variable widths of up to 500 m, tapering at both ends. The extraction level is at elevation 625 m, which is slightly higher than the portal elevation of 600 m and thus does not require major ramps or shafts for access or ventilation. Mine access is primarily through two of the intake portals, one of which directly connects to the footprint area as a main ventilation intake drive, and the other which eventually parallels the UG conveyor as a service decline that connects to the crusher area. The total production tonnage from lift 1 will be 525 million tonnes. Figure 24.22 shows the general arrangement of Lift 1.

Figure 24.22 Deep Kerr Lift 1 (625L Mine Area) General Arrangement

Source: Golder (2020)

Lift 2 is smaller in tonnage and footprint size than Lift 1 with an area of 329,000 m2. The footprint shape is an irregular and approximately 1,080 m in length with widths varying from 200 m up to 700 m, tapering at both ends. The extraction level is at elevation 130 m. It requires two decline ramps for access and muck conveyance, and seven bored shafts for ventilation. The total production tonnage from Lift 2 will be 300 million tonnes. Figure 24.23 shows the general arrangement of Lift 2.

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Figure 24.23 Deep Kerr Lift 2 (130L Mine Area) General Arrangement

Source: Golder (2020)

Figure 24.24 shows the typical level arrangement for the Deep Kerr mine.

Figure 24.24 Typical Level Arrangement for Deep Kerr

| Note: | Light blue parallel undercut drifts and red parallel extraction panel drifts are spaced 30 m apart. Dark blue drawpoints are spaced 18 m apart. Source: Golder (2020)

The underground mine development occurs on the following seven levels for each lift:

| · | the pre-conditioning level is comprised of parallel drifts from which hydrofracking is performed to improve fragmentation. It is accessed by one perimeter drift on the west side only

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| · | the undercut level is comprised of a series of parallel drifts to perform drilling and blasting required to develop a void and initiate caving. The undercut level is 40 m below the preconditioning level

| · | the extraction level is comprised of parallel drifts spaced 18 m apart where the draw points and dumping points are located. Load-Haul-Dump (LHD) equipment extract the caved material from the draw points and haul it to the nearest dumping points for the ore pass. The extraction level is 17 m below the undercutting level

| · | the exhaust ventilation level is 22 m below the extraction level and this level exhausts air from the other levels via vertical ventilation raises from the extraction level

| · | the intake, or fresh air, ventilation level is 16 m below the exhaust level and conveys fresh air to the other levels via vertical ventilation raises

| · | the haulage level is 17 m below the fresh air level. It receives the material dumped by diesel powered LHD’s into ore passes on the extraction level. Each ore pass is fitted with a stationary rockbreaker to size dumped material and a loading chute at the bottom to enable truck loading. Diesel haul trucks transport the material to one of two gyratory crushers located just outside the footprint

| · | the dewatering level collects all drainage water via gravity through a series of sumps and boreholes connecting all the levels. The dewatering level is 16 m below the haulage level.

On the extraction level, an El Teniente-type drawpoint layout is planned. The routing for LHD haulage is shown in Figure 24.25. The design incorporates extraction drifts that are 4.5 m x 4.5 m in cross-section and spaced 30 m apart, and drawbell drifts that are 4.5 m x 4.5 m in cross-section, aligned 60° to the extraction drifts, and spaced 18 m apart. The drawbells are created by drilling and blasting and once the drawbells connect to the undercut level, two drawpoints per drawbell are constructed. These drawpoints are supported with heavy support in the brows to withstand the attrition and erosion from the mineralized material as it is mucked. The resulting drawpoint spacing is 18 m by 15 m appropriate for the expected estimated fragmentation sizing of the mill feed.

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Figure 24.25 Drawpoint Configuration for Deep Kerr

Source: Golder (2020)

Overall, the mine plan assumes that to produce from Lifts 1 and 2 there will be 14 different levels and approximately 230 km of drifts and raises. Table 24.16 outlines the main drift types and dimensions used in the mine design.

Table 24.16 Main Drift Types and Dimensions

Type | Dimensions
Fresh/Return Air Cut-outs, Undercut, Pre-conditioning | 4.0 m W x 4.2 m H
Drawpoint Drifts | 4.5 m W x 4.2 m H
Extraction Drifts | 4.5 m W x 4.5 m H
Access Ramps, Dewatering Drifts, Perimeter Drifts | 5.0 m W x 5.5 m H
Haulage Levels | 5.4 m W x 6.1 m H
Conveyor Access | 6.0 m W x 5.4 m H
Main Fresh/Return Air Drifts, Service Drift | 6.0 m W x 7.0 m H

Vertical development will be constructed primarily using raisebore machines and raisebore-and-slash techniques. Table 24.17 summarizes the various types of vertical development and sizes.

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Table 24.17 Vertical Development Dimensions

Type | Dimensions | Type
Intake Raises | 3.5 m Diam. | Bored Raise
Ore Passes | 3.5 m Diam. | Bored Raise
Exhaust Raises | 2.4 m Diam. | Bored Raise
Crusher Ore Passes | 5 m Diam. | Bored Raise
Main Ventilation Raises | 5 m Diam. | Bored Raise

Ventilation

The quantity of the required airflow was determined using a VentSim model and was based on the current mine design indicated earlier and 0.06 m3/s of ventilating air for each kilowatt of diesel-powered equipment. Mine equipment and engine utilization factors were also considered in developing the total ventilation needs.

The peak ventilation requirement was estimated to be 1,629 m3/s, which is sufficient to dilute noxious gases, particulate matter, and heat produced by the mining equipment and the operational activities on each mining level. Ventilation is provided by four main intake drifts and four main exhaust drifts 6 m wide and 7 m high The second lift is ventilated via 3 fresh air raises and 4 exhaust raises of 5 m diameter which connect to the Lift 1 drifts, and to the fresh air main ramp and fresh air service ramp.

Heating of mine air in the winter months is included in the design and cost estimates, and this will be accomplished using mine heaters located at each of the two main fan installations. A mine air heating system of 18 MW capacity is required to heat 1,629 m3/s of air from an average minimum of -5.7°C to 2°C for six months per year (November through April). The propane requirement is 7.8 million liters per year at peak production.

A ventilation schematic is shown in Figure 24.26.

Figure 24.26 Ventilation Schematic

Source: Golder (2020)

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Mine Dewatering

The water inflow rate for a 1 in 200-year storm event is predicted to be approximately 5 m3/s for Lift 1 and 9.0 m3/s at the completion of Lift 2. Other inflows that need to be dealt with are small quantities of groundwater and the production process water from drilling and dust suppression activities.

For Lift 1 most of the inflow water will be intercepted on the production level and routed away from the footprint to flow by gravity along the ventilation drifts to the portal at approximately 600 elevation. From here it will be directed into a 120 cm (48 inch) surface pipeline (separate from dewatering pipelines utilized for the Kerr open pit). This pipeline will convey the water to the WTP platform at 543 elevation by gravity and then it will be pumped up to the WSF. The 120 cm pipeline to the WTP is designed to handle 4 m3/s flow and any additional inflows that occur under rare extreme seasonal conditions will be treated as an emergency discharge.

For Lift 2, it is planned to intercept inflow water as it enters the production level and route it to sixteen 3,700 kW pumps (approximately 60,000 kW [80,000 HP]). These pumps have a capacity of 8 m3/s (65,000 gpm) to pump water up to the Lift 1 ventilation drifts from where it will flow under gravity to the ventilation drift portals. From the portals, as for Lift 1, water will be directed into two 120 cm surface pipes (an additional one added for Lift 2 mining). These pipes will convey the water to the WTP the same as for Lift 1. The estimated inflow rate over a 20-year period was used to determine an average annual pump utilization factor of 3.5% which was used to calculate the annual pump power requirements. The mine dewatering general arrangement is shown in Figure 24.27.

Figure 24.27 Mine Dewatering General Arrangement

Source: Golder (2020)

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Production Schedule

The proposed production schedule for Deep Kerr, as developed in FF software, is shown in Figure 24.28. Note that the years shown apply to the overall KSM mine production schedule.

The ramp-up period is 5 years and is governed by the draw point opening rate of 120 per year and maximum draw rate of 80 m per year (220 mm per day). Both are consistent with industry experience. The total peak production rate of 80,000 tpd is achievable based on the available footprint area. Starting around Year 29 there is a dip in total production as lift 1 depletes and Lift 2 commences. This lower production interval is the result of a 60-degree offset applied to the Lift 2 mining relative to Lift 1 to ensure that the Lift 1 crusher and associated access drifts on 625 m level are not impacted by caving from Lift 2. There is an additional 5 years (about 53 Mt) of ramp down production from the second lift that was removed from the mine plan due to limitations imposed by the KSM tailings deposition plan.

Figure 24.28 Deep Kerr Mine Production Plan

Source: Golder (2020)

Only Indicated and Inferred Mineral Resources are included in the production plan. All the unclassified materials are treated as waste by setting the grades to zero. Table 24.18 shows the production tonnes and grade in the 2020 PEA mine plan based on their classifications.

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Table 24.18 Deep Kerr Production Tonnes and Grade in 2020 PEA Mine Plan

Lift 1 625RL | Measured | Indicated | Inferred | Total*
Tonnes (millions) | - | 47.9 | 477.4 | 525.3
NSR (Cdn$/t) | - | 34.03 | 33.76 | 33.74
Au (g/tonne) | - | 0.25 | 0.30 | 0.27
Cu (%) | - | 0.53 | 0.48 | 0.44
Ag (g/t) | - | 1.3 | 1.7 | 1.6
Lift 2 130RL | Measured | Indicated | Inferred | Total*
Tonnes (millions) | - | - | 299.7 | 299.7
NSR (Cdn$/t) | - | - | 35.76 | 35.76
Au (g/tonne) | - | - | 0.33 | 0.33
Cu (%) | - | - | 0.49 | 0.49
Ag (g/t) | - | - | 1.8 | 1.8
Combined | Measured | Indicated | Inferred | Total*
Tonnes (millions) | - | 47.9 | 777.2 | 825.1
NSR (Cdn$/t) | - | 34.03 | 34.53 | 34.31
Au (g/tonne) | - | 0.25 | 0.31 | 0.29
Cu (%) | - | 0.53 | 0.49 | 0.46
Ag (g/t) | - | 1.3 | 1.7 | 1.6

* Note: Includes 124 Mt of mineralized dilution (the portion of Indicated and Inferred material that is below the final mining operating cost of Cdn$6.21 discussed in Section 24.21.2 plus process and G&A operating cost of Cdn$7.20 [Cdn$0 < NSR < Cdn$13.41]) and includes unclassified at zero grade.

The mining sequence supporting the production is initiated on Lift 1 at a location where NSR values are the highest so that early value is maximized. The mining then propagates radially from this location and continues down the southern limb of the footprint. Lift 2 mining starts on the northern end and propagates southward to allow for concurrent production from both lifts.

Development Schedule

The pre-production phase of the Deep Kerr mine is estimated to be 5 years, which includes mine access, some initial footprint development, and construction of major mine infrastructure such as an underground crusher, material handling system, shops, dewatering system and primary ventilation. The development schedule was generated using Deswik software, and it assumes advance rates of 4 m/day and a maximum of 6 active headings in the first 2 years and 10 active headings thereafter.

The overall development strategy is to provide access to the planned crusher location as soon as possible and to prioritize the development of the service drift and MTT conveyor drift. Following this, extraction drifts, drawpoints, undercut, and pre-conditioning levels will be advanced to allow for cave initiation. Two intake and two exhaust airways will be established during pre-production with the remaining airways driven as production requires.

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Blasting and Explosives

Explosives will be received from suppliers and stored in the main surface explosive magazine. The bulk of the material will be ANFO for horizontal development and emulsion for up-holes, plus initiators and accessories.

Underground explosives magazines will be constructed for both Lift 1 and Lift 2 to service development and secondary reduction needs. Each magazine will have a one-week capacity.

Explosives will be used underground for five main purposes:

| · | preparation for the development of mineralized areas

| · | draw bell construction

| · | undercutting

| · | development in barren material

| · | secondary blasting of boulders and hang ups

Mining Equipment

Primary and support equipment requirements at full production in Year 17 are summarized in Table 24.19.

Table 24.19 Mobile Equipment Requirements for Deep Kerr (Year 17)

Development Equipment | Unit | Production Equipment | Unit | Support Equipment | Unit
Jumbo | 1 | LHD | 20 | Electricians Truck | 2
Haul Truck | 1 | Truck | 20 | Grader | 3
LHD | 1 | Emulsion Loader | 1 | Lube Truck | 3
Bolter | 1 | Drill Rig | 3 | Mobile Crane | 1
ANFO Charger | 1 | Shotcrete Sprayer | 1 | Mechanics Truck | 2
Scissor Lift | 1 | Transmixer | 1 | Mine Ambulance | 1
Shotcrete Sprayer | 1 | Bolter | 1 | Fire Truck | 1
Transmixer | 1 | Block Holer | 7 | Mine Rescue Truck | 1
Rehab Equipment | Unit | Mobile Rockbreaker | 4 | Backhoe | 3
Jumbo | 1 | Water Cannon | 4 | Water Truck | 1
ANFO Charger | 1 | Raisebore | 1 | Sludge Truck | 1
Shotcrete Sprayer | 1 | Scissor Lift | 1 | Cable Bolter | 3
Transmixer | 1 | Boom Truck | 1 | Telehandler | 3
Bolter | 1 |  |  | Small Personnel Carrier | 10

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Mine Workforce

The estimated workforce for Deep Kerr is approximately 391 persons at the peak of the pre-production construction period, and 641 persons at the peak of Lift 2 development period.

| 24.16.5 | Mine Production Schedule (Open Pit and Underground Combined)

The summarized production schedule results are shown in Table 24.20 and Figure 24.29, including both open pit and underground mining. The mine production plan starts using conventional large-scale equipment before transitioning into block cave underground bulk mining. The pits at Mitchell, Sulphurets, and Kerr, and underground operations at Iron Cap and Deep Kerr are sequenced from higher grade and lower development and operating cost areas first (higher profitability to lower) to provide optimal cash flow. Mitchell pit with its lower preproduction and development capital cost is scheduled to provide first mill feed followed by the underground operations at Iron Cap and then Deep Kerr.

The production from the block caves will ramp up and then ramp down towards the end of their life. Deep Kerr block cave ramp up begins after Iron Cap ramp up is completed to avoid overlap of block cave development. The pit mining and stockpile reclaim is then scheduled to augment the mill feed up to the targeted mill through-put rate.

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Table 24.20 Summarized Production Schedule – Open Pit and Underground

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Figure 24.29 KSM Mill Feed Production Schedule

Source: MMTS (2020)

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| 24.17 | Recovery Methods

The proposed KSM plant will have an average process rate of 170,000 t/d. The process plant will receive mill feed from the Mitchell, Kerr and Sulphurets open pits and the Iron Cap and Deep Kerr underground mines. The Mitchell open pit and Iron Cap and Deep Kerr underground mines will be the main source of mill feed, contributing about 92% of the total plant feed over the LOM, as follows:

| ● | 25.8% Open pit – Mitchell

| ● | 34.7% Underground mine - Deep Kerr

| ● | 31.3% Underground mine - Iron Cap.

Based on the available information, test work performed on various samples provides a reasonable indication of the mineralogical characteristics of the material examined in this 2020 PEA for both the surface and underground mine extraction scenarios. The process flowsheet considered in the prior studies is also appropriate for the current 2020 PEA given the nature of the mill feed and the level of sampling performed through the deposits.

Several metallurgical test programs have been carried out to assess the recoverability of copper, gold and silver values using the PEA flowsheet. The results of the programs performed on KSM’s mineralized material indicate that the mineral samples from the Mitchell, Sulphurets, Iron Cap and Deep Kerr deposits are amenable to flotation followed by a cyanide leaching process.

Detailed characterization and metallurgical test work on Mitchell, Sulphurets, Iron Cap, and Deep Kerr samples are presented in the pertinent test work reports, which are summarized in Section 13.4, with recent test work presented in Section 13.5. This test work examined comminution (crushing and grinding), concentration by flotation of copper and gold into a saleable concentrate, and the further recovery of gold from flotation tailings by leaching. Metallurgical performance projections from this test work were produced by Tetra Tech and presented in Section 13.7. Wood reviewed this test work and considers it adequate for the purposes of reference for the 2020 PEA. The recent test work conducted after 2016 PEA have been used to improve the 2020 PEA design with two separate mineral process lines in one plant, which is a more refined mineral process plant design than 2016 PEA. Copper recovery models for Iron Cap and Kerr showed clear relationship of copper flotation recovery to copper head grade, respectively. These results can be considered as confirmatory. For gold, there was no clear relationship between gold recoveries vs. copper recovery or gold head or copper head grades.

Comminution test work results indicate that the samples from all the deposits are moderately hard for SAG and ball milling. Comminution tests also showed that the samples tested were amenable to particle size reduction by the HPGR approach as preparation for ball mill grinding.

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Flotation test work (batch and locked cycle) indicates that the mineralization is amenable for concentration into a saleable copper-gold concentrate with no significant penalty elements. Following flotation, cyanidation tests showed that it was possible to recover gold and silver bullion by CIL leaching of gold-rich pyrite concentrates derived from pyrite flotation of rougher flotation tailing and tailing from first cleaner scavenger separation. Trade-off studies indicate that cyanidation of Kerr and Iron Cap gold-bearing pyrite materials may not be economic due to elevated copper/gold ratios and the resulting ineffective CIL carbon absorption. Additional test work and economic trade offs are required to validate this observation. Cyanidation of gold bearing pyrite materials are therefore limited to Mitchell and Sulphurets for the 2020 PEA. The metallurgical test results obtained from the various test programs have been used to predict plant metallurgical performance parameters for copper, gold and silver. The metallurgical performance projections of the four KSM mineralized zones are summarized in Section 13.7. In addition, work has been performed to determine the consumption of reagents and grinding media in comminution, flotation and cyanidation. Test work on cyanide recovery was also performed and the results incorporated into the operating costs.

For the purpose of the 2020 PEA, the process circuit incorporates three stage crushing, one stage milling, conventional flotation, and cyanidation processes for the recovery of copper, gold and silver.

The overall process flow diagram developed for the 2020 PEA relative to prior studies includes these changes:

| ● | elimination of the molybdenum recovery circuit as the economic benefits of molybdenum extraction are not material to the 2020 PEA economics

| ● | the process plant has been split into two distinct parallel process lines with the same layout and equipment to enable separate treatment of high and low copper mill feed from different mining areas. Mitchell and Sulphurets mill feed are processed through dedicated lines separate from Iron Cap and Deep Kerr mill feed.

| ● | an increase in equipment sizing to larger and proven units available in the market to process a higher throughput; 170 ktpd for the 2020 PEA vs 130 ktpd described in the approved EA.

| ● | optimization of the plant footprint and energy consumption for the tonnes processed

| ● | redesign of the facilities has been limited to optimizing the layout provided by the use of the larger equipment in the 2020 PEA.

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| 24.17.1 | Process Plant

The process plant is designed appropriately in accordance with the test work and will consist of five separate facilities for the handling and processing of mill feed:

| ● | primary crushing and a handling facility at the Mitchell Ore Processing Center (OPC)

| ● | truck delivery of mill feed from Sulphurets and Kerr open pit mines to the Mitchell OPC for primary crushing

| ● | conveyor transport of primary crushed mill feed from the Mitchell pit area and the Deep Kerr and Iron Cap underground mines to train loading hoppers

| ● | train transportation of mill feed through the MTT to the process plant

| ● | the process plant is split into two lines which are able to separately process Mitchell/Sulphurets and Iron Cap/Kerr feeds. Each line includes:

| o | Coarse mill feed stockpiles

| o | Secondary crushing

| o | Fine mill feed stockpiles

| o | Tertiary crushing

| o | Ball mill grinding

| o | Copper Flotation and regrinding

| o | Pyrite flotation and regrinding

| o | CIL.

| ● | copper concentrate dewatering – common system

| ● | doré production – common system

| ● | cyanide destruction and tailing disposal – common system

| ● | tailing treatment prior to deposition in the TMF.

| 24.17.2 | flowsheet description

Comminution

The primary crushing facilities are located at the Mitchell OPC, which will reduce the Mitchell ROM particle size to approximately 80% passing 150 mm using two 1,524 mm x 2,260 mm gyratory crushers. Mill feed is trucked from the Sulphurets open pit mine site to the Mitchell OPC for crushing. Similarly, mill feed from the Kerr open pit mine site is trucked to the Mitchell OPC for crushing. Deep Kerr and Iron Cap mill feed is crushed underground and transported separately to the MTT system via different conveyor systems.

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Crushed mill feed is conveyed and loaded onto the MTT train for transport to the main process plant site located at Treaty OPC, approximately 23 km northeast of the mine site.

Mill feed from the mine site through the MTT tunnel is delivered to two parallel process lines. In the early years, Mitchell will supply both process lines, and at the onset of mining of underground material from Iron Cap and Deep Kerr, one process line will be fed Mitchell/Sulphurets mill feed and the other will be fed Iron Cap/Kerr mill feed. This is to enable the continued use of the CIL when processing Mitchell and Sulphurets mill feed, and bypassing CIL when processing Kerr and Iron Cap mill feed. There will be two coarse mill feed stockpiles, each with a live capacity of 30,000 tonnes. The reclaimed feed material from each stockpile will be conveyed to the secondary crushing facility and fed to splitter chutes, which feed two vibrating screens. The oversize from each screen will feed two MP1250 cone rushers. The cone crusher product will be returned to the screen feed conveyor. Screen undersize product that is finer than 50 mm will be delivered to one of the two enclosed fine mill feed stockpiles by the conveyor; each stockpile will have a 30,000-tonne live capacity.

Crushed mill-feed material from secondary crushing will be reclaimed from the fine mill feed stockpile by the reclaim apron feeders onto two HPGR feed conveyors. These conveyors will deliver the mill-feed material to two tertiary crusher HPGR feed surge bins.

Reclaimed mill-feed material will be further crushed by four HPGR crushers; two for each process line. Two belt feeders will withdraw the reclaimed mill-feed material from each of the two HPGR feed surge bins and feed each of the four HPGR crushers separately.

Each HPGR crusher is in closed circuit with double deck vibrating screens with the HPGR discharge wet-screened at a cut size of 6 mm. The screen oversize will return, via conveying, to the HPGR feed bin while the screen undersize will leave the crushing circuit and report to the ball mill grinding circuits. The four HPGR crushing lines will produce a nominal process feed to the ball mills of 7,535 t/h.

The grinding circuit will employ four conventional ball mills in two process lines to grind the HPGR product to a particle size of 80% passing 125 µm to 150 µm. Each ball mill will be in closed-circuit with a cluster of cyclones. The cyclone underflow will gravity-flow to the ball mill feed chute, while the overflow will gravity flow to two copper-gold rougher flotation lines. The overall capacity of the grinding circuits is designed to have a nominal processing rate of 7,535 t/h.

Flotation

The products from the primary grinding circuits will feed the copper-gold rougher/scavenger flotation circuit. This circuit is composed of two banks of four 600 m3 flotation tank cells. The feed density is nominally 37 wt%. The latest test work on Iron Cap/Kerr indicates the possibility of improved copper recovery with a feed density of 22 wt% feed. There is an allowance in the layout for two extra 600 m3 flotation tank cells for the Iron Cap/Kerr line, should they be required, and can be installed in Year 6, ready for the commencement of Iron Cap/Kerr mill feed processing in Year 7. The cells of each bank will produce copper rougher flotation concentrates (at an overall mass pull of 6%) which will then be reground to a particle size of 80% passing 20 µm in two tower mills (each with an installed power of 3,360 kW).

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The reground copper-gold rougher concentrate will then be upgraded in a cleaner flotation circuit comprised of two banks of cells with three stages of copper cleaner flotation producing a copper-gold concentrate with a target grade of 25% copper. The first cleaning stage will consist of two banks of four 100 m3 tank cells, the second stage of two banks of five 25 m3 tank cells and in the third stage two banks of three 25 m3 tank cells will be used. First cleaner flotation tailing will be further floated in two banks of three cleaner scavenger flotation tank cells each with a 50 m3 capacity. The concentrate product from the cleaner scavenger flotation will be sent to the first cleaner cells while the tailing will report to the gold leaching circuit if the leaching circuit is in operation. The tailing from the second and third cleaner flotation stages will be returned to the first cell in the prior cleaner flotation circuit.

The upgraded copper-gold concentrate will be thickened in a 21 m diameter high rate thickener. The thickener underflow will be directed to the three 160 m2 copper-gold concentrate pressure filters to further reduce water content to 9% moisture. The copper-gold concentrate will be stockpiled on site and then transported by trucks to a seaport at Stewart, B.C. where the concentrate will be stored and loaded into ships for ocean transport to overseas smelters.

Located at the tailing of each copper rougher scavenger bank is a further two lines of four 600 m3 tank cells which are used for pyrite flotation. These cells will produce a pyrite concentrate which is gold bearing. The final pyrite flotation tailing will be sent to the TMF. Pyrite concentrate will be reground in two tower mills to a particle size of 80% passing 20 µm and floated to remove liberated copper minerals. The copper flotation tailing along with bulk copper and gold cleaner tailing from Mitchell and Sulphurets mill feed are fed to CIL leaching for gold recovery.

Pyrite concentrate and first cleaner scavenger tailing from Iron Cap/Kerr mill feed is routed to the lined CIL residue storage pond in the TMF.

Pyrite concentrate and first cleaner scavenger tailing for Mitchell and Sulphurets report to the gold leach circuit.

Gold Recovery from the Gold-Bearing Pyrite Products

Both the reground gold-bearing pyrite concentrate, and the first cleaner scavenger tailing will each be thickened in dedicated 43 m diameter thickeners to a solid density of 65%. The underflow of each thickener will be pumped to its own dedicated cyanide leaching line. Each line will consist of two pre-treatment tanks and six conventional CIL tanks.

Loaded carbon leaving the CIL circuit will be transferred to an elution circuit for gold recovery followed by the reactivation of the barren carbon in an electrically heated rotary kiln. Reactivated carbon will be circulated back to the CIL circuit. Tailing from the CIL lines will pass over a safety screen system prior to being processed for cyanide recovery.

Pregnant solution from the elution system will go through the electrowinning cell where a precious metal sludge will be produced. The sludge will be filtered, dried and smelted in an induction furnace to produce gold and silver doré.

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Treatment of Leach Residues

The combined residue from the CIL circuits will be pumped to a single two-stage conventional CCD washing circuit. The CCD circuit will consist of two 50 m diameter high-rate thickeners in series. Thickener overflow from the first stage washing will be pumped to the cyanide recovery system. There, the barren leach solution containing copper cyanide complexes from cyanidation is treated with sulphuric acid and sodium sulphide to precipitate copper as copper sulphide, which is thickened and recovered. This precipitate will be blended with the copper-gold concentrate for sale. The acidification converts part of the cyanide to hydrogen cyanide gas allowing it to be recovered in a wet scrubber system as a high strength cyanide solution for recycle back to the gold leach circuit. The underflow (washed residues) of the second thickener will be sent to a SO2/air cyanide destruction circuit prior to being pumped to the lined CIL residue storage pond in the TMF.

Tailing Management

The flotation tailing and the treated CIL residue will separately flow to the TMF located southeast of the main process plant. Flotation tailing and CIL residue will be stored in separate areas within the TMF. Based on the PEA production plan and the metallurgical test results, the sulphide residue/tailing reporting to centre pond is estimated to be 454 Mt (or approximately 19.4% of the total tailing) and the rougher scavenger tailing to the north and south cells comprise the balance of tailing volume placed in the TMF.

CIL residue (pyrite concentrate in the case of Iron Cap/Kerr) will be placed in the lined CIL residue storage pond. The residue will be covered with the supernatant to prevent sulphide oxidation. The residue will be eventually covered by the flotation tailing, from which most sulphides have been removed. Supernatant from the CIL residue pond will be reclaimed by pumping to the CIL circuit for reuse. Excess water will be combined with the water from the flotation tailing storage cells within the TMF and reused in the flotation circuit as process water.

There will be two flotation tailing pipelines directing the flotation tailing to the TMF. The flotation tailing from one of the tailing pipelines will be classified to produce coarse tailing sands by two stages of cyclone classification. The coarse fraction will be used to construct tailing dam embankments and the fines will directly report to the TMF cells. The second tailing pipeline will report directly to the TMF. The supernatant from the tailing impoundment area will be reclaimed to the process water tank by two stages of pumping. The water will be used as process water for the flotation circuits.

Reagents

All the reagents will be prepared in a dedicated reagent preparation and storage facility within a containment area. Liquid reagents will be added in the undiluted form via metering pumps. Solid reagents will be prepared into adequate strength solutions in dedicated mixing tanks and stored in holding tanks to be added to the processes via metering pumps.

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Water and Pressurized Air Supply

Three separate water supply systems will be provided to support the operation:

| ● | A fresh water system.

| ● | A process water system for grinding/flotation circuits.

| ● | A process water system for CIL/gold recovery circuits.

Plant air service systems will supply blower air to flotation, moderate-pressure air to leach, cyanide recovery and cyanide destruction, and high-pressure air to filtration and general plant and instrumentation services.

Assay and Metallurgical Laboratory

The assay laboratory will be equipped with necessary analytical instruments to provide routine assays for the mine, process, and environmental departments.

The metallurgical laboratory, with laboratory equipment and instruments, will undertake all necessary test work to monitor metallurgical performance and to improve the plant production and metallurgical results.

Process Control and Instrumentation

The plant control system will consist of a DCS with PC-based OIS located in control rooms at the process facilities. Process control will be enhanced with the installation of an automatic sampling system. The system will collect samples from various lines for on-line analysis and the daily metallurgical balance.

For the protection of operating staff, cyanide and sulphur dioxide monitoring/alarm systems will be installed at site where needed. In addition, CCTV support will be provided at various locations at the crushing and plant facilities to ensure comprehensive site monitoring.

Process Flow Diagram

Figure 24.30 below illustrates the overall process block flow diagram. The major design consideration in the process plant equipment sizing and layout is the use of the largest equipment sizing available in order to minimize pumping and piping requirements, process building footprint and capital costs. Figure 24.31 and Figure 24.32 illustrate the main process building layout.

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Figure 24.30 Overall Process Block Flow Diagram

Source: Wood (2020)

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Figure 24.31 Main Process Plant Layout

Source: Wood (2020)

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Figure 24.32 Leaching Plant Layout

Source: Wood (2020)

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| 24.17.3 | Mill Design Criteria

The process plant design criteria for the 2020 PEA flowsheet are based on the following sources of information and analysis:

| ● | Metallurgical test work analysis and recovery assessment by Tetra Tech.

| ● | Crushing and grinding calculations by Wood (Wood, 2020)

Table 24.21 summarizes the main design criteria established for the 2020 PEA. The criteria are for a processing plant of 170,000 t/d capacity with a plant availability of 94%.

Table 24.21 Major Process Design Criteria

Parameter | Units | Value
Plant Feed Rate
Process Plant Availability | % | 94
Annual Processing Rate | Mt/a, dry | 62
Daily Processing Rate | t/d, dry | 170,000
Hourly Processing Rate, Nominal | t/h, dry | 7,535
Primary Grind Size, P80 | µm | 125 – 150
Concentrate Regrind Size, P80 | µm | 20
Process Methods |  | Flotation+Cyanidation
Leach Circuit | - | CIL
Elution Circuit | - | Pressure Zadra
Carbon Regeneration | - | Electric Kiln
Residue Management | - | CCD
Cyanide Recovery Method | - | AVR
Cyanide Destruction Method | - | SO2 + Air
Head Grades
Head Copper Grade, Average | % Cu | 0.37
Head Gold Grade, Average | g/t Au | 0.53
Head Silver Grade, Average | g/t Ag | 2.5

Major Equipment Sizing Criteria

Metallurgical calculation output for comminution circuit equipment sizing and installed power is listed in Table 24.22, Table 24.23 and Table 24.24.

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Table 24.22 Comminution – Major Equipment Sizing

Equipment | Unit | Value
Primary Crushing – Gyratory Crusher (OP)
Number of Units | - | 2
Installed Motor per Unit | kW | 600
Recommended Gyratory Crusher Size | mm | 1,524 x 2,260
Secondary Crushing – Cone Crusher
Number of Units | - | 4
Installed Motor per Unit | kW | 933
Tertiary Crushing - HPGR
Number of Crushers | - | 4
Installed Power (per unit) | kW | 7,400
Recommended HPGR Size | mm Ø x mm W | 2,600 x 1,750
Ball Mill
Number of Ball Mills – Mitchell/Sulphurets |  | 2
Dimensions | m (ø x length EGL) | 8.2 x 12.8
| ft. (ø x length EGL) | 27 x 42
Installed motor (per unit) | MW | 19,900
Number of Ball Mills – Iron Cap/Kerr | - | 2
Dimensions | m (ø x length EGL) | 8.5 x 13.4
ft. (ø x length EGL) | 28 x 44
Installed motor (per unit) | MW | 23,100

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Table 24.23 Copper-Gold Flotation – Major Equipment List

Equipment | Unit | Value
Rougher/Scavenger Flotation
Number of Trains | - | 2
Number of Cells/Train | - | 4
Cell Type |  | Tank Cell
Cell Volume | m3 | 600
Installed Motor | kW | 450
Bulk Concentrate Regrind
Mill Type | - | Vertimill
Number of Mills | - | 2
Installed Power (per unit) | kW | 3,360
1st Cleaner Flotation
Number of Trains | - | 2
Number of Cells/Train | - | 4
Cell Type | - | Tank Cell
Cell Size | m3 | 100
Installed Motor | kW | 132
1st Cleaner Scavenger Flotation
Number of Trains | - | 2
Number of Cells/Train | - | 3
Cell Type | - | Tank Cell
Cell Size | m3 | 50
Installed Motor | kW | 65
2nd Cleaner Flotation
Number of Trains | - | 2
Number of Cells/Train | - | 5
Cell Volume | m3 | 25
Installed Motor | kW | 55
3rd Cleaner Flotation
Number of Trains | - | 2
Number of Cells/Train | - | 3
Cell Volume | m3 | 25
Installed Motor | kW | 55
Copper Concentrate Dewatering
Thickener Type | - | High Rate
Thickener Diameter | m | 21
Filter Type | - | Vertical
Number of Filters | - | 3
Filter Size | m2 | 160

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Table 24.24 Pyrite Flotation & Cyanidation – Major Equipment List

Equipment | Unit | Value
Pyrite Flotation
Number of Trains | - | 2
Number of Cells/Train | - | 4
Cell Volume | m3 | 600
Installed Motor | kW | 450
Pyrite Concentrate Regrind
Mill Type | - | Vertimill
Number of Mills | - | 2
Installed Power (per unit) | kW | 3,360
Pre-Leach Thickeners
Number of Thickeners | - | 2
Thickener Size | m, diameter | 43
Pre-Aeration Tanks
Number of Trains | - | 2
Number of Tanks/Train | - | 2
Tank Volume | m3 | 573
Cyanidation Tanks
Number of Trains | - | 2
Number of Tanks/Train | - | 6
Tank Volume | m3 | 2,474
Cyanide Recovery Circuit
Number of Thickeners | - | 2
Thickener size | m, diameter | 50

Interpretations and Conclusions

The increased throughput envisioned in this 2020 PEA can be achieved using larger proven units available in the market for each unit operation without significantly redesigning the facilities proposed in the approved EA.

| 24.18 | Project Infrastructure

| 24.18.1 | On-Site Infrastructure

The water and waste management structure designs in the 2020 PEA are based on those described in the approved EA with conceptual-level adjustments to the 2020 PEA mine plan layouts, staging plan, and capacities required for the 2020 PEA mine production schedule.

Due to alignment changes and simplification of designs for the 2020 PEA (KCB 2020c), both initial and ultimate water management tunneling requirements in the 2020 PEA on the mine site are less than in prior studies. The approximate length at start up of these tunnels in the 2020 PEA is 18 km, whereas prior studies showed 25 km. Through the 2020 PEA life of mine in Mitchell Valley, the ultimate water management tunnels on the mine site total 20 km, whereas prior studies included 51 km.

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Rock Storage Facility

The 2020 PEA rock storage requirements are significantly less than in the approved EA due to waste reductions during open pit phases of mining (1.3 Bt versus 3.0 Bt). The 2020 PEA waste management plan consolidates all mine waste into a single facility, including minor waste volume from initial underground development. As a result, the McTagg RSF and the Sulphurets Pit Backfill RSF are eliminated in the 2020 PEA, with a reduction to the Mitchell RSF footprint. The Mitchell RSF geotechnical and water management design criteria, operating concepts, and closure design concepts from prior studies are retained for the 2020 PEA. The selenium drain configuration for the Mitchell RSF is reduced in the 2020 PEA, as a result of the elimination of its extension into the McTagg RSF.

Water Management

The overall 2020 PEA operational water management and water treatment designs are similar to those in the approved EA, with adjustments in some areas based on the revised mine plan and recently collected field data. For closure, the 2020 PEA water management plan remains the same as in the approved EA, except for the elimination of McTagg RSF and the associated closure channels.

Due to elimination of Mitchell Block Cave and inclusion of a smaller Mitchell Open Pit, the MDT design basis is reduced from a 1,000 year to 200 year flow. This reduction is justified by the lower sensitivity of an open pit to flood flows compared to a block cave. As a result, the twin MDT tunnels from prior studies are reduced in the 2020 PEA to a single MDT tunnel. Additionally, the MDT is re-aligned to divert water to Sulphurets Valley more directly, in order to avoid mineralized areas and to reduce tunnel length. The normally wetted depth of the revised MDT is concrete-lined to improve reliability, reduce erosion and to reduce potential for loadings from mineralized wall rock. Additionally, the revised alignment improves the reliability and constructability of the structure with reduced risk. The revised outlet location; however, precludes MDT hydropower development, due to the challenging terrain for powerhouse and power transmission line development.

Elimination of the block cave and a smaller Mitchell Open Pit allow shortening and realignment of the NPWDA that carries contact water from above Mitchell Pit. Glacier recession and the elimination of the large inlet setback, that was previously required to avoid block cave subsidence cracking, allowed replacing the two additional stages of the surface inlets shown in prior studies with a single stage in the 2020 PEA.

In the 2020 PEA configuration, McTagg Valley catchments are diverted around Mitchell RSF and the WSF by an inlet at the same location as in the approved EA. MTDT’s design was changed to a single bore design in the 2020 PEA that is fully concrete-lined. This updated design reduces tunneling volume, improves reliability, and matches better the rock types along the route. The redesigned MTDT includes provision for hydroelectric development from start-up, instead of only in the later stages as shown in prior studies, thus replacing what was eliminated at MDT. The redesigned MTDT hydroelectric facility has a larger catchment than in prior studies. The revised tunnel route and outlet portal location improves winter constructability and has direct access to the muck pads via a short construction access tunnel bypassing an avalanche area. The MTDT construction access tunnel also shortens the length of the penstock to the McTagg Power Plant.

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The WSD crest elevation is kept at the same elevation for the 2020 PEA. With the reduced tonnage of waste rock, WSD storage and WSD crest elevation could potentially be reduced, although these optimizations have not been adopted for the 2020 PEA.

To improve constructability, the asphalt-covered concrete plinth in the lower section of the dam foundation is enlarged, to allow inclusion of a concrete grouting gallery which allows remedial grouting, if needed. During construction, this gallery also decouples grouting from dam raising, which reduces risks to the construction schedule caused by grouting delays.

The 2020 PEA includes improvements in definition of borrow sources for the WSD. Within the Mitchell Open Pit waste, a large source of monzonite rockfill for WSD construction is identified on the north valley wall. Geotechnical and geochemical testing confirmed that Mitchell P8 monzonites are equivalent to the Sulphurets Quarry P8 Monzonites, which were identified as the WSD rockfill source in prior studies (KCB, 2019b). The Mitchell P8 Monzonites have a shorter haul distance to the WSD, and access can be developed earlier in the construction sequence than for the Sulphurets Quarry. Developing a construction rock source in Mitchell valley benefits the development plan by eliminating the need for local quarries around the WSD spillway (and above the WSD along the access road) in the 2020 PEA.

The waste management plan to consolidate all mine waste rock into a single facility resulted in elimination of the Sulphurets Pit Backfill RSF. The associated backfill cover and Sulphurets RSF selenium collection system are removed as unnecessary. Kerr waste will be hauled to the Mitchell RSF to utilize the selenium collection system there. The use of truck haul eliminates need for crushing Kerr waste to place on a conveyor. The elimination of crushing will reduce exposed surface area of the waste material and may reduce ML/ARD generation from this waste product.

HDS Water Treatment Plant

The 2020 PEA catchment area has reduced from the EA due to a reassessment of catchment areas, a change in pit areas, and reducing the total waste rock tonnage. This reduction in catchment area from 37.3 km2 to 29.1 km2 has resulted in a 22% reduction in the amount of run off from the catchment area that has to be treated prior to discharge to the environment (KCB, 2020a). The direct effect of this reduction is that the design rate for the HDS water treatment plant (freshet maximum), that processes the mine water run-off, is now reduced from 7.5 m3/s to 5.9 m3/s, resulting in a reduction in the number of HDW water treatment trains required from seven for the EA to six for the 2020 PEA. Figure 24.33 shows the number of trains required and constructed for the 5.9 m3/s mine water run-off flow rate.

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Figure 24.33 Water Treatment Trains Required and Constructed

Source: Wood (2020)

Tailings Management

TMF configurations for the 2020 PEA are generally consistent with the TMF designs and the staged Cell configuration presented in the approved EA, except for slight modifications in starter dam heights and sizing of the CIL cell. The 2020 PEA has a higher peak throughput with an annual average of 170,000 t/d during early operations, so starter dams for the North Dam, Splitter Dam, and Saddle Dam need to be 10 m higher than in prior studies, therefore increasing the TMF initial capital costs. The relatively higher 2020 PEA throughput reduces TMF Stage 1 North Cell operational duration from 25 to 19 years. TMF Stage 2 requires construction of the Southeast Dam and operation of the South Cell. Due to reduced throughput during the underground mine production-only TMF Stage 2, there is no change in starter dam height for the Southeast Dam.

Final crest elevations for all dams and the total design capacity of 2.3 Bt are similar to prior studies and total storage is within the contingency storage provided. The 2020 PEA CIL Residue Cell is enlarged to account for CIL tailings storage of approximately 20% of total tailings produced from the process plant. This is due to the increased sulphide content of Deep Kerr mineralization and relatively higher portion of total mill feed comprised of this material source. The lined CIL Residue Cell is enlarged to accommodate the additional storage by relocating the Saddle Dam to the south.

A new glacial till deposit adjacent to the TMF has been identified, with a potential to supply dam core borrow material for the life of mine. As a result, the TMF contingencies for potential bentonite amendment to meet a low permeability specification have been removed.

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MTT Rail System

MTT rail system will transport crushed mill feed, freight, fuel and personnel between the Mitchell OPC and Treaty OPC.

Crushed mill feed will be conveyed to one of two 15,000 t capacity underground bins. Loading chutes under the bin will feed mineralized material into train wagons for transport to Treaty OPC where the wagons will bottom dump the crushed mill feed into a 15,000 t capacity ore bin. Apron feeders will then reclaim the mill feed to a belt conveyor which will report to the 60,000 t live capacity COS.

Ten mill feed trains will be needed to deliver an average of 170,000 t/d of mill feed to the process plant. Each mill feed train will be made-up of a 140 t electric locomotive and 16 off 42 m3 bottom-dump mill feed wagons.

Specially configured personnel and freight trains will transport freight, fuel, and personnel through the MTT. Staging areas at each end of the MTT for marshalling and loading/unloading trains at each end will separate these activities from those for mill feed transport. Personnel, freight, and fuel handling will be scheduled during the day shift operations. A maintenance shop and siding for rolling stock will be located in the Treaty staging area.

Train operations, with the exception of train loading, will be controlled by an automated controls and scheduling system. The main control room will be located in the Treaty staging area. Locomotives will be unmanned and not require engine drivers.

To meet the daily production requirements and provide sufficient surge capacity during periods of peak demand, the rolling stock fleet will be as follows:

| ● | 12 off 140 t locomotives

| ● | 192 off 42 m3 bottom-discharge mill feed wagons

| ● | 12 off 50 t multi-purpose freight wagons

| ● | 6 off ISO fuel transport tanks wagons

| ● | 6 off ISO lime slurry transport wagons

| ● | 6 off 60-seat personnel carriages

| ● | 2 off battery shunting wagons

| ● | 1 off rail maintenance vehicle

| ● | 6 off heavy-duty drop frame tri-axle rubber tired ISO-tank chassis.

The MTT rail system can increase short term upside capacity when required for production recovery by adjusting track time allocation between mill feed, freight and personnel transport. Longer term recovery requirements will need to be addressed in future studies with considering Treaty stockpiling, detailed track logistics studies, and rolling stock availability.

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Facilities, Buildings, and Services

Facilities, buildings and services in the mine site and Treaty OPC will include the following:

| ● | site access roads

| ● | fuel receiving, storage, and dispensing

| ● | sewage treatment

| ● | medical buildings and ambulance stations

| ● | administration offices

| ● | consulting engineers’ and contractors’ offices and storage/laydown areas

| ● | warehouses, cold storage, and maintenance buildings

| ● | truck maintenance and wash bay

| ● | batch plants

| ● | accommodation facilities including receptions, lunch rooms and cafeterias, mine dry and wash cars, recreation facilities, kitchen, and parking areas

| ● | potable water treatment

| ● | potable, process, fire and fresh water systems

| ● | container storage

| ● | laydown and equipment and materials storage areas

| ● | landfill

| ● | waste management storage and handling and Incinerators

| ● | security fencing and gates

| ● | helipads

| ● | explosives magazines and AN prill storage areas

| ● | explosives manufacturing plant

| ● | communication systems

| ● | auxiliary truck and support vehicles

| ● | electrical power supply and distribution.

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| 24.18.2 | Off-site Infrastructure

| 24.18.2.1 | Electrical Power Supply

Electrical power will be supplied to the KSM Site from the BC Hydro NTL 287 kV transmission line that runs from the Skeena Substation near Terrace, BC, to a substation near Bob Quinn Lake, further north of the KSM Site. KSM will connect to the NTL transmission line at the new Treaty Creek Terminal (Switching Station) located adjacent to Highway 37, approximately 18 km south of Bell II. This station, designated TCT by BC Hydro, is currently at the preliminary design stage. A new 30 km long, 287 kV transmission line tap will connect this substation to the KSM Main Substation No. FLT1 at the Treaty OPC.

BC Hydro has completed an updated System Impact Study (SIS) that confirms a Contract demand of 245 MW for KSM. BC Hydro has now started a Facilities Study that will define the preferred interconnection option. Upon completion of the Facilities Study , the BC Hydro substation project can proceed to the construction phase upon signing of a Facilities Agreement.

The cost of electric power at the site main substations, including transformer and line losses, is Cdn$0.0614 per kilowatt hour for the 170,000 tpd operation with HPGR comminution which qualifies for reduced power rates under BC Hydro’s Power Smart program. These costs are based on BC Hydro Rate Schedule 1823, Transmission Service, Stepped Rate, which has two tiers of energy pricing to encourage conservation.

The KSM 287 KV transmission line tap will terminate at the compact gas insulated (GIS) main substation designated FLT1 at the Treaty OPC. This substation steps the voltage down to 25 kV to supply the various process plant and other area feeders and 138 kV to supply the cable feed thought the MTT to the Mitchell area. The substation has an installed capacity of 500 MVA based on two stages of transformer fan cooling operating (under emergency conditions) and provides full redundancy of electric power supply and flexibility for maintenance.

The 25 km long 138 kV cables through the MTT will terminate at GIS substation FLT2 located at the Mitchell OPC. This station steps the voltage down to 25 kV for local distribution and 69 kV for distribution to more remote facilities. The substation has an installed capacity of 270 MVA with two stages of transformer fan cooling in operation (under emergency conditions) and provides full redundancy of electric power supply and flexibility for maintenance.

The 2020 PEA includes 3 energy recovery systems. These systems recover energy from the McTagg diversion tunnel, the discharge from the WSF to the WTP, and the North Tailings Line. The plants generate an average of 60,122 megawatt hours per annum over the first 10 years of operation with an annual value of Cdn$5,200,000, based on the BC Hydro Tier 2 rate for energy. Subsequent to the first decade, the generation varies somewhat, but the McTagg and WTP energy recovery systems will continue to operate after mine closure.

The electrical infrastructure also includes a large number of construction phase diesel power plants and associated distribution to supply power to all aspects of the construction. The installed capacity, including units required for redundancy, is 38,560 KW at prime rating.

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| 24.18.2.2 | Access Roads

The access roads to the KSM site will be as follows:

| ● | Eskay Creek Access Road

| ● | CCAR

| ● | TCAR

| ● | Lower NTAR

| ● | Upper NTAR

| ● | Cut-off Ditch Access Road

| 24.18.2.3 | Logistics

Inbound equipment and materials will be transported either by barge to Stewart, BC, or by rail to Terrace, BC, where these loads will be consolidated at local marshalling yards or staging areas for onward transport by truck to KSM.

Copper concentrate will be transported by truck to a deep water port in Stewart, BC, where it will be held in storage until loaded onto oceangoing vessels.

| 24.18.3 | Iron Cap Project Infrastructure

Road and Logistics

The main access for personnel and materials to the Iron Cap operations is through the MTT tunnels originating in the Treaty Valley and connecting to the Mitchell OPC. Access is then via the Iron Cap access tunnel collared nearby. All personnel will stay in a mining camp located in the Mitchell Valley area. Materials required for the operation of the mine will be delivered via the MTT tunnels to warehouses in the Mitchell Valley.

Underground Crushers and Conveyors

The Iron Cap block cave will produce 90kt/d at steady state production. The mill feed will be transported by LHD directly from the drawpoints to one of three ore passes located along each extraction drift. Each pass is fitted with a stationary rock breaker and the sized material gravity feeds to the haulage level where continuous loading chutes transfer the uncrushed material into train cars. Electric locomotives then transport the cars to one of two crusher unloading stations where the material is bottom dumped into the crushers. Two 60″ x 89″ gyratory crushers will be installed to crush the mill feed to 152 mm size.

The crushed material will be transported by a 2.4 km long series of conveyors feeding two ore bins that in turn feed the MTT trains for transfer to the process plant. The conveyors will vary in width from 1,000 mm (48″) to 14,00 mm (54″) depending on location and requirement. The belts do not exceed a 15% grade.

Conveyor structures will be back mounted to provide ease of cleaning under the belts. Conveyor drives and tail and head pulleys will be sill mounted for ease of installation and maintenance.

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Underground Mine Service Water System

The mine will require 83 m3/h of process water for the bolters and for the development and production drills. In addition, it is estimated that 50 m3/h of water will be required for the conveyor fire suppression systems. This water will be supplied through the Iron Cap access tunnel from pumps in the Mitchell Valley. The water will be delivered to the working face through steel pipes.

Underground Electrical Reticulation System

The Iron Cap mine will require approximately 19 MW of average electrical power at peak operation. The main contributors to this demand are the crushers, conveyors, ventilation fans, and drill jumbos. The annual power consumption is estimated to be 166,000 MW hours.

The main power will be supplied to the underground from Mitchell area substation FLT2 through two redundant 25 kV cables hung from the back of the Iron Cap access tunnel and MTT train tunnel to create a ring main style system. Each of the main levels will have a 25 kV feed except the haulage level will have separate feeders for the west and east train loops and the extraction level will also have two separate feeders. The 25 kV distribution voltage will be stepped down to the required utilization voltage by skid mounted dry-type unit substation transformers. Equipment that draws larger loads (e.g., ventilation fans, conveyors and crushers) will be fed from local permanent transformers.

Waste Storage Facilities

Approximately 4 Mt of development rock (mineralized and non-mineralized) will be generated during the four-year pre-production period and hauled by truck to the Mitchell OPC. The 2020 PEA mine plan assumes the NPAG portion of this material will be used in construction of pads and the other material will either be hauled to the Mitchell RSF or fed into the Mitchell mill feed. After Year 5, once one of the crushers and the conveyor are installed, all development material will be crushed and conveyed to the MTT bins and hauled to the mill by the trains. Most of this material will be mineralized rock from within the footprint.

Water Management

The underground mine dewatering system (refer to 24.16.3 Mine Dewatering) is discharged into the North Pit Wall Dewatering Adit (NPWDA) via graded drifts that allow for water to drain by gravity. From there, it will flow to the WSF.

Surface Infrastructure

The 2020 PEA design assumes the surface infrastructure required to support the Iron Cap operations will be integrated with the overall site infrastructure. This will include major equipment maintenance and fuel provision.

Surface infrastructure required to support operations include:

| ● | equipment shop for major repairs

| ● | explosives magazine

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| ● | warehouse

| ● | concrete plant

| ● | water pump stations

| ● | general mine offices

| ● | control room and training center

| ● | first aid station and mine rescue room

| ● | construction and operations camp

| ● | mining contractor facilities

| ● | change house

| ● | instrumentation shop

Underground Infrastructure

Underground facilities at Iron Cap include:

| ● | equipment shop

| ● | warehouse

| ● | fuel bays

| ● | explosives magazines

| ● | sampling room

| ● | refuge stations (fixed and portable)

| ● | lunchrooms

| ● | offices, map and meeting room, training room

| ● | first-aid station

| ● | comfort stations

| ● | water tanks

| ● | fire water tanks and fire suppression systems, primarily for crushers, conveyor transfers and shops.

| 24.18.4 | Deep Kerr Project Infrastructure

Road and Logistics

The main access for personnel and materials to the mining complex in Mitchell Valley is through the MTT tunnels connecting the Treaty Valley to the Mitchell OPC. From the Mitchell OPC, a series of roads developed for the overall KSM site will lead to the Deep Kerr surface complex area just outside of the Deep Kerr portals. All personnel will stay in the mining camp located in the Mitchell Valley area. Materials required by the Deep Kerr mine will be delivered via the MTT tunnels to the warehouses in the Mitchell Valley.

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Deep Kerr mine is accessed through a series of portals and ventilation adits along the northwestern base of Kerr Mountain. Mine access is primarily through two of the intake drifts, one of which directly connects to the footprint area as a main ventilation intake drive, and the other eventually parallels the UG conveyor as a service drive that leads to the crusher area. These two primary drifts will connect via ramps to the lower lift (Lift 2) to provide access to the lower elevation.

Underground Crushers and Conveyors

The Deep Kerr block cave will produce 80 kt/d at steady state production. The mined material will be transferred from production drawpoints to ore passes fitted with stationary rock breakers to size the material before it gravity feeds to the haulage level below. At the haulage level, 55 t trucks will haul the material to the crushers.

The mine will produce from two cave lifts over LOM, the 625 level (Lift 1) and the 130 level (Lift 2). Each lift will be serviced by two, 60 inch x 89 inch gyratory crushers on the haulage level. These crushers are sufficient to support the peak production rate of 80 kt/d. Each crusher will have three dumping points for rear dump articulated underground haul trucks. The crushers will discharge to incline conveyor systems which connect to a series of conveyors to transfer the material to the MTT transfer bin at the portal.

There will be a total of 10.2 km of underground conveyor belts with widths ranging from 1,000 mm (48″) to 1,400 mm (54″). Conveyor structures will be back mounted to provide ease of cleaning under the belts. Conveyor drives and tail and head pulleys will be sill mounted for ease of installation and maintenance. Most conveyor inclinations are at 15% or less and none exceed 17%.

Underground Mine Service Water System

It is estimated that the mine will require 91 m3/hr of process water for the bolters and the development and production drills. In addition, it is estimated that 50 m3/hr will be required by the conveyor fire suppression systems. Mine service water is sourced from the surface complex near the Deep Kerr portal area and piped underground through a series of pipes to the main areas for each lift. Smaller pipelines will branch off the main feeder pipe on each level to support excavation of mine development headings, longhole drilling, shop activities, etc.

Underground Electrical Reticulation System

The Deep Kerr mine will require an average of 39.1 MW electrical power at peak operation. The main contributors to this demand are the crushers, conveyors, ventilation fans, and drill jumbos. This load is larger than typical as it includes requirements for significant pumping plus the long conveyor to the MTT portal. The annual power consumption is estimated to be 342,770 megawatt hours.

The main power will be supplied to the underground from the Mitchell area substation FLT 2 via a 69 kV distribution line to a modular 69 to 25 kV surface substation near the mine. A 25 kV ring main system will feed the mine via Teck cables from this substation. Each of the main levels will have a 25 kV line which will be stepped down to the required voltage by skid mounted dry-type transformers. Equipment that draws larger loads (e.g., ventilation fans, conveyors and crushers) will be equipped with a permanent transformer.

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Waste Storage Facilities

Approximately 5 Mt of waste rock (mineralized and non-mineralized) will be generated and brought to surface during the initial 5-year pre-production period. NPAG portion of this material will be hauled by truck to the Mitchell RSF. The mineralized material will be hauled to the Mitchell OPC and fed into the mill feed.

After Year 11, once one of the crushers and the conveyor to the MTT are installed, all development material will be crushed and conveyed to the MTT bins and hauled to the mill by the trains. Most of this material will be mineralized rock from within the footprint.

Water Management

The underground mine dewatering system (refer to Section 24.16.4Mine Dewatering) is discharged to the surface complex established outside of the Deep Kerr portal areas. From there the water will gravity flow via pipeline to the WTP platform and then be pumped to the WSF in the Mitchell Valley.

Underground Electrical Reticulation System for Dewatering

The mine dewatering system at Deep Kerr requires an average of 4 MWh with a maximum of 30 MWh during a peak storm event, which is greater than the power requirements of the mine under normal conditions. The strategy during a peak storm event will be to shut down or reduce operations in the underground mine along with other site facilities when the high-powered pumps are required. This will allow power to be diverted from normal operations to power the pumps.

The main power will be supplied to the underground from the Mitchell area substation FLT2 through a 25 kV cable hung from the back of the access ramp and conveyor tunnel to create a ring main style system. Each of the main levels will have a 25 kV line which will be stepped down to the required voltage by skid mounted dry-type transformers. Equipment that draws larger loads (e.g., ventilation fans, conveyors and crushers) will be equipped with a permanent transformer.

Surface Infrastructure

The 2020 PEA design assumes the surface infrastructure required to support the Deep Kerr operations will be integrated with the overall site infrastructure.

Surface infrastructure required to support operations will include:

| ● | explosives magazine

| ● | warehouse

| ● | water pump stations

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| ● | concrete plant

| ● | general mine offices

| ● | control room and training center

| ● | first-aid station and mine rescue room

| ● | construction and operations camp

| ● | mining contractor facilities

| ● | change house

| ● | instrumentation shop.

Underground Infrastructure

Underground facilities at Deep Kerr include the following on each lift:

| ● | extraction level equipment shop

| ● | haulage level equipment shop

| ● | warehouse

| ● | fuel bay

| ● | explosives magazine

| ● | concrete / shotcrete slickline and loading system (for Lift 2 only)

| ● | sampling room

| ● | refuges

| ● | lunchrooms

| ● | offices, map and meeting room, training room

| ● | first-aid station

| ● | comfort stations

| ● | service water tanks

| ● | fire water tanks and fire suppression systems, primarily for crushers, conveyor transfers and shops.

| 24.19 | Market Studies and Contracts

Copper Concentrate

Seabridge engaged NSA to provide opinion reports on marketing inputs for the 2016 PEA and review the 2019 copper-gold concentrate market and related concentrate treatment charges, excluding offsite transportation costs. The information and options in this section come from the opinion reports. No smelter contracts are currently in place or being negotiated. All currency amounts used in this section are in US dollars, unless otherwise specified.

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Marketability

When considering the marketability of copper concentrates, quality and quantity are determining factors. There are considerable variations in the quality of concentrates and the requirements of various smelters do vary; such variations relate to the technical abilities of the smelter and its overall concentrate feed and blend. Gold, silver and copper are key elements in determining concentrate salability, as well as iron, sulphur and other impurities.

Based on the impurity levels determined by metallurgical testwork, Tetra Tech expects that concentrates from the 2020 PEA mill production plan should be relatively clean. Depending on the market situation at the time of contract negotiations, penalties will likely be minimal, if at all applicable. Mill feeds from different deposits and spatial locations will be blended, therefore clean mill feeds from certain deposits should effectively mitigate penalizable elements in mill feeds from other deposits. As pointed out by NSA, some smelters, such as in Japan, South Korea, and Europe, are expected to have more interest in copper concentrates with high gold content.

Copper Concentrates Smelting Terms

NSA suggests that the annual benchmark terms realized at the time of the 2016 PEA are likely to be a guide to future levels. For 2016 PEA, a copper treatment charge of $100/dmt was assumed, with a $0.10/lb copper refining charge. The current copper-gold concentrate market review by NSA in Q4 2019 foresees that the treatment charges and refining charges are expected to be slightly lower than the 2016 projections. For 2020 PEA, Tetra Tech suggests keeping the smelting terms unchanged. Also, the payable terms used for the 2016 PEA should be kept unchanged as below.

Payable Metals

| Copper | Pay 96.5% with a minimum deduction of 1 unit (amount deducted has to equate to a minimum of 1% of the agreed concentrate copper assay).

| Silver | If over 30 g/dmt pay 90%.

| Gold | A scale is applicable with some variations of the following:

| ● | less than 1 g/dmt, no payment

| ● | 1 to 3 g/dmt, pay 90%

| ● | 3 to 5 g/dmt, pay 93%

| ● | 5 to 7 g/dmt, pay 95%

| ● | 7 to 10 g/dmt, pay 96.5%

| ● | 10 to 20 g/dmt, pay 97.0%

| ● | over 20 g/dmt pay 97.5%

| ● | over 30 g/dmt pay 97.75%.

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Refining Charges

| Copper | $0.10/lb payable copper

| Gold | $6.00 to $8.00/oz payable gold

| Silver | $0.50/oz payable silver

Treatment Charges

| Treatment Charge | $100.00/dmt CIF-FO main smelter port.

Price Participation

Not applicable at present.

Penalties

| Arsenic: | $2.50 to $3.00 per 0.1% over 0.1% up to 0.5% arsenic

| Antimony: | $3.00 to $4.00 per 0.1% over 0.1% antimony

| Lead: | $2.00 to $3.00 per 1% over 0.5% to 1.0% lead

| Zinc: | $2.00 to $3.00 per 1% over 2% to 3% zinc

| Mercury: | $2.00 per each 10 ppm over 10 ppm mercury

| Bismuth: | $3.00 to $5.00 per 0.01% over 0.03 to 0.05% bismuth

| Selenium: | $3.00 to $5.00 per 0.01% over 0.05% selenium

| Tellurium: | $4.00 to $5.00 per 0.01% over 0.02% to 0.03% tellurium

| Fluorine | $1.00 to $2.00 per 100 ppm over 300 ppm fluorine

| Chlorine | $1.00 to $3.00 per 100 ppm over 300 ppm chlorine.

Furthermore, penalties may also vary from smelter to smelter. It should be noted that for the elements where a percentage range is used, this relates to ranges of penalty thresholds that are negotiated.

Other Offsite Costs

Various indirect costs other than smelter charges include:

| ● | losses are assumed to be 0.1% or less.

| ● | marine transport insurance is assumed to be in the range of 0.10 to 0.15% of net invoice value of the concentrate.

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| ● | third party supervision, assaying and umpire costs are assumed to be approximately US$1.00/dmt.

| ● | the cost of marketing varies with concentrate tonnage, location, and number of smelters to be shipped. For the 2020 PEA, the estimated marketing cost is in the range of US$5.00 to US$10.00/dmt.

| ● | the copper concentrate transportation costs are based on the following assumptions by Tetra Tech:

| o | trucking: US$38.60/wmt

| o | port storage and handling: US$14.40/wmt

| o | ocean transport to Asian port: US$46.00/wmt.

Gold-Silver Doré

There are no gold and silver doré marketing studies conducted for the 2020 PEA. No smelter contracts are currently in place or being negotiated. The terms used for this study are based on the typical terms currently used in the markets. The general payment terms for gold and silver doré are assumed as below:

| ● | gold: pay 99.8% of content less a refining charge of US$1.00/accountable oz. Doré transportation is assumed to be US$1.00/oz

| ● | silver: pay 90.0% of content less a refining charge of US$1.00/accountable oz. Doré transportation is assumed to be US$ 1.00/oz

| ● | insurance and assay costing is assumed to be in the range of 0.10 to 0.15% of net invoice value of the doré.

| 24.20 | Environmental Studies, Permitting and Social or Community Impact

KSM successfully underwent a joint federal provincial harmonized environmental assessment with Federal, Provincial and the Nisga’a environmental assessment approvals granted in mid to late 2014. The governments conducted the EA cooperatively, in accordance with the principles of the Canada-BC Agreement on Environmental Assessment Cooperation (Cooperation Agreement 2004). The regulatory approvals for KSM were granted in 2014. Further details on the environmental/permitting regime may be found in Section 20.0.

The 2020 PEA which incorporates the inclusion of Deep Kerr as a block cave in addition to an open pit, involves the same environmental issues as the previous studies which successfully received its environmental approvals in 2014. Additionally, the 2020 PEA includes an expanded footprint for the Iron Cap block cave over what was shown in the previous studies. In 2020 PEA, KSM is situated in exactly the same geographical area, has the same environmental characteristics and will involve similar disturbances required to develop infrastructure and the mine as was previously assessed and approved.

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Based on this comparison, and upon completion and evaluation of additional technical studies required to support the inclusion of Deep Kerr block cave material into the mine plan and to identify and examine the net environmental benefit of proposed project changes, it is anticipated that the 2020 PEA would be approved to operate by the appropriate regulatory authorities. This regulatory approval would be forthcoming only after the appropriate engagement and information sharing has occurred with the Nisga’a Nation, the Tahltan Nation and First Nations whom have an interest in KSM

Potential design changes resulting in an overall net environmental benefit associated with the 2020 PEA include the following items:

| ● | a reduction in the size of proposed open pit mines

| ● | an increased reliance of underground block cave mines with a subsequent smaller surface disturbance relative to open pit mining

| ● | a reduction of the proposed tonnage of waste rock from 3 Bt to 1.3 Bt, exclusively placed in Mitchell valley

| ● | a reduction in the volume of water requiring treatment from the WSF

| ● | a reduction in the overall KSM footprint.

The closure plan costs outlined in the previous studies have been updated to reflect the proposed mine plan changes outlined in the 2020 PEA. Closure costs comprise physical closure (e.g. grading, covering and revegetation); access costs for post-closure activities; monitoring, inspection, and maintenance; and post-closure water treatment.

Seabridge has developed long term respectful relationships with the Nisga’a Nation, the Tahltan Nation and the three other First Nations groups whom are potentially influenced by the KSM development, over the past twelve years. These relationships, including the requirements contained within the Benefits Agreement signed with the Nisga’a Nation in June 2014, the Impact Benefit Agreement signed with the Tahltan in June 2019 and the Sustainability Agreement negotiated with the Gitanyow Wilps, also signed in June 2014 respectively, would remain in good standing and the agreements would continue to be adhered to by the Property operating company as the proposed mine plan outlined in the 2020 PEA is implemented.

| 24.21 | Capital and Operating Cost Estimates

The capital cost estimate is the product of engineering developed during the previous studies and the 2020 PEA. The unit rates used for the 2020 PEA capital and operating estimates (labour, power, fuel and other consumables) are based on recent price quotes and estimates. Well established internal benchmarks were used to check capital and operating costs to verify reasonableness of cost levels. These benchmarks are internal to each company.

The capital cost estimates were produced by the consulting firms named in Table 24.25 with the area of responsibility for each firm identified.

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Table 24.25 Capital Cost Estimate Responsibilities by Firm

Capital Cost Area | Firm Responsible
Open pit mine development and related infrastructure, equipment, wall depressurization, capitalized operating costs, MTT and associated infrastructure, indirects, and contingency | MMTS
Open pit dewatering and wall depressurization | BGC
Underground mine development and related equipment, dewatering, ventilation and services, infrastructure, indirects, and contingency | Golder
Water diversion tunnels (MDT, MTDT, MVDT, SSDA, NPWDA) | Quantities by KCB and costing by MMTS
Underground material handling for Iron Cap and Deep Kerr block caves | Wood
Process plant and related ancillary facilities | Wood
Electrical power supply and energy recovery | W. N. Brazier Assoc.
TMF, water management and water treatment | Quantities by KCB and costing by Wood
Construction indirects (except for mining), owner’s cost, and contingency (except for mining) | Wood

| 24.21.1 | Capital Cost Estimate

The 2020 PEA level estimate includes:

| ● | direct field costs of executing the mining, construction, installation and commissioning of all structures, utilities, materials, and equipment

| ● | indirect costs associated with spares, initial fills, freight/logistics, commissioning, EPCM, and vendor assistance

| ● | owner’s costs

| ● | contingency.

See Table 24.26 for a summary of the 2020 PEA capital cost estimate, summarized by activity.

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Table 24.26 2020 PEA Capital Cost Estimate Summary in

Area | Initial | Sustaining | Total
US$ M | US$ M | US$ M
Direct Costs |  |  |
Mine | 1,193 | 4,406 | 5,599
Process | 1,359 | 190 | 1,549
TMF | 488 | 509 | 998
Environmental | 14 | - | 14
On-site Infrastructure | 22 | 0 | 22
Off-site Infrastructure | 106 | 10 | 116
Power Supply/Energy Recovery | 159 | 214 | 373
Total Direct Capital | 3,342 | 5,330 | 8,672
Indirect cost | 802 | 648 | 1,450
Owners cost | 152 | - | 152
Contingency | 859 | 907 | 1,767
Total Estimated Total | 5,156 | 6,886 | 12,041

| Notes: | 1Sums may not add due to rounding, PST not included in table please see section 24.22 for details.

2All costs associated with closure cost are addressed separately in financial analyses outside of capital.

3Most of the sustaining on-site infrastructure is included in the mine category.

Basis of Estimate

This estimate falls under the AACE® Class 5 Estimate classification and its accuracy is expected to be within -30% to +50% of final 2020 PEA cost.

Initial capital cost is defined as all costs associated with development of the operation until first mill feed in Year 1. It includes mine, process, TMF, environmental, on-site/off site infrastructure and power supply.

Indirect Costs and Contingency

Allowances were made for indirect costs based on first principle calculations, in-house data, or 2020 PEA parameters. The overall factor applied to the direct capital costs to estimate indirect costs was 24%.

An overall contingency of 20% was applied for the 2020 PEA. Contingency is applied to the sum of direct, indirect, and owner’s costs.

Owner Capital Costs

KSM estimated owner’s costs based on the 2020 PEA development schedule. In total, owner’s costs are estimated at US$152 M.

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Sustaining Capital

Sustaining capital costs are capital costs incurred after the start of production. Sustaining capital includes items required either to replace worn-out or exhausted assets or to support planned mine development that does not increase production capacity. This includes projects that improve operational efficiency, safety, or decrease costs. Additionally, sustaining capital comprises costs associated with the development of new deposits that will maintain the throughput of the operation. The majority of this comprises opening up Sulphurets and Iron cap open pits, and development of the Iron Cap and Kerr block cave mines. This explains why mine site sustaining capital is higher than initial.

| 24.21.2 | Operating Cost Estimate

The operating cost estimate is the product of engineering developed during the previous studies and the 2020 PEA study. The unit rates used for the 2020 PEA estimates (labour, power, fuel and other consumables) are based on recent price quotes and estimates. Table 24.27 provides a listing of 2019 Q4 commodity prices used in the 2020 PEA estimates.

Table 24.27 2019 Q4 PEA Commodity Pricing

Commodity | Units | 2019 Q4 Values | Source
General Consumables |  |  |
Electrical Power Cost | Cdn$/kWhr | 0.0614 | W. N. Brazier Assoc.
Diesel Fuel for Off Road | Cdn$/litre | 0.983 | W. N. Brazier Assoc.
Propane | Cdn$/litre | 0.67 | W. N. Brazier Assoc.
Labour (mid-level total employee compensation) | Cdn$/yr | 126,000 | Seabridge

The 2020 PEA operating cost estimates were produced by consulting firms named in Table 24.28 with the area of responsibility for each firm identified.

Table 24.28 Operating Cost Estimate Responsibilities by Firm

Operating Cost Area | Firm Responsible
Open Pit Mining | MMTS
Underground mining | Golder
Underground material handling for Iron Cap and Deep Kerr block caves | Estimated by Wood and reviewed by Golder
Processing, G&A, site services, and tailings placement | Estimated by TT and reviewed by Wood
HDS water treatment | Quantities by KCB and costing by Wood
Se Water Treatment | BQE Water Inc.
Water pumping, and energy recovery | W. N. Brazier Assoc.
PST | Wood

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Basis of Estimate

Operating costs detailed in the 2020 PEA were derived from a variety of sources including, but not limited to, benchmarking analysis, derivation from first principles, and factoring costs from previous studies where possible.

The operating cost estimate is considered to have a level of accuracy of –25% to +35%. Overall assumptions for operating costs:

| ● | costs are presented in 2020 US dollars, unless stated otherwise. When required, certain costs in this report have been converted using a fixed currency exchange rate of Cdn$1.00 to US$0.76

| ● | the costs per dry metric tonne of material milled (US$/dmt) provided in this report are the average costs over the LOM.

The average LOM operating cost per tonne milled for the 2020 PEA is estimated at US$9.98/dmt. Details are presented in Table 24.29. The LOM average unit operating cost is calculated by dividing the total LOM operating costs by the LOM dmt milled. The costs exclude pre-production mining, reclamation and fund costs for servicing long-term water treatment that are accounted for separate from operating costs.

Table 24.29 2020 PEA Average Operating Costs

Operating Costs | Cost | Cost
(US$/dmt milled) | (LOM US$ million)
Mining (OP & UG) | 3.91 | 9,283
Milling/TMF/Train Transport | 4.32 | 10,250
G&A | 0.89 | 2,120
Site Services | 0.33 | 793
Water Management | 0.48 | 1,138
Annual Energy Recovery | (0.07) | (161)
PST | 0.11 | 269
Operating Costs/dmt milled | 9.98 | 23.692

Open Pit Mine Operating Costs

Open pit operating cost were estimated using an Excel® based cost model utilizing first principle build up for the haulage costs and benchmarking from a similar study for all other cost areas. The most significant components of the operating costs are fuel and labour. The average LOM cost for the open pit mine is US$1.53 per tonne mined. On a per tonne milled basis, the cost is US$3.96/dmt (see Table 24.30) Open pit mining represents 34% of the mill feed source.

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Table 24.30 Open Pit Mine Operating Costs

Open Pit Mining Operating Costs | Units | Cost
Total Cost | M US$ | 3,198
Cost per expit tonne mined | US$/t | 1.53
Cost per mill feed tonne milled | US$/dmt | 3.96

Underground Mine Operating Costs

The major components of the underground mining cost are labour and mobile equipment maintenance. Operating costs include the following indirect costs:

| ● | mine management, supervision, technical staff

| ● | freight on bulk materials

| ● | PPE and training.

Underground mining costs were estimated for each block cave operation using similar methodology as used for the capital cost estimate. The LOM cost per tonne mined varies between US$2.95 (Iron Cap) and US$4.72 (Deep Kerr). The Iron Cap mining costs are lower due to the application of battery electric loaders, electric haulage trains and partial automation of both loaders and trains, while the Deep Kerr mine employs a conventional diesel loader and truck haulage fleet. The cost reductions result primarily from lower diesel fuel, equipment maintenance and labour costs and less ventilation power demand. Iron Cap also benefits from its proximity to the MTT and lower conveying costs. The LOM underground mining cost per total dmt milled is estimated at US$3.88 (see Table 24.31). Underground mining represents 66% of the mill feed source.

Table 24.31 Underground Mine Operating Costs

Underground Mining Operating Costs | Units | Cost
Total Cost (LOM) | M US$ | 6,085
Cost per mill feed tonne milled | US$/dmt | 3.88

Process Operating Costs

The most significant process costs are reagents, grinding media, and power. Minor costs include labour, maintenance parts and assaying. The LOM average process cost is US$3.86/dmt milled. With the inclusion of tailing costs and train transport costs, the total cost for processing is US$4.32/dmt (Table 24.32).

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Table 24.32 Process Operating Costs

Cost Area | Cost
(US$/dmt milled)
Processing | 3.86
TMF | 0.24
Train Transport | 0.22
Total | 4.32

Reagent costs were based on test work and 2019 Q4 price quotes or in-house data. Grinding media and power costs are based on 2019 Q4 price quotes. Labour costs are based on a manpower list to reflect the 2020 PEA plant design while utilizing the same compensation structure from previous studies.

The costs are based on 170,000 t/d plant throughput. As annual processed tonnage decreases at the tail end of the mine life, the annual cost for reagents, steel consumables and power decrease relative to reduction in throughput. However, for the minor cost components the annual costs decrease more conservatively to reflect the requirement for minimum manpower in plant operations and other constraints.

Process costs are also specific to each deposit’s physical and metallurgical characteristics with each responding to differences in reagent, power and steel consumption. The annual cost reflects the proportional influence of each source of mill feed in the production schedule. These typical costs reflect both concentration paths: flotation for each of the mineral deposits and cyanidation if in operation. In general, if the mill feed is predominately from the Iron Cap and Kerr block cave deposits, there will be no cyanidation.

Tailing storage and handling costs are estimated at US$0.24/dmt milled and included in the processing operating cost. They are derived from the previous and current studies.

Tunnel and material handling costs for the MTT train are also included in the process operating cost. They represent US$0.22/dmt milled of the process operating cost.

General and Administrative Operating Costs

G&A costs are costs that do not relate directly to mining or processing operating costs. G&A costs include:

| ● | personnel: executive management, accounting, supply chain and logistics, human resources, external affairs functions, and other G&A departments

| ● | expenses: including insurance, off site offices, administrative supplies, medical services, legal services, human resources related expenses, community and environmental programs, accommodation/camp costs, air/bus crew transportation, regional and property taxes, and external assay/testing.

The G&A costs were estimated based on the 2020 PEA head count and factored costs from previous studies. G&A costs average US$0.89/dmt milled over the LOM.

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Site Services Operating Costs

Site services operating cost are estimated at US$0.33/dmt milled. They were derived from the previous and current studies and include the following elements:

| ● | personnel – general site services human power

| ● | site mobile equipment and light vehicle operations

| ● | portable water and waste management

| ● | general maintenance including yards, roads, fences, and building maintenance

| ● | off-site operation expense

| ● | building heating

| ● | power supply to surface facilities

| ● | avalanche control.

Water Management Operating Costs

Water management costs are estimated at US$0.48/dmt milled and were derived from previous and current studies and adjusted to reflect the 2020 PEA requirements and to reflect current commodity pricing. Table 24.33 shows the average LOM operating costs for the HDS treatment plant, the selenium water treatment plant, and the water pumping system.

Table 24.33 Water Management Costs

Description | US$/t
HDS Water Treatment | 0.29
Se Water Treatment | 0.17
Water Pumping | 0.02
Total | 0.48

Annual Energy Recovery Operating Costs

Annual energy recovery operating costs appear as a negative value to account for the revenue credit derived from on site power generation from energy recovery plants at the outlet of tunnels and on select pipelines. The value varies over the LOM and its average is estimated at US$-0.07/dmt.

Provincial Sales Tax (PST) Costs

Provincial sales tax (PST) is a retail sales tax that applies when a taxable good or service is purchased, acquired or brought into B.C., unless a specific exemption applies. The proportion of goods and services within the operating costs that are non-exempt from PST was determined in previous studies and applied to the 2020 PEA operating costs to estimate PST costs. The life of mine PST cost is estimated at US$ 0.11/dmt based on applying the rates shown in Table 24.34.

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Table 24.34 PST Rates

Area | Units | Cost
PST Mining | US$/t Mined | 0.044
PST Processing | US$/dmt Milled | 0.041

| 24.22 | Economic Analysis

The results of the economic analysis in the 2020 PEA represents forward-looking information that is subject to a number of known and unknown risks, uncertainties and other factors that may cause actual results to differ materially from those presented here. Forward-looking statements in the 2020 PEA section of this Report include, but are not limited to, timing and amount of future cash flows from mining operations, forecast production rates and amounts of copper, gold, and silver produced from the KSM mining operation, estimation of the Mineral Resources and the realization of the Mineral Resource estimates within the 2020 PEA mine plans, the time required to develop the mine based on the 2020 PEA mine design, statements with respect to future price of copper, gold and silver, currency exchange rate between the US dollars and Canadian dollars, assumptions regarding mine dilution and losses, the expected grade of the material delivered to the mill, metallurgical recovery rates, initial capital and sustaining capital costs, as well as mine closure costs and reclamation, timing and conditions of permits required to initiate mine construction, maintain mining activities, and mine closure, and assumptions regarding geotechnical and hydrogeological factors.

The reader is cautioned that the actual results of mining operations may vary from what is forecast. Risks to forward-looking information include, but are not limited to, unexpected variations in grade or geological continuity, as well as geotechnical and hydrogeological assumptions that are used in the mine designs. There could be seismic or water management events during the construction, operations, closure, and post-closure periods, that could affect predicted mine production, timing of the production, costs of future production, capital expenditures, future operating costs, permitting timelines, potential delays in the issuance of permits, or changes to existing permits, as well as requirements for additional capital. The plant, equipment or metallurgical or mining processes may fail to operate as anticipated. There may be changes to government regulation of mining operations, environmental issues, permitting requirements, and social risks, or unrecognized environmental, permitting and social risks, closure costs and closure requirements, unanticipated reclamation expenses, title disputes or claims and limitations on insurance coverage.

A portion of the Mineral Resources in the mine plans, production schedules, and cash flows include Inferred Mineral Resources, that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the 2020 PEA will be realized. Due to the conceptual nature of the 2020 PEA, none of the Mineral Resources in the 2020 PEA have been converted to Mineral Reserves and therefore do not have demonstrated economic viability.

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| 24.22.1 | Methodology Used

The 2020 PEA has been evaluated using a discounted cash flow (DCF) analysis. Cash inflows consist of annual revenue projections for the mine. Cash outflows such as capital, including the six years of pre-production costs, operating costs, taxes, and royalties are subtracted from the inflows to arrive at the annual cash flow projections. Cash flows are taken to occur at the end of each period.

To reflect the time value of money, annual net cash flow (NCF) projections are discounted back to the valuation date using several discount rates. The discount rate appropriate to a specific project depends on many factors, including the type of commodity; and the level of project risks related to mine construction and operation, such as market risk, technical risk and political risk. The discounted, present values of the cash flows are summed to arrive at the NPV.

In addition to NPV, IRR and payback period are also calculated. The IRR is defined as the discount rate that results in an NPV equal to zero. Payback is calculated as the time require to achieve positive cumulative cash flow following first metal production.

| 24.22.2 | Financial Model Parameters

Basis of Analysis

The financial analysis was based on the Mineral Resources presented in Section 14, the mine and process plan and assumptions detailed in Sections 24.16 and 24.17, the projected infrastructure requirements outlined in Section 24.18, the doré and concentrate marketing assumptions in Section 24.19, the permitting, social and environmental regime discussions in Section 24.20, and the capital and operating cost estimates detailed in Section 24.21.

Metal Pricing

The 2020 PEA metal pricing and US$/Cdn$ exchange rate are based on April 9, 2020 3-yr trailing averages. 3-yr trailing average pricing was likewise used for the 2016 PEA. For comparison, an alternate case is presented in Table 24.36 that utilizes the 2016 PEA metal prices and US$/Cdn$ exchange. An additional upside case is also presented in Table 24.36 that incorporates recent spot prices for gold, copper, silver and the US$/Cdn$ exchange rate. Metal prices for each scenario are presented in Table 24.35.

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Table 24.35 Metal Price Assumptions in 2020 PEA

Metal Prices | 2020 PEA Comparative Case | 2020 PEA 3 Yr Average* Base Case | 2020 PEA Alternate Upside Case
Gold (US$/oz) | 1,230 | 1,340 | 1,700
Copper (US$/lb) | 2.75 | 2.80 | 2.50
Silver (US$/oz) | 17.75 | 16.25 | 16.00
US$/Cdn$ Exchange Rate | 0.80 | 0.76 | 0.72

Note: *April 9, 2020 3 Year Trailing Average

Royalties

Under the Benefits Agreement with the Nisga’a Nation and the Co-operation and Benefits Agreement with the Tahltan Nation, the project owners make an annual financial payment to each Nation once operations have commenced. The combined annual payments to these Nations are payable in two forms; payments that are a percentage of the tax payable (the “Mineral Tax”) under the Mineral Tax Act (British Columbia) (the “Mineral Tax Act”), which is a tax on net operating profit, and payments that are based on net smelter returns. The LOM payments required by these agreements sum to approximately US$969 million.

Working Capital

Working capital cash outflow and inflows are included in the financial model. The calculations are based on the assumptions that accounts payable will be paid within 60 days and accounts receivable within 60 days. The impact of the working capital on the pre-tax NPV 5% is approximately US$140 million.

Closure and reclamation

Closure costs are estimated at US$938 million, including US$526 million set aside in a sinking fund from operating cash flow to pay for estimated post-closure water treatment obligations. This number leverages work completed for previous studies adjusted to reflect the 2020 PEA mine development plan. In addition to physical reclamation, closure and reclamation costs also include costs for bonds and costs for post-closure maintenance.

Taxes

The post-tax financial estimates consider all applicable Canadian Federal and BC Provincial taxes including:

| ● | Canadian Federal Income Tax

| ● | BC Provincial Tax

| ● | BC Mineral Tax

| ● | Provincial Sales Tax

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Canadian Federal and BC Provincial Income Tax Regime

The federal and BC provincial corporate income taxes are calculated using the currently enacted rates of 15% for federal and 12% for BC. Over the LOM, approximately US$4,947 million in federal and US$3,917 in provincial corporate income taxes are paid. For both federal and provincial income tax purposes, capital expenditures are accumulated in tax pools that can be deducted against mine income at different prescribed rates, depending on the type of capital expenditures.

BC Mineral Tax Regime

The BC Mineral Tax regime is a two-tier tax regime, with a 2% tax and a 13% tax. Over the LOM, approximately US$4,896 in BC Mineral Tax is paid.

The 2% tax is assessed on “net current proceeds”, which is defined as gross revenue from the mine less mine operating expenditures. The 13% tax is assessed on “net revenue”, which is defined as gross revenue from the mine less any accumulated expenditures.

BC Mineral Tax is deductible for federal and provincial income tax purposes.

Provincial Sales Tax

Provincial sales tax (PST) is a retail sales tax that applies when a taxable good or service is purchased, acquired or brought into B.C., unless a specific exemption applies. The proportion of goods and services within the capital costs that are non-exempt from PST was determined to be 18% in previous studies. The PST retail sales tax is 7%. For estimating PST that applies to capital, the 7% tax was applied to 18.5% of the annual capital and sustaining capital estimates within the financial model. The LOM PST for capital and sustaining capital costs was US$67 million and US$89 million respectively.

Financing

The model does not include any costs associated with financing.

Inflation

There is no adjustment for inflation in the financial model; all cash flows are based on 2020 dollars.

Financial ResultsTable 24.36 summarizes the financial results. The post-tax NPV at a 5% discount rate over the estimated mine life is US$6.01 billion. The post-tax IRR is 14.0%. The post-tax payback of the initial capital investment is estimated to occur in 4.0 years after the start of production. Table 24.37 shows the cash flow broken out on an annualized basis.

Note that the years presented in the table are for illustrative purposes only and do not necessarily represent commitment to the start dates or actual production.

The average life of mine operating cost per ounce of gold recovered is a negative US$-472 and US$511 on a byproduct and a coproduct basis, respectively; and the average life of mine operating cost per pound of copper recovered is a negative US$0.14 and US$1.17 on a byproduct and a coproduct basis, respectively. The total cost in Table 24.36 is calculated similar to the operating cost but includes US$13.1 billion in capital and closure costs.

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Table 24.36 2020 PEA Financial Analysis Summary

Financial Results | 2020 PEA Comparative Case | 2020 PEA Base Case 3yr Avg. | 2020 PEA Alternate Upside Case
Metal Prices: |  |  |
Gold (US$/ounce) | 1,230 | 1,340 | 1,700
Copper (US$/pound) | 2.75 | 2.80 | 2.50
Silver (US$/ounce) | 17.75 | 16.25 | 16.00
Exchange Rate: (US$/Cdn$) | 0.80 | 0.76 | 0.72
Cost Summary: |  |  |
Op cost (US$)/ Au oz recovered byproduct | (410) | (472) | (327)
Total cost (US$)/ Au oz recovered byproduct | 92 | 4 | 124
Op cost (US$)/ Au oz recovered coproduct | 498 | 511 | 605
Total cost (US$)/ Au oz recovered coproduct | 707 | 717 | 839
Op cost (US$)/ lb Cu recovered net of byproduct | 0.09 | (0.14) | (0.79)
Total cost (US$)/ lb Cu recovered net of byproduct | 0.90 | 0.63 | (0.06)
Op cost (US$)/ lb Cu recovered coproduct | 1.24 | 1.17 | 0.96
Total cost (US$)/ lb Cu recovered coproduct | 1.69 | 1.59 | 1.30
Initial Capital (US$ Billion) | 5.4 | 5.2 | 4.9
Sustaining Capital (US$ Billion) | 7.2 | 6.9 | 6.5
Unit Operating Cost On-site (US$/t milled) | 10.50 | 9.98 | 9.45
Pre-Tax Results: |  |  |
Net Cash Flow (US$ Billion) | 31.34 | 36.79 | 43.40
NPV @ 5% Discount Rate (US$ Billion) | 8.20 | 10.26 | 13.22
Internal Rate of Return (%) | 14.9% | 17.5% | 21.9%
Payback Period (Years) | 4.3 | 3.6 | 2.9
Post-Tax Results: |  |  |
Net Cash Flow (US$ Billion) | 19.58 | 23.03 | 27.22
NPV @ 5% Discount Rate (US$ Billion) | 4.67 | 6.01 | 7.92
Internal Rate of Return (%) | 11.9% | 14.0% | 17.4%
Payback Period (Years) | 4.9 | 4.0 | 3.2

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| 24.22.3 | Sensitivity Analysis

Sensitivity to metal prices, operating costs, capital costs, and recovery were analyzed for the Base Case post-tax cash flow, NPV 5%, and IRR. The sensitivities are shown graphically in Figure 24.34 to Figure 24.36 for post-tax undiscounted cash flow, NPV 5%, and IRR respectively.

Figure 24.34 Post-Tax Undiscounted Cash Flow Sensitivity Analysis

Source: Wood (2020)

Figure 24.35 Post-Tax NPV 5% Sensitivity Analysis

Source: Wood (2020)

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Figure 24.36 Post-Tax IRR Sensitivity Analysis

Source: Wood (2020)

The financials are most sensitive to changes in metal prices and grade and less sensitive to changes in capital costs and operating costs.

Due to the recent volatility in the commodities markets and the counter cyclical nature of gold and copper, alternate metal price cases are presented in Table 24.38 that independently vary the gold and copper prices. All alternate metal price cases assume the Base Case exchange rate, silver price, operating costs, and capital costs.

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Table 24.37 2020 PEA Cashflow on an Annualized Basis

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Table 24.38 Alternate Cases in 2020 PEA

Gold Price ($1,200/oz)
Copper Price (US$/lb Cu) | 2.25 | 2.50 | 2.75 | 3.00 | 3.25 | 3.50 | 4.00
Net CF (US$ B) | 24.4 | 28.3 | 32.3 | 36.3 | 40.3 | 44.2 | 52.2
Pre-Tax NPV5% (US$ B) | 6.2 | 7.4 | 8.6 | 9.8 | 11.0 | 12.2 | 14.6
Pre-Tax IRR (%) | 13.6% | 14.7% | 15.7% | 16.6% | 17.5% | 18.3% | 19.9%
Pre-Tax Payback (Years) | 4.5 | 4.3 | 4.1 | 3.9 | 3.8 | 3.7 | 3.5
Post-Tax NPV5% (US$ B) | 3.4 | 4.2 | 5.0 | 5.7 | 6.5 | 7.2 | 8.8
Post-Tax IRR (%) | 10.8% | 11.7% | 12.6% | 13.4% | 14.1% | 14.8% | 16.1%
Post-Tax Payback (Years) | 5.3 | 4.9 | 4.7 | 4.4 | 4.2 | 4.0 | 3.8
Op Cost Byproduct (US$/oz Au) | (161.5) | (305.6) | (449.7) | (593.9) | (738.1) | (882.3) | (1,170.7)
Total Cost Byproduct (US$/oz Au) | 315.0 | 170.9 | 26.8 | (117.4) | (261.6) | (405.8) | (694.1)
Op Cost Coproduct (US$/oz Au) | 530.8 | 502.8 | 477.7 | 455.1 | 434.5 | 415.9 | 383.1
Total Cost Coproduct (US$/oz Au) | 748.9 | 709.0 | 673.3 | 641.0 | 611.8 | 585.2 | 538.5
Op Cost Byproduct (US$/lb Cu) | 0.04 | 0.05 | 0.07 | 0.09 | 0.10 | 0.12 | 0.15
Total Cost Byproduct (US$/lb Cu) | 0.81 | 0.83 | 0.84 | 0.86 | 0.88 | 0.89 | 0.92
Op Cost Coproduct (US$/lb Cu) | 1.09 | 1.15 | 1.21 | 1.26 | 1.31 | 1.36 | 1.45
Total Cost Coproduct (US$/lb Cu) | 1.49 | 1.57 | 1.65 | 1.72 | 1.79 | 1.85 | 1.96
Gold Price ($1,350/oz)
Copper Price (US$/lb Cu) | 2.25 | 2.50 | 2.75 | 3.00 | 3.25 | 3.50 | 4.00
Net CF (US$ B) | 28.3 | 32.3 | 36.3 | 40.2 | 44.2 | 48.2 | 56.1
Pre-Tax NPV5% (US$ B) | 7.7 | 8.9 | 10.1 | 11.3 | 12.5 | 13.7 | 16.1
Pre-Tax IRR (%) | 15.5% | 16.5% | 17.5% | 18.3% | 19.1% | 19.9% | 21.4%
Pre-Tax Payback (Years) | 3.9 | 3.8 | 3.6 | 3.5 | 3.4 | 3.4 | 3.2
Post-Tax NPV5% (US$ B) | 4.4 | 5.2 | 5.9 | 6.7 | 7.4 | 8.2 | 9.7
Post-Tax IRR (%) | 12.3% | 13.2% | 13.9% | 14.7% | 15.4% | 16.0% | 17.3%
Post-Tax Payback (Years) | 4.4 | 4.2 | 4.0 | 3.9 | 3.8 | 3.7 | 3.5
Op Cost Byproduct (US$/oz Au) | -154.8 | -299.0 | -443.2 | -587.4 | -731.6 | -875.8 | -1164.3
Total Cost Byproduct (US$/oz Au) | 321.7 | 177.5 | 33.3 | -110.9 | -255.1 | -399.3 | -687.8
Op Cost Coproduct (US$/oz Au) | 571.8 | 543.6 | 518.2 | 495.2 | 474.2 | 455.1 | 421.2
Total Cost Coproduct (US$/oz Au) | 804.0 | 763.8 | 727.5 | 694.7 | 664.8 | 637.5 | 589.2
Op Cost Byproduct (US$/lb Cu) | -0.20 | -0.18 | -0.16 | -0.15 | -0.13 | -0.12 | -0.09
Total Cost Byproduct (US$/lb Cu) | 0.58 | 0.59 | 0.61 | 0.63 | 0.64 | 0.66 | 0.69
Op Cost Coproduct (US$/lb Cu) | 1.03 | 1.10 | 1.16 | 1.21 | 1.26 | 1.31 | 1.40
Total Cost Coproduct (US$/lb Cu) | 1.41 | 1.50 | 1.58 | 1.65 | 1.71 | 1.78 | 1.89

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Gold Price ($1,500/oz)
Copper Price (US$/lb Cu) | 2.25 | 2.50 | 2.75 | 3.00 | 3.25 | 3.50 | 4.00
Net CF (US$ B) | 32.3 | 36.2 | 40.2 | 44.2 | 48.2 | 52.1 | 60.1
Pre-Tax NPV5% (US$ B) | 9.2 | 10.4 | 11.6 | 12.8 | 14.0 | 15.2 | 17.6
Pre-Tax IRR (%) | 17.4% | 18.3% | 19.1% | 19.9% | 20.7% | 21.4% | 22.8%
Pre-Tax Payback (Years) | 3.5 | 3.4 | 3.3 | 3.3 | 3.2 | 3.1 | 3.0
Post-Tax NPV5% (US$ B) | 5.4 | 6.1 | 6.9 | 7.6 | 8.4 | 9.2 | 10.7
Post-Tax IRR (%) | 13.8% | 14.5% | 15.3% | 16.0% | 16.6% | 17.2% | 18.4%
Post-Tax Payback (Years) | 3.9 | 3.8 | 3.7 | 3.6 | 3.5 | 3.4 | 3.3
Op Cost Byproduct (US$/oz Au) | -148.3 | -292.5 | -436.7 | -580.9 | -725.2 | -869.4 | -1157.9
Total Cost Byproduct (US$/oz Au) | 328.2 | 184.0 | 39.8 | -104.4 | -248.7 | -392.9 | -681.4
Op Cost Coproduct (US$/oz Au) | 609.3 | 581.1 | 555.6 | 532.3 | 511.0 | 491.5 | 456.9
Total Cost Coproduct (US$/oz Au) | 854.0 | 813.8 | 777.3 | 744.1 | 713.8 | 685.9 | 636.5
Op Cost Byproduct (US$/lb Cu) | -0.43 | -0.41 | -0.40 | -0.38 | -0.37 | -0.35 | -0.32
Total Cost Byproduct (US$/lb Cu) | 0.35 | 0.36 | 0.38 | 0.39 | 0.41 | 0.42 | 0.46
Op Cost Coproduct (US$/lb Cu) | 0.99 | 1.05 | 1.11 | 1.16 | 1.21 | 1.26 | 1.35
Total Cost Coproduct (US$/lb Cu) | 1.35 | 1.43 | 1.51 | 1.58 | 1.65 | 1.71 | 1.82
Gold Price ($1,650/oz)
Copper Price (US$/lb Cu) | 2.25 | 2.50 | 2.75 | 3.00 | 3.25 | 3.50 | 4.00
Net CF (US$ B) | 36.2 | 40.2 | 44.2 | 48.1 | 52.1 | 56.1 | 64.1
Pre-Tax NPV5% (US$ B) | 10.7 | 11.9 | 13.1 | 14.3 | 15.5 | 16.7 | 19.1
Pre-Tax IRR (%) | 19.1% | 20.0% | 20.8% | 21.5% | 22.2% | 22.9% | 24.2%
Pre-Tax Payback (Years) | 3.2 | 3.2 | 3.1 | 3.0 | 3.0 | 2.9 | 2.8
Post-Tax NPV5% (US$ B) | 6.3 | 7.1 | 7.8 | 8.6 | 9.4 | 10.1 | 11.6
Post-Tax IRR (%) | 15.2% | 15.9% | 16.5% | 17.2% | 17.8% | 18.4% | 19.5%
Post-Tax Payback (Years) | 3.6 | 3.5 | 3.4 | 3.3 | 3.3 | 3.2 | 3.1
Op Cost Byproduct (US$/oz Au) | -141.9 | -286.1 | -430.4 | -574.6 | -718.9 | -863.1 | -1151.6
Total Cost Byproduct (US$/oz Au) | 334.7 | 190.4 | 46.2 | -98.1 | -242.3 | -386.6 | -675.1
Op Cost Coproduct (US$/oz Au) | 643.7 | 615.7 | 590.1 | 566.8 | 545.3 | 525.6 | 490.4
Total Cost Coproduct (US$/oz Au) | 899.7 | 859.7 | 823.2 | 789.8 | 759.2 | 731.0 | 680.8
Op Cost Byproduct (US$/lb Cu) | -0.66 | -0.65 | -0.63 | -0.61 | -0.60 | -0.58 | -0.55
Total Cost Byproduct (US$/lb Cu) | 0.11 | 0.13 | 0.14 | 0.16 | 0.18 | 0.19 | 0.22
Op Cost Coproduct (US$/lb Cu) | 0.94 | 1.01 | 1.06 | 1.12 | 1.17 | 1.22 | 1.31
Total Cost Coproduct (US$/lb Cu) | 1.28 | 1.37 | 1.45 | 1.52 | 1.58 | 1.65 | 1.76

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Gold Price ($1,800/oz)
Copper Price (US$/lb Cu) | 2.25 | 2.50 | 2.75 | 3.00 | 3.25 | 3.50 | 4.00
Net CF (US$ B) | 40.2 | 44.1 | 48.1 | 52.1 | 56.1 | 60.1 | 68.0
Pre-Tax NPV5% (US$ B) | 12.3 | 13.5 | 14.7 | 15.9 | 17.1 | 18.3 | 20.6
Pre-Tax IRR (%) | 20.8% | 21.6% | 22.3% | 23.0% | 23.7% | 24.4% | 25.6%
Pre-Tax Payback (Years) | 3.0 | 2.9 | 2.9 | 2.8 | 2.7 | 2.7 | 2.6
Post-Tax NPV5% (US$ B) | 7.3 | 8.0 | 8.8 | 9.6 | 10.3 | 11.1 | 12.6
Post-Tax IRR (%) | 16.5% | 17.2% | 17.8% | 18.4% | 19.0% | 19.5% | 20.6%
Post-Tax Payback (Years) | 3.3 | 3.2 | 3.2 | 3.1 | 3.0 | 3.0 | 2.9
Op Cost Byproduct (US$/oz Au) | (135.5) | (279.8) | (424.0) | (568.3) | (712.5) | (856.8) | (1,145.3)
Total Cost Byproduct (US$/oz Au) | 341.0 | 196.7 | 52.5 | (91.8) | (236.0) | (380.3) | (668.8)
Op Cost Coproduct (US$/oz Au) | 675.5 | 647.7 | 622.3 | 598.9 | 577.4 | 557.6 | 522.1
Total Cost Coproduct (US$/oz Au) | 941.8 | 902.0 | 865.7 | 832.3 | 801.6 | 773.3 | 722.5
Op Cost Byproduct (US$/lb Cu) | -0.89 | -0.88 | -0.86 | -0.85 | -0.83 | -0.82 | -0.79
Total Cost Byproduct (US$/lb Cu) | -0.12 | -0.10 | -0.09 | -0.07 | -0.06 | -0.04 | -0.01
Op Cost Coproduct (US$/lb Cu) | 0.90 | 0.96 | 1.02 | 1.08 | 1.13 | 1.18 | 1.27
Total Cost Coproduct (US$/lb Cu) | 1.23 | 1.31 | 1.39 | 1.46 | 1.53 | 1.59 | 1.71
Gold Price ($2,000/oz)
Copper Price (US$/lb Cu) | 2.25 | 2.50 | 2.75 | 3.00 | 3.25 | 3.50 | 4.00
Net CF (US$ B) | 45.4 | 49.4 | 53.4 | 57.4 | 61.4 | 65.3 | 73.3
Pre-Tax NPV5% (US$ B) | 14.3 | 15.5 | 16.7 | 17.9 | 19.1 | 20.3 | 22.7
Pre-Tax IRR (%) | 23.0% | 23.7% | 24.3% | 25.0% | 25.6% | 26.2% | 27.4%
Pre-Tax Payback (Years) | 2.7 | 2.7 | 2.6 | 2.6 | 2.5 | 2.5 | 2.4
Post-Tax NPV5% (US$ B) | 8.6 | 9.3 | 10.1 | 10.8 | 11.6 | 12.3 | 13.9
Post-Tax IRR (%) | 18.2% | 18.8% | 19.4% | 19.9% | 20.5% | 21.0% | 22.0%
Post-Tax Payback (Years) | 3.0 | 3.0 | 2.9 | 2.9 | 2.8 | 2.8 | 2.7
Op Cost Byproduct (US$/oz Au) | -127.1 | -271.3 | -415.6 | -559.8 | -704.1 | -848.3 | -1136.8
Total Cost Byproduct (US$/oz Au) | 349.4 | 205.2 | 60.9 | -83.3 | -227.6 | -371.8 | -660.3
Op Cost Coproduct (US$/oz Au) | 714.4 | 687.0 | 661.9 | 638.7 | 617.2 | 597.4 | 561.6
Total Cost Coproduct (US$/oz Au) | 992.9 | 953.8 | 917.8 | 884.7 | 854.0 | 825.5 | 774.4
Op Cost Byproduct (US$/lb Cu) | -1.21 | -1.19 | -1.17 | -1.16 | -1.14 | -1.13 | -1.10
Total Cost Byproduct (US$/lb Cu) | -0.43 | -0.42 | -0.40 | -0.38 | -0.37 | -0.35 | -0.32
Op Cost Coproduct (US$/lb Cu) | 0.86 | 0.92 | 0.97 | 1.03 | 1.08 | 1.13 | 1.22
Total Cost Coproduct (US$/lb Cu) | 1.16 | 1.24 | 1.32 | 1.39 | 1.46 | 1.52 | 1.64

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| 25.0 | Interpretations and Conclusions

| 25.1 | Introduction

This Report presents a summary of the results of two separate alternative studies for two separate mine development options for KSM. The first study is the 2016 PFS prepared for Seabridge in 2016. The Qualified Persons (QPs) of this Report have verified that the 2016 PFS remains current. The second study is the 2020 PEA which has been prepared as an alternative option to the 2016 PFS. The 2020 PEA evaluates a different approach to KSM mine development by emphasizing low-cost block cave mining and reducing the number and size of the open pits, that significantly reduces the surface disturbance of the KSM mine. The 2020 PEA assesses the potential impacts of incorporating higher grade Inferred Mineral Resources delineated at Deep Kerr and Iron Cap into the mine design, and increasing the annual average maximum mill throughput from 130,000 t/d designed in the 2016 PFS to 170,000 t/d in the 2020 PEA.

The results of the 2016 PFS remain valid and represent a viable option for developing KSM, with the 2020 PEA assessing an alternative development option at a conceptual level.

KSM received its EA approvals from both the provincial and federal governments in July and December of 2014, respectively, including numerous provincial permits covering the first years of construction. Those reviews and subsequent decisions concluded that KSM would not result in significant adverse effects to the environment, identifying KSM as a responsible mine development.

The Mineral Resources were constrained within the various KSM mineralized zones using open pit shells and block cave shapes to establish reasonable prospects for eventual economic extraction as outlined in the CIM Definition Standards for Mineral Resources and Mineral Reserves (CIM, 2014). Details of the Mineral Resources are presented in Section 14.0.

The 2016 PFS maintains the alternative of KSM as a large-tonnage open pit and underground block cave mining operation, at a nominal rate of 130,000 t/d of ore fed to the processing plant capable of producing a copper/gold/silver concentrate for transport by truck to the nearby deep-water sea port at Stewart, BC. A gold-silver doré, and a separate molybdenum concentrate, will also be produced at the processing facility.

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| 25.2 | 2016 Prefeasibility Study Conclusions

| 25.2.1 | 2016 PFS Data Verification

The QPs for the 2016 PFS concluded that the Mineral Reserves would not materially change if the mine plans were updated using the December 2019 Mineral Resource estimates, Q4 2019 capital and operating costs, and 2020 metal prices and metallurgical recoveries. Therefore the 2016 PFS Mineral Reserves are verified as being current and are suitable to be used unchanged in the 2016 PFS that is to be included in the 2020 Technical Report on the KSM Property.

The QPs verified that if the mine capital and operating cost estimates, metal prices and exchange rates were updated, any changes would not cause a material difference in the outcomes of the pre-tax and post-tax economic analyses of the 2016 PFS on the KSM property. Therefore the 2016 PFS pre-tax and post-tax economic analyses are considered current and suitable to be used unchanged in the 2020 Technical Report.

| 25.2.2 | Mineral Reserves

From the section above, the stated reserves have not been changed from the 2016 PFS Technical Report even though there have been changes to the Mineral Resource model described in Section 14.0. Mineral Reserves are based on NSR values that were calculated in the block model using US$1,200.00/oz of gold, US$2.70/lb of copper, US$17.50/oz of silver, US$9.70/lb of molybdenum and a foreign exchange rate of US$0.83 per Cdn$1.00, with varying process recoveries for the different mining areas, and applicable off-site charges. The NSR values were used as a dynamic cut-off grade for defining ore and waste in the open pit, with a minimum of Cdn$9.00/t. Mining loss and dilution parameters for the open pits and underground mine plans are applied as described in Section 15.0. The estimated Proven and Probable Mineral Reserves as of July 31, 2016 are 38.8 Moz of gold and 10.2 Blb of copper (2.2 Bt at an average grade of 0.55 g/t gold, 0.21% copper, 2.6 g/t silver and 42.6 ppm molybdenum per tonne).

The methods used in this estimate comply with CIM Definition Standards, with reasonable engineering practices for a PFS-level study, and economic estimates based on the technical and economic parameters stated in this Report.

| 25.2.3 | Mining Methods

25.2.3.1 Open Pit Mining

The open pit mine plan in the 2016 PFS establishes the economic mining limits of the Mitchell, Sulphurets, and Kerr Mineral Resource areas using large-tonnage mining methods capable of providing mill feed at a nominal rate of 130,000 t/d when the open pit is supplying all ore. The LOM plan accommodates the local adverse conditions comprising snow, cold, remoteness, and steep terrain. Waste and water management designs are incorporated into the mine plan, as specified in the current site plans, as reviewed and approved in the Application/EIS (Rescan 2013) review process completed in 2014.

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The chosen mine equipment is well known and suitable for the expected operating conditions, and the productivity assumptions are reasonable and achievable. The resultant unit mining costs are comparable when benchmarked against other similar operating mines when considering the site operating conditions. Given the stated design parameters and assumptions, the open pit mine plan would achieve the forecast production schedule and the annual levels and costs within the expected range of accuracy of the 2016 PFS estimate.

25.2.3.2 Underground Mining

The 2016 PFS underground block cave mining plans for the Mitchell deposit, beneath the previously mined Mitchell open pit, and for the Iron Cap deposit were developed from assessments of the economic shut-off at the drawpoints using GEOVIA’s PCBC™ software. These assessments incorporate dilution associated with sub-economic and waste rock that mixes with the mineralized rock above shut-off value as it is drawn down and mucked at the drawpoint. The GEOVIA PCBC™ assessments were based on industry standard approaches of developing production and grade schedules, and incorporated benchmarked production controls for ramp-up rates, maximum draw rates, and drawpoint construction rates. The size of the block cave footprints on the production levels at Mitchell and Iron Cap are much larger than the minimum estimated area required to initiate and propagate the cave to surface, and so there are no concerns about these deposits caving.

The drawpoint layout and spacing were based on estimates of fragmentation and cave mining experience in maintaining favourable interaction between adjacent draw columns that control premature reporting of waste dilution at the drawpoints. The chosen mucking equipment, the haulage distances from drawpoints to ore passes, the handing of oversize muck by secondary blasting, and the design of the ore passes and grizzly system allow for production rates of 55,000 t/d at Mitchell and 40,000 t/d at Iron Cap. The production ramp-up periods for these operations are six years and four years, respectively, which are within the ranges that have been achieved at other block cave operations. Production from Mitchell, and subsequently from Iron Cap, allow the total KSM production to be maintained at 130,000 t/d through Year 35 as production from the open pits decrease. Production from underground then dominates mill feed for the final third of the LOM.

Based on the various factors discussed above, the block cave mine plan is expected to achieve the forecasted production schedule and the annual levels and costs within the expected range of accuracy of the 2016 PFS estimate.

| 25.2.4 | Recovery Methods

Several wide-ranging metallurgical test programs have been conducted since 2007 to assess the metallurgical responses of the mineral samples from the KSM deposits, especially the samples from the Mitchell deposit. The test results indicate that the mineral samples from the four separate mineralized deposits are amenable to the flotation-cyanidation combined process.

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The 2016 PFS processing plant is designed based on the flowsheet developed from the testwork results. The proposed flotation process is projected to produce a copper-gold concentrate containing approximately 25% copper. Copper and gold flotation recoveries will vary with changes in head grade and mineralogy. The average copper and gold recoveries to the concentrate are projected to be 81.6% and 55.3%, respectively, for the LOM mill feed containing 0.55 g/t gold and 0.21% copper. As projected from the testwork, the cyanidation circuit will increase the overall gold recovery to a range of 60 to 79%, depending on gold and copper head grades. Silver recovery from the flotation and leaching circuits is expected to be 62.7% on average. A separate flotation circuit will recover molybdenite from the copper-gold-molybdenum bulk concentrate when higher-grade molybdenite mineralization is processed.

The process flowsheet proposed for KSM is conventional and has been widely used in processing porphyry copper-gold ores. The equipment type and sizing selected for KSM are common in other mining projects.

In general, the copper and molybdenum concentrates produced are anticipated to be acceptable by most of the copper and molybdenum smelters. On average, the impurity contents in the copper and molybdenum concentrates should be lower than the penalty thresholds set by most of the smelters.

| 25.2.5 | Project Infrastructure

25.2.5.1 MTT Transportation System

In the 2016 PFS, transportation of ore, freight, and personnel through the MTT will be achievable with a two tunnel automated train transport system, at an average rate of 130,000 t/d and a peak capacity of 10,000 t/h. The required deliveries of freight, including bulk transport of fuel and lime, mine site consumables, as well as personnel movement for periodic crew changes and daily requirements between the Mitchell OPC and Treaty OPC are also achievable using the tunnel and rail infrastructure with appropriate traffic control management systems.

The system is scalable and flexible, as it has the ability to add or remove train sets to meet higher or lower throughput requirements, and components can be taken out of operation for maintenance without compromising total system operation.

This scale of underground tunnel transport via trains has proven effective in other mines globally, utilizing equipment and infrastructure specified within the 2016 PFS. The planned train system operations can be reasonably accomplished at the 2016 PFS’s estimated productivities and costs.

25.2.5.2 Tunnelling

For 2016 PFS, the conventional drill and blast methodology for excavating infrastructure and water tunnels is a valid basis for the scheduling and costing of the long tunnels required in this PFS. Using a twinned tunnel for the MTT provides advantages in construction with three sets of advancing twin headings, enabling the use of one tunnel at each heading as a fresh airway, and the other tunnel as a return airway, which has a significant impact on ventilation and advance rates for long tunnels. Other infrastructure and water tunnels would be excavated with a single advancing heading from each portal, a well proven method in the mining industry.

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Advance rates and excavation costs for the MTT have been determined by two contractors using detailed cycle time calculations for the varying ground conditions. The contractor-developed advance rates and costs have been adapted for the other tunnel excavations. The contractor’s estimates have also been benchmarked, indicating the estimates are within the accuracy of the 2016 PFS. The License of Occupation for the MTT route was issued by the BC Government in September 2014.

25.2.5.3 Infrastructure Dams

Tailings Management Facility

In 2016 PFS, the TMF will be a conventional storage facility with three main cells, using a starter dam construction with ongoing centerline dam raises using cyclone sand. The TMF layout and sequencing allows for staged construction of the facility, which optimizes both start-up and operating costs, while managing geotechnical and environmental risks. Using a three-cell system provides advantages for storage of CIL tailings, and allows progressive, staged development followed by closure of the facility to mitigate environmental and closure risks. The cell system reduces TMF discharge of excess water and also allows for progressive development of water management facilities.

A Best Available Tailings Technology (BATT) assessment (KCB, 2016a), completed at the same time as the 2016 PFS, shows that the current TMF design is the most appropriate to meet environmental, operability, and geotechnical criteria. Dry stack options were either not the most appropriate or did not meet these criteria.

Key conclusions include:

| ● | TMF starter dams safely store 18 to 24 months of tailings under the range of start-up assumptions assessed as established by the TMF design criteria. The storage plan allows for operational flexibility during start up, with a minimum of one winter season and potentially two winter seasons accommodated

| ● | TMF dam designs (from starter dams through closure) are stable under static and pseudo-static conditions, as designed, and these designs meet all applicable regulatory criteria.

Water Storage Facility

The WSF provides environmental containment of runoff water for the mine site. The facility includes a rock fill-asphalt core WSD to retain contact water collected from the mine site for treatment at the HDS WTP. The facility is capable of storing inflows from extreme events and is designed to minimize seepage. The WSD will be built to full height before start-up to provide 50 Mm3 of storage for a 200-year wet year runoff from the mine site catchments. Key conclusions include:

| ● | the rock fill-asphalt core WSD design is confirmed to be the preferred structure type to meet the KSM environmental, durability, and geotechnical design criteria. A value engineering study (KCB, 2012d) showed that this design has both low-seepage rates, as well as constructability and cost advantages over other types of dam structures analyzed for this location and purpose

| ● | in 2016, a dam construction contractor provided improved reliability estimates of structure cost and construction schedule duration for completion of the WSD.

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| 25.2.6 | Economic Analysis

Tetra Tech prepared an economic evaluation for the 2016 PFS based on a pre-tax financial model. The tax component of the model was prepared and reviewed by other consultants (please see Section 22.0 for further details). Based on this tax analysis, Tetra Tech prepared the post-tax economic evaluation of the 2016 PFS.

For the 53-year LOM and 2.198 Bt Mineral Reserve, Table 25.1 summarizes the economic analysis results presented in Section 22.3 and 22.5.

Based on the economic results of Table 25.1, it can be concluded that the 2016 PFS:

| ● | has a short payback period compared to the long mine life

| ● | has low cash and total costs per ounce of gold produced net of by-product credits

| ● | costs include capital cost commitments resulting from the EA review process and thus represent a realistic cost basis moving forward as compared to those projects which have yet to complete an environmental review

| ● | closure costs including long term water treatment costs are included in the economic evaluation

| ● | based on study results herein, is considered economically viable, and thus merits additional study in the next design phase.

| Table 25.1 Summary of Major Pre- and Post-tax Results by Metal Price Scenario

| Unit | 2016 Base Case | 2016 Design Case | 2016 Spot Case | 2016 Alternate Case
Metal Price
Gold | US$/oz | 1,230.00 | 1,200.00 | 1,350.00 | 1,500.00
Copper | US$/lb | 2.75 | 2.70 | 2.20 | 3.00
Silver | US$/oz | 17.75 | 17.50 | 20.00 | 25.00
Molybdenum | US$/lb | 8.49 | 9.70 | 7.00 | 10.00
Exchange Rate | US$:Cdn$ | 0.80 | 0.83 | 0.77 | 0.80
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| Unit | 2016 Base Case | 2016 Design Case | 2016 Spot Case | 2016 Alternate Case
Pre-tax Results
Cumulative Cashflow | US$ million | 15,933 | 13,727 | 16,101 | 26,319
NPV (at 3%) | US$ million | 6,217 | 5,128 | 6,461 | 11,138
NPV (at 5%) | US$ million | 3,263 | 2,510 | 3,507 | 6,541
NPV (at 8%) | US$ million | 960 | 475 | 1,175 | 2,928
IRR | % | 10.4 | 9.2 | 11.1 | 14.6
Payback | years | 6.0 | 6.5 | 5.6 | 4.1
Cash Cost/oz Au | US$/oz | 277 | 311 | 404 | 183
Total Cost/oz Au | US$/oz | 673 | 720 | 787 | 580
Post-tax Results
Cumulative Cashflow | US$ million | 9,983 | 8,537 | 10,109 | 16,721
NPV (at 3%) | US$ million | 3,513 | 2,789 | 3,691 | 6,696
NPV (at 5%) | US$ million | 1,539 | 1,028 | 1,718 | 3,663
NPV (at 8%) | US$ million | -2 | -343 | 161 | 1,282
IRR | % | 8.0 | 7.0 | 8.5 | 11.4
Payback | years | 6.8 | 7.4 | 6.4 | 4.9

| Notes: | Operating and total costs per ounce of gold are after base metal credits. Total costs per ounce include all start-up capital, sustaining capital, and reclamation/closure costs.

| 25.3 | 2020 Preliminary Economic Assessment Conclusions

| 25.3.1 | Mining Methods

The 2020 PEA mining study took a different approach to the 2016 PFS. The 2020 PEA mine plan was carried out with the aim of reducing the amount of waste rock produced in the open pits with the mill feed drawing more on the underground resources. The 2020 PEA design is a combined open pit/underground block cave mining operation that is planned to operate for 44 years. The mine production plan starts in lower-cost open pit areas using conventional large scale equipment before transitioning into block cave underground bulk mining later in the mine life. Relatively smaller pits were designed for the Mitchell, Sulphurets and Kerr deposits as well as mining the Iron Cap and Deep Kerr deposits by underground mining methods. This approach shrinks the KSM mine footprint

The Mitchell open pit and Iron Cap and Deep Kerr underground mines will be the main source of mill feed, contributing approximately 92% of the total plant feed over the LOM, supplemented by the Sulphurets and Kerr open pits in the mine production plan.

25.3.1.1 Open Pit Mining

The 2020 PEA open pit mining is designed as a conventional truck-shovel operation. The pit design for the Mitchell area includes three nested phases to balance stripping requirements while satisfying the process plant requirements. Open pit production will be carried out to meet the 170 ktpd mill feed requirement incorporating the underground production schedules. Mitchell open pit will be the only operating mine area during initial capital payback. Mining will then transition to Iron Cap and Deep Kerr underground mines, supplemented by Kerr and Sulphurets open pits during the transition. Mill feed will be derived entirely from Iron Cap and Deep Kerr underground mines from Year 17 to Year 28 when Iron Cap production begins to steadily decrease. The M3 Mitchell pit will be mined from Year 26 to 34 and the remaining Deep Kerr underground will continue providing mill feed at 80,000 tpd until depleted by the end of the life of mine (LOM) in Year 44.

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25.3.1.2 Underground Mining

The 2020 PEA underground block caving mine designs for Iron Cap and Deep Kerr are based on modeling using GEOVIA’s PCBC™ Footprint Finder software. Iron Cap has been designed for using battery-electric loaders and electric trains for material handling and employs both tele-operation and automation for these units. This enables a reduction to the number of primary ventilation intake and exhaust drifts due to less ventilation demand and cost savings from lower diesel fuel, labour and equipment maintenance costs.

The first underground mill feed production from Iron Cap and Deep Kerr comes in Years 6 and 12, respectively. Iron Cap is estimated to have a production ramp-up period of six years, steady state production at 32.9 Mt/a for 17 years, and then ramp-down production for another six years. Deep Kerr is estimated to have a production ramp-up period of five years, steady state production at 29.2 Mt/a for 20 total years with a seven year production dip during years where the operation transitions from the first to second lift. Underground production ends first at Iron Cap in Year 34 and finally at Deep Kerr in Year 44 of production.

| 25.3.2 | Recovery Methods

Based on the available information, test work performed on various samples provides a reasonable indication of the mineralogical characteristics of the material examined in the 2020 PEA for both the surface and underground mine extraction scenarios. The process flowsheet considered in the prior studies is also appropriate for the 2020 PEA given the nature of the mill feed and the level of sampling performed through the deposits.

Several metallurgical test programs have been carried out to assess the recoverability of copper, gold and silver values using the 2020 PEA flowsheet. The results of the programs performed on KSM’s mineralized material indicate that the mineral samples from the Mitchell, Sulphurets, Iron Cap and Deep Kerr deposits are amenable to flotation followed by a cyanide leaching process.

Detailed characterization and metallurgical test work on Mitchell, Sulphurets, Iron Cap, and Deep Kerr samples are presented in the pertinent test work reports, which are listed in Section 27. These test work examined comminution (crushing and grinding), concentration by flotation of copper, gold and silver into a saleable concentrate, and the further recovery of gold and silver from flotation tailings by leaching. Metallurgical performance projections from this test work were produced by Tetra Tech. Wood reviewed this test work and considers it adequate for the purposes of reference for the 2020 PEA. Two Kerr and one Iron Cap copper recovery models showed clear relationship of copper flotation recovery to copper head grade. These results can be considered as confirmatory. For gold, there was no clear relationship between gold recoveries vs. copper recovery or gold head or copper head grades.

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Comminution test work results indicate that the samples from all the deposits are moderately hard for SAG and ball milling. Comminution tests also showed that the samples tested were amenable to particle size reduction by the HPGR approach as preparation for ball mill grinding.

Flotation test work (batch and locked cycle tests) indicates that the mineralization is amenable for concentration into a saleable copper-gold concentrate with no significant penalty elements. Following flotation, cyanidation tests showed that it was possible to recover gold and silver bullion by CIL leaching of gold-rich pyrite concentrates derived from pyrite flotation of rougher flotation tailing and tailing from first cleaner scavenger separation. Trade-off studies indicate that cyanidation of Kerr and Iron Cap gold-bearing pyrite materials may not be economic due to elevated copper/gold ratios and the resulting ineffective CIL carbon absorption. Additional test work and economic assessments are required to validate this observation. Cyanidation of gold bearing pyrite materials are therefore limited to Mitchell and Sulphurets for the 2020 PEA. The metallurgical test results obtained from the various test programs have been used to predict plant metallurgical performance parameters for copper, gold and silver. The metallurgical performance projections of the four KSM mineralized zones are summarized in prior studies. In addition, work has been performed to determine the consumption of reagents and grinding media in comminution, flotation and cyanidation. Test work on cyanide recovery was also performed and the results have been incorporated into the operating costs.

| 25.3.3 | Project Infrastructure

Tailings and Water Management

The water and waste management structure designs in the 2020 PEA are based on those described in the approved EA with conceptual-level adjustments to the 2020 PEA mine plan layouts, staging plan, and capacities required for the 2020 PEA mine production schedule.

The 2020 PEA rock storage requirements are significantly less than in the approved EA due to waste reductions during open pit phases of mining (1.3 Bt versus 3.0 Bt). The 2020 PEA waste management plan consolidates all mine waste rock into a single facility, including minor waste volume from initial underground development.

The overall 2020 PEA operational water management and water treatment designs are similar to those in the approved EA, with adjustments in some areas based on the revised mine plan and recently collected field data. For closure, the 2020 PEA water management plan remains the same as in the approved EA, except for the elimination of McTagg RSF and the associated closure channels.

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The WSD crest elevation is kept at the same elevation for the 2020 PEA. With the reduced tonnage of waste rock, WSD storage and WSD crest elevation could potentially be reduced, although these optimizations have not been adopted for the 2020 PEA.

TMF configurations for the 2020 PEA are generally consistent with the TMF designs and the staged cell configuration presented in the approved EA, except for slight modifications in starter dam heights and sizing of the CIL cell.

| 25.3.4 | Economic Analysis

The 2020 PEA metal pricing and US$/C$ exchange rate are based on April 9, 2020 3-yr trailing averages. For comparison, an alternate case is presented that utilizes the 2016 PEA metal prices and US$/C$ exchange rate. An additional upside case is also presented in Table 25.2 that incorporates recent spot prices for gold, copper, silver and US$/C$ exchange rate. For all cases, the 2020 PEA financials are positive on both a Pre-tax and Post-tax basis Table 25.2 provides financial results for all cases. For the base case, additional key financial metrics include:

| ● | 4.0-year payback on US$5.2 B initial capital

| ● | average annual gold production during payback over 1.37 Million oz

| ● | average annual pre-tax Free Cash Flow of US$1.45 Billion from 1.34 Million oz gold and 265 Million pounds copper produced per year during the initial 5-years of production

| ● | first five year operating cost profile co-product of US$210/oz Au or US$0.86/lb Cu.

| Table 25.2 2020 PEA Financial Analysis Summary

Financial Results | 2020 PEA Comparative Case | 2020 PEA Base Case 3yr Avg. | 2020 PEA Alternate Upside Case
Metal Prices: |  |  |
Gold (US$/ounce) | 1,230 | 1,340 | 1,700
Copper (US$/pound) | 2.75 | 2.80 | 2.50
Silver (US$/ounce) | 17.75 | 16.25 | 16.00
Exchange Rate: (US$/Cdn$) | 0.80 | 0.76 | 0.72
Pre-Tax Results: |  |  |
Net Cash Flow (US$ Billion) | 31.34 | 36.79 | 43.40
NPV @ 5% Discount Rate (US$ Billion) | 8.20 | 10.26 | 13.22
Internal Rate of Return (%) | 14.9% | 17.5% | 21.9%
Payback Period (Years) | 4.3 | 3.6 | 2.9
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Financial Results | 2020 PEA Comparative Case | 2020 PEA Base Case 3yr Avg. | 2020 PEA Alternate Upside Case
Post-Tax Results: |  |  |
Net Cash Flow (US$ Billion) | 19.58 | 23.03 | 27.22
NPV @ 5% Discount Rate (US$ Billion) | 4.67 | 6.01 | 7.92
Internal Rate of Return (%) | 11.9% | 14.0% | 17.4%
Payback Period (Years) | 4.9 | 4.0 | 3.2

The 2020 PEA offers a viable option for development of KSM and reduces start up risk by focusing on a single mine initially. By including large block cave mines at Iron Cap and Deep Kerr, annual average maximum throughput of 130,000 t/d designed in the 2016 PFS could be increased to 170,000 t/d in the 2020 PEA without significant redesign of facilities. Increased throughput increases the metal production, reduces payback periods, and improves estimated projected IRRs and NPVs.

| 25.4 | 2016 Prefeasibility Study Risks

There are risks that could affect the economic viability of the KSM mine development. Many of these risks are based on the current extent of sufficiently detailed information and engineering in specific project areas. These risks can be further managed as more drilling, sampling, testing, design, and engineering are completed in the next study stage.

However, several risks have been reduced through activities completed by Seabridge and its team since 2012. Specifically, the permitting risk has been addressed substantially with the receipt of environmental approvals granted by both the federal and provincial governments in 2014, and the granting of the early-stage construction permits for KSM, including the MTT License of Occupation. It can be effectively stated that Seabridge earned “the social license” for KSM through the successful completion of the environmental review process with the support of the nearby communities through the submission of letters of support and indigenous groups support. On June 17, 2014, Seabridge Gold entered into a comprehensive Benefits Agreement with the Nisga’a Nation in respect of the KSM Property. On June 10, 2019, the Tahltan Nation and Seabridge Gold Inc. announced that the parties reached agreement on the terms of a Co-operation and Benefits Agreement in connection with KSM.

One of the significant unknowns associated with KSM is related to the extent of available geotechnical data for the tunnels. Since several tunnels are on the critical path for construction, potential delays in construction could occur if unforeseen rock mass conditions or groundwater inflows are encountered that cause durations in excess of those anticipated in the preliminary construction schedule.

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These risks are common to most mining projects, many of which can be mitigated with adequate engineering, planning, and pro-active management. Some external risks, such as metal prices, exchange rates, and government legislation are beyond the control of the developer and operator and are difficult to anticipate and mitigate, although in some instances measures for risk reduction have already been included in conservative design such as in selection of economic mining limits. Risk reduction measures are also included in tunnel and access road design, and through the inclusion of early scheduling for items potentially subject to risk of delay. The means to address risk for a project the size of KSM moving forward is to establish a formal risk management program during advanced study phases that continues through development and into mine operation. The KSM project team will systematically review risks and opportunities during project development and construction, and take appropriate action to minimize the impact on overall costs and scheduling.

| 25.4.1 | Open Pit Mining

The Mitchell pit is a very large open pit with exceptional slope heights. The Sulphurets and Kerr pits have slope heights typical of other open pit metal mines in BC and other parts of the world. The main mining risks for the open pits relate to slope stability, safety, and production delays. Throughout the planning, design, and engineering of the open pits, the KSM project team has identified hazards and developed mitigation plans to reduce and manage risks related to the open pit slopes. Previous studies included a comprehensive risk analysis of the open pit mine design and operating issues. Geotechnical and hydrogeological drilling and test work, geohazard studies, and weather and snow studies have been completed. The results of these have been incorporated into the mine designs for the 2016 PFS.

The following issues were considered in the 2016 PFS:

| ● | slope deformation or rock falls from unstable slopes requiring restricted access to the pit, slope rehabilitation, and lost or reduced mine production. These risks are mitigated via geotechnical design and assessment prior to mining, a comprehensive slope monitoring and management plan during mining, the addition of extra-wide benches at regular 150 m intervals in the pit slope configurations, and standard operating procedures that include good wall control practices and mine operations planning

| ● | snow avalanches at the pit crests resulting in restricted access to the pit. These risks are mitigated via an avalanche management plan coordinated with mine operations

| ● | visibility or weather shutdowns resulting in reduced productivity in the open pits. These risks are mitigated via snow handling crews and equipment, planned lost days and the use of ore stockpiles

| ● | slope depressurization to reduce high pore water pressures in the pit slopes that might result in pit slope instability. These risks are mitigated by an extensive slope depressurization plan that includes vertical wells, horizontal drains, and dewatering adits as a multi-layer system to achieve the design depressurization targets.

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Surface water reporting to the pits due to failure or inundation of water management structures. Excessive surface water into the pits may result in localized slope failures, reduced mine productivity or increased pumping costs. These risks have been mitigated by adjusting the mine plan to limit the exposure of haul ramps to potential erosion from surface water sources and adequate geotechnical design of the water management infrastructures.

The risks associated with the open pits of the 2016 PFS are common for many open pit mining projects with large open pits mined in areas of mountainous terrain with high precipitation. The height of the highwalls also needs consideration. To address the height of the pit walls the mining and geotechnical teams have drawn on experience from design work in similar terrain and operations in these conditions and applied it to the designs in this study to reduce the risk through design, and operating practices, and procedures integrated into the plan.

| 25.4.2 | Underground Mining

A conservative approach was adopted for the design of the critical aspects of the Mitchell and Iron Cap block caves that might impact meeting the production and mill feed schedules. Examples of this include the rate at which the construction of the drawbells is accomplished, the production ramp-up curve and maximum rate of draw of individual drawpoints, the number of drawpoints available for production at any one time, and the travel distance for the LHD vehicles. Design assumptions for all of these are within demonstrated industry experience and are considered achievable. The development advance rate during the critical pre-production period, and subsequently in the production period, have been demonstrated to be industry achievable, and with the technology advances that are currently being developed, may be demonstrated to be conservative in the future. As further conservatism, technological advances with regard to automated and battery powered equipment have not been incorporated in the 2016 PFS design. The adoption of these more advanced technologies may lead to operational efficiencies and reduced costs in the future as they are studied further and adopted into the design.

In developing the production and grade schedules using GEOVIA PCBC™ software, it has been assumed that an infinite supply of zero grade waste rock is available above the columns being drawn. This and other factors controlling the estimates of ore dilution using GEOVIA PCBC™ have been chosen conservatively.

There are recognized uncertainties in predicting the fragmentation of the caved rock at the drawpoints, and the relatively low-fracture intensity of the rock mass at Mitchell, and to a lesser extent Iron Cap, led to the estimated fragmentation being assessed in some detail. Other block caving mines such as Palabora Mine have demonstrated that with careful planning and the availability of equipment to deal with oversize rock, drawpoint production rates comparable to those proposed at Mitchell and Iron Cap are achievable. Furthermore, if the ore pass and grizzly systems do not achieve the planned production rate, and the material transported to the passes is coarser than expected, mitigation measures can be introduced such as employing additional mobile breakage equipment and redesigning the undercut blasting to enhance the fragmentation during the early stages of the column draw.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

As an additional measure to mitigate against possible adverse fragmentation issues at the drawpoints, the designs for Mitchell and Iron Cap incorporate pre-conditioning of the rock mass by hydraulic fracturing, which was not applied at Palabora Mine. This is expected to enhance the fragmentation of the material reporting to individual drawpoints. It also mitigates against possible concerns regarding adverse impacts from stress-induced rock bursts at Mitchell, although the proposed depth of mining and the geotechnical characteristics of the rock mass do not indicate that this will likely pose a major issue. The stress conditions at Iron Cap are estimated to be relatively benign because the deposit is located above the valley floor in a relatively stress-free zone.

The geotechnical characterization of the Mitchell deposit is based on a number of well-positioned geotechnical boreholes that have been geotechnically logged in detail and tested hydrogeologically. Some geotechnical holes have been drilled in the general vicinity of the proposed access development and these holes provide adequate indications of the quality of the rock mass to minimize major uncertainties. Further mitigation of additional uncertainties will be achieved by undertaking further geotechnical drilling in the future.

The geotechnical characterization of the Iron Cap deposit has been based on several holes that were drilled for geotechnical purposes, but reliance was also placed on comparisons of the geological, lithological, and alteration characteristics between the Mitchell and Iron Cap deposits, and developing geotechnical correlations between the two. The correlations were shown to be quite favourable, but further confirmation will be required by undertaking additional geotechnical drilling in the future.

The spacing and layout of the drawpoints of 15 m by 15 m was conservatively selected to achieve favourable interaction between draw columns, minimize early entry of waste dilution material, and to maintain good draw control. Future studies may indicate that this conservatism is more than is required, and an increased drawpoint spacing and layout can possibly be adopted with associated enhanced project economics.

Industry standard approaches have been adopted to support the drawpoints and to maintain their stability. If future analyses of abutment stresses and the interaction of the progressive advance of the undercut on the underlying drawpoints indicate more adverse stress conditions than currently estimated, this can be mitigated by installing higher capacity and more resilient support. It can also be mitigated by adopting a stress-shadowing advance undercut approach, instead of the currently proposed concurrent undercut approach.

The maximum height of draw (HOD) assumed in the design, which controls the maximum tonnage that is drawn at an individual drawpoint, has been limited to 500 m. Based on favourable experience at a number of block cave mines operating in good quality rock, the trend in the industry under these circumstances is to plan to increase the maximum HOD to beyond 500 m. However, the more conservative 500 m limit has been adopted for the Mitchell and Iron Cap designs. There is still the potential requirement for ongoing rehabilitation of drawpoints and mucking drives as a result of changing stress conditions and the adverse impact of secondary blasting. An estimate of the cost of this rehabilitation has been made and this has been included in the block cave operating costs.

Seabridge Gold Inc. | 25-14 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

The extent of surface disturbance from caving and the associated formation of the crater and more peripheral surface cracking have been estimated conservatively for the relatively good geotechnical quality of the rock mass at Mitchell and Iron Cap. Estimates have also been made of the influence of this disturbance on the stability of the walls of the Mitchell open pit and adjacent valley walls. Based on the analyses to date, all permanent infrastructure is beyond the potential impact zone. Major infrastructure that might be critically impacted has been conservatively located well beyond the potential impact zone. If further assessments and future monitoring indicate a potential impact of some of the less critical infrastructure, there is sufficient space on surface for this to be re-located at modest additional cost. The Iron Cap block cave design does not include infrastructure that might be influenced by the extent and nature of the surface disturbance, other than the surface water runoff that might enter the cave that is discussed further in this section.

The estimate of the inflow of surface runoff from rainfall and snowmelt within the catchment area formed by natural slopes, the open pit (in the case of Mitchell), and the crater and adjacent surface cracking, has been based on a 1-in-200 year event. A conservative approach has been adopted in estimating the runoff coefficient, water retention, and impeding factors which control the intensity of this runoff water reaching the production levels underground. The inflow quantities and rates reaching the underground workings are potentially very large for a 1-in-200 year event.

A water management system has been developed for Mitchell, which does not rely on any temporary storage of water in any of the mine openings otherwise required for the operation of the mine. Instead, dedicated water diversion tunnels with a large dedicated storage capacity have been included in the design, with a dedicated shaft and pumping system to extract the water to surface. If experience in the future indicates that the storage capacity is too small to prevent flooding of the operational part of the mine, the pumping rate storage volume can readily be increased or more water diversion tunnels can be added for extra storage. Also, if there is concern about the capacity and functionality of the partial diversion system beyond the rim of the pit that helps control some of the inflow water, or if there is concern that an even larger event than the 1-in-200 year event might occur, consideration can be given to diverting water underground to temporarily flood areas that are benign and do not house critical underground infrastructure, and/or to installing bulkhead doors that can be closed under extreme conditions to limit flood damage to critical underground infrastructure.

The elevation of the production level at Iron Cap is higher than the Mitchell valley floor. This provides for a much simpler water management system than at Mitchell. This is based on gravity flow of water from the entry points underground to a short shaft that connects to the North Slope Depressurization Tunnel. The water then gravity drains from this tunnel to the WSF for treatment.

Seabridge Gold Inc. | 25-15 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

The development of voids beneath the back of the active cave, and the associated concerns about hazardous conditions developing and air blasts occurring, is mitigated by maintaining good knowledge about the cave profile as cave mining progresses. This knowledge is obtained by ongoing monitoring of the geometry of the caved material using techniques such as microseismic monitoring and seismic tomography, which are important mitigation measures to maintain necessary draw control.

Experience at other block cave mines, where there is more potential for mud rushes to occur because of the increased presence of fines (naturally occurring or generated by the attrition of caved rock as it is drawn down) than those at Mitchell or Iron Cap, has shown that such concerns can be mitigated. The measures that can be adopted include monitoring of the flow of water from individual drawpoints, temporary closure of certain areas that are deemed vulnerable until water flows at drawpoints decrease sufficiently, and the adoption of remote mucking until conditions are deemed to be acceptable to return to full entry.

| 25.4.3 | Tunnels

The 2016 PFS involves approximately 77km of tunneling at start up. This includes both infrastructure and water diversion tunnels to accommodate the challenges of building a mine in mountainous terrain and to reduce project risks. In later stages of the mine operation, the cumulative constructed tunnel length rises to over 100 km. Tunnels provide direct routes for ore transport from the mining areas to the Treaty OPC and divert surface water around mining areas and facilities. Tunnels were assessed as having lower operational risks than alternative surface routes. Surface routes are more susceptible to conflicts with other surface infrastructure and ongoing mine operations, and also present risks from climate and geohazards. The permit covering the MTT route was secured from the BC Government in September 2014.

The PFS level of design for the tunnels in this study is based on reasonable assumptions made from preliminary-level investigations that include geological mapping of tunnel routes, geophysical surveys, drilling, and hydrogeological/geotechnical sampling and testing. Poor ground conditions caused by unforeseen faults, areas of weaker than anticipated rock, and/or higher than assumed differential stresses are issues that could cause major disruptions to tunnel construction and operation. If not accounted for properly in design, tunnel collapse or poor performance issues causing business interruptions could result, thus tunnel design represents a significant risk for mine operations and warrants additional investigation in subsequent design phases. There are many successful tunnel projects, including comparable tunnel projects in the KSM area, such as the nearby Granduc Tunnel, that have addressed similar risks successfully. Higher speed methods for long axis tunnelling are becoming more prevalent in mining and civil applications and tunnelling technologies are developing rapidly. However, the cost of mitigation versus the impact of schedule delays must be considered in a risk analysis. Mitigations already incorporated include: the use of twinned tunnels to provide opportunities to advance alternate headings in problem areas, drilling and geophysical testing from surface, the use of probe and pilot drilling in areas of difficult ground conditions, and obtaining the now in place permitting allowing early tunnelling starts.

A Risk Analysis in future studies will lead to cost-benefit analyses of additional investigations, assessment of risk sensitivities for alternate tunnelling methods, review of available changes in designs or routings, and in some cases alternative overland systems so that costs and risks of the alternatives can be considered in any decisions.

Seabridge Gold Inc. | 25-16 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 25.4.4 | Construction Critical Path

The construction critical path runs through early establishment of pioneering roads along the TCAR and CCAR alignments, to establish access to the MTT portals, and then four years of tunnel excavation to complete the twinned 23 km tunnels. Pioneering road construction will be performed from the middle of Year -6 through Q2 Year -5 and includes multiple headings supported by helicopters to install culverts, bridges, and construction of the pioneering road that will expand to full width before the start of the second construction winter in Year -5.

While considered feasible, the schedule is considered aggressive, and additional measures to mitigate potential impacts completion of these roads may have on the critical path is recommended for action in the next design phase. A notable mitigation to the construction critical path is the fact that road construction permits have been granted to Seabridge, allowing road construction to begin earlier in the KSM development plan.

| 25.4.5 | Metallurgical Performance and Process

Lower than expected metal recoveries and copper grade of the concentrate produced is the principal process risk, particularly through the capital payback period, due to significant deviations in metallurgical performance. Unexpected inferior metallurgical performance would affect the 2016 PFS economics.

Additional testwork is required to better characterize metallurgical response of mineralization from the different ore sources. Further process condition and flowsheet optimization would improve process design to accommodate potential variations in metallurgical performance. Current recoveries and economic returns are estimated based on the process design, mainly developed from the metallurgical test results of the Mitchell samples. Samples from the Kerr, Sulphurets, and Iron Cap deposits were also subjected to metallurgical tests and responded similarly to the Mitchell mineralization.

The 2016 PFS proposes to use energy efficient HPGR as a part of the comminution process. The test results from the Mitchell and Sulphurets samples show that this mineralization is amenable to HPGR treatment, although the Sulphurets samples are more resistant to HPGR treatment as compared to the Mitchell samples. The HPGR circuit performance is more sensitive to mill feed moisture and clay contents than other portions of the comminution circuit. Lower than expected performance from the HPGR comminution circuit due to higher than expected moisture and/or clay content can result in a reduction of mill production that could have an impact on the 2016 PFS economics.

Seabridge Gold Inc. | 25-17 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 25.5 | 2020 Preliminary Economic Assessment Risks

| 25.5.1 | Open Pit Mining

The Mitchell pit is a very large open pit with substantial slope heights. The Sulphurets and Kerr pits have slope heights typical of other open pit metal mines in BC and other parts of the world. The main mining risks for the open pits relate to slope stability, safety, and production delays. Throughout the planning, design, and engineering of the open pits, the KSM project team has identified hazards and developed mitigation plans to reduce and manage risks related to the open pit slopes. The reduced reliance on open pit mining in the PEA plan particularly in the early years. The higher pit wall slopes can be left for mining beyond the PEA plan when more is known about the geotechnical characteristics of the rock and the operating conditions.

| 25.5.2 | Underground Mining

A conservative approach was adopted for the designs of the critical aspects of the Iron Cap and Deep Ker block cave designs that might impact meeting the production and mill feed schedules. Examples of this include the rate at which the construction of the drawbells is accomplished, the production ramp-up curve and maximum rate of draw of individual drawpoints, the number of drawpoints available for production at any one time, and the travel distance for the LHD vehicles. Design assumptions for all of these are within demonstrated industry performance standards and are considered achievable. The development advance rate during the critical pre-production period, and subsequently in the production period, have been demonstrated to be industry achievable, and may be demonstrated to be conservative in the future.

Notwithstanding the levels of conservatism that have been applied, there are inherent risks associated with meeting the production targets of 90 kptd for Iron Cap and 80 kptd for Deep Kerr. This has been given careful consideration, and recognizing that the rates of construction of drawpoints and the maximum draw rate per drawpoint are well within what has been achieved in industry, it is considered that the production rates are achievable. The footprints are large in area, but the tramming distances are typical of industry standards, and the infrastructure layout is considered conducive to efficient movement of machinery and materials. Further efficiency studies will be undertaken during the next level of studies to confirm this.

In developing the production and grade schedules using FF software, it has been assumed that an infinite supply of zero grade waste rock is available above the columns being drawn. This and other factors controlling the estimates of mill feed dilution using have been chosen conservatively.

There are recognized uncertainties in predicting the fragmentation of the caved rock at the drawpoints, particularly with the modest fracture intensity of the rock mass expected at Iron Cap and Deep Ker. Other block caving mines such as Palabora Mine have demonstrated that with careful planning and the availability of equipment to deal with oversize rock, drawpoint production rates comparable to those proposed are achievable. Furthermore, mitigation measures can be introduced such as employing additional mobile breakage equipment and redesigning the undercut blasting to enhance the fragmentation during the early stages of the column draw. As an additional measure, pre-conditioning of the rock mass by hydraulic fracturing has been included in the design.

Seabridge Gold Inc. | 25-18 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Only a limited number of boreholes have been drilled specifically for geotechnical and hydrogeological characterization purposes at Iron Cap only. Some geotechnical information has been obtained from limited logging of the exploration holes, and much reliance has been placed on comparisons of the geological, lithological, and alteration characteristics between the Mitchell rock mass and the Iron Cap and Deep Kerr rock masses, developing geotechnical correlations between the two. The correlations were shown to be quite favourable, but further confirmation will be required by undertaking additional geotechnical drilling in the future.

Industry standard approaches have been adopted to support the drawpoints and to maintain their stability. If future analyses of abutment stresses and the interaction of the progressive advance of the undercut on the underlying drawpoints indicate more adverse stress conditions than currently estimated, this can be mitigated by installing higher capacity and more resilient support. It can also be mitigated by adopting a stress-shadowing advance undercut approach, instead of the currently proposed concurrent undercut approach.

The maximum height of draw (HOD) assumed in the design is 750 m. While this is more than many of the caving mines around the world, there is precedence for achieving even higher HODs than this in rock masses with similar geotechnical characteristic and maintaining the stability and integrity of the drawpoints. It is recognized that there will be a requirement for ongoing rehabilitation of drawpoints and mucking drives as a result of changing stress conditions and the adverse impact of secondary blasting. Accordingly the impact of this on production and costs has been taken into account in the study.

The extent of surface disturbance from caving and the associated formation of the crater and more peripheral surface cracking have been estimated conservatively for the relatively good geotechnical quality of the rock mass at Iron Cap and Deep Kerr. Estimates have also been made of the influence of this disturbance on the stability of the walls of the Mitchell open pit and adjacent valley walls. Based on the analyses to date, all permanent infrastructure is beyond the potential impact zone.

Estimates of the inflow of surface runoff from rainfall and snowmelt within the catchment areas, and the crater and adjacent surface cracking, have been based on a 1-in-200 year event at Iron Cap and Deep Kerr. A conservative approach has been adopted in estimating the runoff coefficient, water retention, and impeding factors which control the intensity of this runoff water reaching the production levels underground. Nevertheless, the available data to predict storm events is limited and the inflow quantities and rates reaching the underground workings are potentially very large. This aspect and potential measures to be adopted to mitigate against adverse impacts on meeting production schedules need to be assessed in future levels of study.

Experience at other block cave mines, where there is more potential for mud rushes to occur because of the increased presence of fines (naturally occurring or generated by the attrition of caved rock as it is drawn down) than those at Iron Cap and Deep Kerr, has demonstrated that that such concerns can be mitigated. The measures that can be adopted include monitoring of the flow of water from individual drawpoints, temporary closure of certain areas that are deemed vulnerable until water flows at drawpoints decrease sufficiently, and the adoption of remote mucking until conditions are deemed to be acceptable to return to full entry. Management plans of this nature will be developed as part of future levels of study.

Seabridge Gold Inc. | 25-19 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| 25.5.3 | Tunnels

Tunnels provide direct routes for mill feed transport from the mining areas to the Treaty OPC and divert surface water around mining areas and facilities. Tunnels have been assessed as having lower operational risks than alternative surface routes. Surface routes are more susceptible to conflicts with other surface activities, and also present risks from climate and geohazards. The permit covering the MTT route was secured from the BC Government in September 2014.

For the 2020 PEA, the geotechnical risk has been mitigated by providing additional concrete lining of the diversion tunnels. Previous designs only included concrete lining in areas with poor rock quality or susceptibility for erosion. Designs for the 2020 PEA incorporate continuous lining for the entire length of the MDT and MTDT tunnels to the flow depth. In addition, an allowance for complete lining is provided for areas with poor rock quality. The systematic use of concrete lining of the diversion tunnels increases reliability, reduces potential for erosional damage and allowed use of single bore tunnels for the MDT and MTDT. The use of tunnels with fully lined wetted surfaces allows steeper gradients with higher flow velocity which provided flexibility to optimize alignments.

| 25.5.4 | Metallurgical Performance and Process

Lower than expected metal recoveries and copper grade due to deviations in metallurgical performance would affect the 2020 PEA economics. The metallurgical performance was estimated based on average test results. Some variations are expected during the operation, especially lower than projected concentrate grade. Although the impurities in the flotation concentrate are expected to be lower than the penalty thresholds that may be applied for by most of the smelters, the levels of impurities in some concentrate produced in some period time may exceed the penalty thresholds, especially for the mill feeds from Iron Cap and Deep Kerr block caves.

The 2020 PEA proposes to use energy efficient HPGR as a part of the comminution process. The design was mainly based on the test results from the Mitchell and Sulphurets samples which show amenable to HPGR treatment. The HPGR circuit performance is more sensitive to mill feed moisture and clay contents than other portions of the comminution circuit. Lower than expected performance from the HPGR comminution circuit due to higher than expected moisture and/or clay content can result in a reduction of mill production that could have an impact on the 2020 PEA economics.

The 2020 PEA proposes two process lines, including front-end comminution circuits, to process Mitchell/Sulphurets mill feed and Iron Cap/Deep Kerr mill feed separately. This will require more complicated mill feed transport and handling management. Due to slightly different process conditions, it is better to process these mill feeds separately. However, blending mill feeds from the four deposits for short durations would not have substantial impacts on the metallurgical performance.

Further process condition and flowsheet optimization could improve process design to accommodate the potential variations in the metallurgical performance.

| 25.5.5 | Marketing Risks

A small number of the minor element assays of the final bulk concentrates from Deep Kerr and Iron Cap deposits indicate that concentrations of arsenic, antimony and mercury may be near or above typical smelter penalty limits. The locally elevated impurity contents can be mitigated by mill feed blending with other deposits.

Seabridge Gold Inc. | 25-20 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

26.0 | Recommendations

26.1 | Introduction

This section outlines areas to investigate for improvements to the two alternative development options for the KSM Property that are presented in this Report. A high-level budgetary estimate for the completion of each recommended item is provided in US dollars.

26.2 | 2016 Prefeasibility Study Recommendations

Should Seabridge decide to advance the 2016 PFS to the next logical phase of study, the QPs that authored the 2016 PFS have the following recommendations to support the next level of study.

26.2.1 | Open Pit Mining and Reserves

During more advanced studies, the following optimization studies are recommended:

| ● | conduct a detailed blasthole drill study in conjunction with drill manufacturers to determine site-specific penetration rates for the rock types from each mining area. The 2016 PFS is based on typical values

| ● | delineate overburden in more detail for pre-production and initial benches of pushbacks. Conduct a detailed haulage simulation study to refine the truck/shovel production performance in the mining schedule for pre-production and the first 5 years of mine operations. Simulations should also investigate opportunities to reduce mining costs (e.g. combine pit and dump design strategies to lower fuel consumption from increased downhill loaded hauls). This optimization could have significant efficiency and cost impacts in mountain mines

| ● | analyze alternative equipment power technologies, such as liquefied natural gas (LNG), battery powered and fuel cell

| ● | evaluate the RSF basal drain design(s) to reduce the volume requirements, which could lead to reduced mining costs and greater flexibility in the timing of waste rock placement

| ● | design in more detail and optimize the pre-stripping plan. Conduct operability studies including short range planning on selected key areas of the proposed mine plan

| ● | refine the startup pit selection and initial cutoff grades to reduce the capital payback time

Seabridge Gold Inc. | 26-1 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| ● | obtain budgetary quotes for the operating cost estimates for an optional leased truck fleet in the early years of the mine schedule (pre-production and initial years of production)

| ● | future pit design work should utilize more detailed parameters such as variable slope angles with wall height.

The total estimated cost for these optimization studies is between US$50,000 to US$75,000.

The open pit mine planning designs will need to be reworked to a feasibility-level of detail. The total estimated cost for this open pit mine planning is between US$200,000 to US$250,000.

Also, close-spaced drilling and a higher resolution Mineral Resource model is recommended for the material in the first several years of open pit operations to test for grade and metallurgical continuity. The estimated cost for the extra drilling, assaying, Mineral Resource modeling, and metallurgical testing, is between US$4 million and US$5 million.

26.2.2 | Pit Slopes

The following geotechnical and hydrogeological work is recommended to support future assessments of the KSM open pit slopes:

| ● | complete a field program of large-scale hydrogeological testing via pumping wells. The results of these tests are required to increase the confidence in the design of the depressurization system for the open pit slopes of the Mitchell, Sulphurets, and Kerr zones. A total of five pumping wells and ten monitoring wells for a total meterage of approximately 3,000 m is required

| ● | complete a field program of geotechnical drilling in the Kerr and Sulphurets zones to support updates to the geotechnical model and slope designs. It is estimated that 10 geotechnical holes with a total meterage of approximately 5,000 m is required between the two zones. A program of televiewer surveys of the drill holes, geotechnical logging of the core and laboratory testing should be undertaken

| ● | update the hydrogeological model for the latest geological model and mine plans. This update is required to improve the reliability of the depressurization quantity estimates

| ● | extend the slope design updates to the east and west walls of the Mitchell open pit. Develop a set of updated slope design recommendations for the whole zone that reflect the updates to the geotechnical model

| ● | update the slope design recommendations for Kerr and Sulphurets open pits based on the results of the recommended geotechnical drilling.

The recommended work is estimated to cost US$3.8 million to US$5.0 million. This estimate includes drilling, drill pads, field support (helicopters, camp), and engineering fees. The range of costs are based on per drilled metre expenditures for previous studies.

Seabridge Gold Inc. | 26-2 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

26.2.3 | Underground Mining and Reserves

The following optimization studies are recommended for underground block cave mining of the Mitchell deposit, beneath the Mitchell open pit, and of the Iron Cap deposit.

Mitchell

| ● | undertake a detailed study of the optimum pit profile to transition from open pit to underground mining

| ● | review available geotechnical logging information within the mineralized rock beneath the pit and if deemed beneficial, geotechnically log additional available core to upgrade assessments of fragmentation, drawpoint hang ups, oversize, and drawpoint layout and spacing

| ● | evaluate the impact of fragmentation, oversize, and hang ups on productivity

| ● | undertake a detailed cost-benefit assessment regarding fragmentation and mitigating rock stress impacts, and evaluate the associated impact on schedules of incorporating the pre-conditioning as presently designed into the overall design and mine plan to confirm it provides a net positive benefit

| ● | undertake a study of the potential to increase the drawpoint layout and spacing from the currently proposed 15 m by 15 m and maintain favourable interaction between draw cones without experiencing adverse dilution and premature reporting of sub-economic rock to the drawpoints

| ● | undertake additional laboratory rock strength tests on core to characterize the rock mass strength of the various lithologies and alteration types and intensities

| ● | undertake computer analyses of the stress conditions controlling the extent to which stress-induced fracturing enhances primary fragmentation in the back of the cave, and the potential to experience associated rock bursting as a result of this stress-induced fracturing

| ● | undertake computer analyses of the abutment and cave front stress concentrations. Assess whether an advanced undercut approach, rather than the presently proposed concurrent undercut approach, needs to be adopted to control the stress concentrations on the drawpoints and maintain adequate drawpoint stability

| ● | undertake additional studies of the progressive formation of the crater and resulting destabilizing influence of this on the overlying pit and natural valley slopes. Conduct more detailed assessments of the overall surface disturbance and the possible influence of this on the surface infrastructure adjacent to the block cave operation

| ● | develop and undertake a geotechnical drilling and logging program to characterize the rock that the peripheral infrastructure will be excavated in, including access ramps, ventilation shafts, water conveyance and storage tunnels, crusher stations, underground excavations to support production activities, etc.

Seabridge Gold Inc. | 26-3 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| ● | undertake a detailed review of the surface water runoff inflows and the underground water management system. Conduct optimization studies to meet conveyance and temporary storage requirements to minimize flooding of the mine and pump the water from the mine in the most cost effective manner.

| ● | undertake detailed GEOVIA PCBC™ scheduling studies to optimize caving front and drawpoint sequencing and draw control requirements

| ● | conduct mine productivity simulations to confirm operability and identify potential restraints and restrictions in achieving the proposed production schedules

| ● | develop feasibility-level mine designs and designs of underground infrastructure including crushers, conveyor systems, shops, and ventilation

| ● | assess opportunities to incorporate mine automation and battery powered equipment into the design to reduce ventilation demand and operating costs.

Iron Cap

| ● | similar studies to those recommended for Mitchell are required for Iron Cap as well, although in a number of cases the results of the assessments for Mitchell can be applied to Iron Cap on a comparative basis, with the geotechnical characteristics of the Mitchell and Iron Cap rock masses being somewhat similar. Studies are not required of a transition from open pit to underground mining, and the underground water management system at Iron Cap does not require temporary storage and pumping

| ● | the current production schedule includes production from Iron Cap starting in Year 32, and there are no transition restraints with other production sources in bringing Iron Cap into production. Based on this, the level of study required for Iron Cap is less than for Mitchell to undertake relevant economic assessments for a feasibility-level study

| ● | notwithstanding this, the most significant gap in advancing Iron Cap to a feasibility level is that no geotechnical drilling and logging has been undertaken specifically directed to characterizing the response of the rock mass to block cave mining. In the current study, this has been done primarily by comparison and inference with the Mitchell deposit, and limited geotechnical drilling for pit stability assessments in the mineralized rock. It is recommended that a limited program of geotechnical drilling appropriate for block cave mining assessments be undertaken at Iron Cap to advance the studies to a feasibility level

| ● | this feasibility-level underground work for both Mitchell and Iron Cap, including a limited geotechnical drilling program at Iron Cap, is projected to cost between US$2.5 million to US$3.0 million.

Seabridge Gold Inc. | 26-4 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

26.2.4 | Rock Storage Facilities

KCB recommends that a geotechnical site investigation program be carried out for further delineation and characterization of foundation soils in Mitchell Valley, within the footprint of the Mitchell RSF, with a focus on the area of a known deposit of weak soils near the MTT muck piles. The characterization would use Standard Penetration (SPT) testing and undisturbed sampling for collection of samples for laboratory strength testing. This program will include:

| ● | geotechnical drillholes in the area adjacent to the MTT muck piles and in the RSF toe above the WSF pond area. The holes should be advanced up to 60 m in overburden, with Shelby tube samples and SPT tests. Mobilization of a suitable rig for this program can be combined with drilling in the TMF area for borrow and foundation assessment

| ● | test pits to assess surface soil conditions may be used to augment drilling (for early-stage investigations these will require helicopter supported equipment mobilization)

| ● | geotechnical testing program including consolidation, triaxial, direct shear, and index testing.

Cost of this drilling and a subsequent geotechnical laboratory program is US$600,000 to US$650,000.

A stability assessment and design review is required to assess the findings of the site investigations and to review suitability of existing designs to mitigate the presence of this known weak soil layer. Cost of the assessment and design review is US$75,000 to US$85,000.

26.2.5 | Metallurgical Testing and Process Engineering

In order to optimize process conditions, establish design-related parameters and more accurate metallurgical performance projections for the next stage of study, further metallurgical testwork is recommended, especially the locked cycle flotation tests and cyanidation tests on various ore composite samples from the Sulphurets, Kerr, and Iron Cap deposits. Tetra Tech makes the following recommendations:

| ● | additional metallurgical test work and mineralogical evaluations should be conducted to optimize process conditions and to establish design-related parameters for the next stage of study. The test work should include variability testing of samples from all deposits, especially from the Sulphurets, Kerr, and Iron Cap zones. The variability tests should be a part of the geo-metallurgical testing program that is recommended for the PFS to better understand metallurgical responses of the mineralization in these resources. The testwork should further investigate the effect of low slurry solid density on copper and gold in bulk rougher flotation, especially for the Iron Cap and Kerr mineralization. The test program will provide inputs for geological modelling development and mine plan update for better mill feed control

Seabridge Gold Inc. | 26-5 | 219221-01-RPT-001
KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| ● | technically and economically feasible gold recovery processes from the gold bearing products rejected from copper and gold bulk flotation of the Iron Cap and Kerr mineralization, such as first cleaner scavenger tailing and pyrite concentrate, should be further investigated, The processes should include further optimizing cyanide leach conditions, improving extracted gold and silver recovery from cyanide leach solution and alternative gold and silver extraction processes, such as thiosulphate leaching, ultrafine regrinding treatment and bacterial oxidation. Additional value that may contribute from associated copper due to co-dissolution during gold and silver extraction should be investigated

| ● | further study should be conducted to optimize the proposed cyanide recovery and destruction methods

| ● | a metallurgical laboratory test program should be performed on ore composites representing each year of the initial seven years of open pit mine production from Mitchell and Sulphurets according to the updated mining plan

| ● | further study into economical water treatment methods is recommended for water from the CIL pond.

The estimated costs of these programs range from US$3.0 million to US$4.0 million inclusive of sample shipping and short-term storage costs. Costs for sample collection (i.e. drilling) are not included as this recommendation assumes samples would be derived from core stored on site or in off-site warehouses.

26.2.6 | Water Management

Water Storage Facility and Mine Site Water Management

KCB recommends additional hydrogeological and geotechnical site investigations in the WSD footprint area including:

| ● | hydrogeological/geological data gap analysis and review of additional water level and time series piezometer data to interrogate interaction of geological model assumptions with hydrogeological model parameters

| ● | additional geological mapping and geophysical surveying of the WSD footprint area. May include high resolution drone lidar to map detailed topography of the WSD canyon area

| ● | additional hydrogeological/geotechnical site investigations consisting of drilling up to 2,700m across the WSD area, with packer testing, televiewer downhole imaging, and installation of multi-level piezometers equipped with data loggers. Based on the results, review locations and requirements for the drilling of 3 large diameter wells for pumping tests, with associated monitoring holes

| ● | geotechnical testing is included for WSD area drill holes to characterize soils in the WSD footprint, including SPT sampling and geotechnical laboratory testing program. Complete seismic and test pit programs over the WSD footprint area to assess soil conditions

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| ● | installation of thermistor chains within new or existing drill holes in the WSD footprint area to obtain annual rock mass temperature variations with depth to assess effect of temperature on grouting performance

| ● | in the area of the Mitchell Glacier diversion inlets, geotechnical drilling is recommended to assist in feasibility level design of sub-glacial inlets.

Based on drilling and geophysical quotes, cost of these site investigation programs is US$4.0 million to US$4.5 million, including an allowance for drill pads, instrumentation, pumping test costs field support (helicopters, camp), and engineering fees.

The following design studies are recommended to address the IGRB recommendations:

| ● | review of potential for WSD gravity discharge structures

| ● | commencing an ongoing assessment of effects of long term exposure of monzonite rockfill to acid water is recommended, using accelerated testing techniques

| ● | assessment of options for capture of selenium seepage from RSFs and from other sources to develop an optimum strategy for control of selenium

| ● | update of WSD and mine site water balance to incorporate revised hydrological and meteorological parameters including updated estimates of response to extreme events

| ● | update of hydrogeological seepage models for WSD based on updated site investigations and site parameters

| ● | advancement of WSD, diversion tunnel, spillway and mine site water management designs to a feasibility level of detail.

Cost of these design studies and updates is US$2.0 million to US$2.5 million.

Water Balance

Design parameters such as the 50-year, 100-year, and 200-year storm and flood events; the 100-year dry year and 200-year wet year have been used in PFS design for the sizing of diversion ditches and tunnels and for evaluation of water quality in receiving waters. These parameters have to date been developed from several years of onsite precipitation and streamflow gauging, the nearby longer period Eskay Creek station, as well as use of Global Climatic Models (GCMs). While these analyses provide suitable design parameters for the PFS, more comprehensive analyses incorporating long-term records in the wider region surrounding the sites are required for advanced design phases. In general terms, the longer-term (20 to 100 plus years) meteorological and runoff data from regional stations provide more reliable statistical distributions that may then be extrapolated to the various areas of the site by correlating with onsite stations, some of which have now been monitored for over 12 years.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

It is important for the PFS design evolution that the meteorological and runoff database is expanded and the design parameters refined by conducting a regional hydrologic evaluation incorporating precipitation and stream flow monitoring data from a larger area surrounding the mine site. Typically, the steps in such an analysis include:

| ● | selection of long-term meteorological and runoff stations

| ● | single station frequency analyses for the parameters of interest (such as precipitation of durations ranging from hours to a year, annual maximum flood peaks, etc.)

| ● | regional analyses to determine frequency curves applicable to the various areas such as the mine site and the TMF area.

Following completion of this metrological and runoff data base review, evaluation is recommended to be undertaken to update the site-specific water balance that will be used as the basis in future design. Cost to complete this work is US$350,000 to US$450,000.

26.2.7 | TMF Area

KCB recommends additional hydrogeological and geotechnical site investigations in the TMF footprint area consisting of:

| ● | hydrogeological data gap analysis and review of requirements for additional water level and time series piezometer data. Data gap analysis to refine requirements for locations of pumping tests or other investigations

| ● | data gap review of geotechnical (foundation) and geochemical (characterization of borrow and diversion excavations) data requirements.

| ● | seismic surveying and heli-portable auger soil sampling to further delineate borrow resources and characterize construction materials

| ● | geotechnical drilling with hydrogeological testing (packer, multilevel piezometer and data logger installations) at the North, Saddle and Southeast seepage dams. Total of up to; 26 drillholes for foundation and borrow, 11 large diameter wells for pump testing and 7 Cone Penetration Tests are recommended for dam foundation assessment. Test pitting for borrow investigations may require helicopter support for equipment mobilization

| ● | seep mapping, overburden characterization, and additional overburden permeability testing in the CIL Residue Cell area to further inform the next design stages for determination of drain requirements beneath the Saddle Dam and CIL Cell liner.

Based on geophysical and drilling quotes, costs for these site investigation and laboratory testing programs is US$9.0 million to US$11.0 million including an allowance for line cutting, drill pads, instrumentation, pump testing, field support (helicopters, camp), and engineering fees.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

The following design updates are recommended to address remaining IGRB TMF area recommendations:

| ● | update of TMF area hydrogeological and seepage models and review of required seepage mitigations based on updated site investigations

| ● | updates to drainage designs for TMF dam and CIL liner footprint areas based on revised hydrogeological data and models with liner ballast assessment

| ● | review of required TMF dam cross sectional designs required to meet seasonal storage, freeboard related stability, starter dams and constructability requirements relative to cycloning design optimizations, borrow availability and feasibility level design requirements

| ● | review and update of volume/storage requirements for each cell, based on revised CIL Residue cell volumes

| ● | review of CIL Residue Cell subaqueous deposition system, and liner and drain placement plans for constructability/operability and cost optimizations

| ● | update to TMF area water balance with design updates for discharge pipeline and diffuser system

| ● | advancement of TMF area water management designs to FS-level.

Cost of these design updates is US$1.0 million to US$1.5 million.

26.2.8 | Tunnels

Geotechnical

KCB recommends additional geological and geotechnical investigations at the portal locations, where these have not been investigated already, and also where feasible along the tunnel alignments in the PFS design. The recommended investigations consist of:

| ● | data gap analysis for tunnel design parameters including geological, geotechnical and geochemical aspects

| ● | additional mapping and rock sampling to better characterize properties of lithological units at the portals and along the alignments for geotechnical and geochemical assessments

| ● | more detailed mapping of portal areas and targeted geophysical investigations to assess locations of potential structures (e.g., faults, contacts or water bearing structures along tunnel alignments)

| ● | drilling at selected locations along the permitted MTT alignment (for example, undrilled opportunities for readily accessible drilling to MTT elevation along the alignment include 200 m deep, 500 m deep, and 750 m deep locations). Drilling will provide an opportunity to obtain geotechnical and geochemical samples for laboratory testing, to assess in situ stress regimes, to perform televiewer and packer testing and to install piezometers along the route. For the next phase of MTT drilling KCB has identified 10 target locations for a total of up to 4,300 m of drilling

| ● | drilling for diversion tunnels and PFS infrastructure tunnels will primarily focus on shallower holes, with focus on portal and inlet area rock mass characteristics and sampling of lithological units. Several longer holes (up to 800 m) are recommended to test regional fault systems

| ● | geotechnical and geochemical laboratory testing on samples obtained from drilling and mapping programs.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

The programs will inform selection of the optimum tunnelling method, allow refinement of tunnelling risk, better identify diversion tunnel lining requirements, portal locations and portal development designs and assist with determination of appropriate contingencies. The cost of these programs is estimated to range between US$4.0 million and US$6.0 million, including drilling costs, drill pads, field support (helicopters, camp), and engineering fees.

Tunnel Design

An expert-level risk assessment of the critical tunnels is recommended following the tunnel geotechnical data gap analysis to identify aspects of tunnels with high consequence to the construction schedule. This should also be done for water diversion tunnels with high impact to future facilities. If the end results are design alternatives that cannot mitigate high-consequence risk, then significant data must be collected regarding ground conditions and geologic structure. The risk analysis, and any subsequent data collection and studies, should be done prior to FS so that the PFS can be re-evaluated before significant FS work starts. The analysis should include re-evaluation of opportunities and risks with sensitivities of other tunneling methods to unforeseen conditions. Methods reviewed should include high speed drill and blast, use of a single tunnel with an internal dividing wall, or mechanical systems such as a continuous miner.

The results of the opportunity and risk analysis will indicate where, or if the following are needed:

| ● | geophysical studies to identify major structures and geology to estimate possible ground conditions

| ● | drilling of areas along the tunnel alignment to confirm structures and geology indicated in the geophysical survey

| ● | a detailed risk study may also inform if alternate mitigations such as tunnel lining or other measures to improve tunnel stability and capacity can be considered, instead of the redundant water diversion tunnels included in the water management designs at certain times during mine life and beyond

| ● | ongoing analysis of tunneling method and machines is needed, as this is a continuously advancing technology.

Following the completion of the above risk analysis, possible data acquisition programs, and assessment of resultant alternative design concepts, advanced design updates and budgetary quotes for costing of mitigation options may be needed at an estimated cost of US$250,000.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

26.2.9 | MTT Transport System

The following optimization studies are recommended:

| ● | optimize the size and configuration of the trains using dynamic train movement modelling

| ● | investigate alternative rail systems such as tracks, ballast and ties, and pre-manufactured systems etc. to optimize supply and construction cost and schedule impacts

| ● | evaluate the construction cost and construction schedule of the MTT with respect to tunnel dimensions and alignment, in conjunction with the above optimization study of the train consists

| ● | update the analysis of freight, fuel, lime and personnel transport requirements through the MTT to support the mine area operations, especially if the other design changes to the PFS affect these components

| ● | define and coordinate ore transport and freight labour requirements between train, processing plant, and Mitchell OPC

| ● | evaluate the effect of adjusting the tunnel grade for train start up at the loadout and stopping at the unloading station as well a tunnel alignment and revise the designs as appropriate

| ● | engage the train logistics specialists to investigate the possibility of delaying the twinning of the most eastern section of the MTT without adversely affecting total train, ore, freight, and personnel requirements with the objective reducing startup expenditures

| ● | engage several train system suppliers to optimize the train fleet prior to FS-level budgetary quotes. Due to the specific requirements of the system, it may be necessary to pay for these studies.

Train system designs should be updated during more advanced studies. The train system engineering portion to FS-level detail is estimated to cost between US$250,000 to US$350,000.

26.3 | 2020 Preliminary Economic Assessment Recommendations

Should Seabridge decide to advance the 2020 PEA alternative development option for the KSM property to the next stage of evaluation, the QPs that authored the 2020 PEA have the following recommendations to support more advanced studies.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

26.3.1 | Geology

IRON CAP EXPLORATION ADIT

The 2020 PEA described in Section 24.0 indicates that underground mining is the preferred mining method for the Iron Cap deposit. In order to upgrade the confidence category of the Inferred Mineral Resources that are included in the 2020 PEA mine plan, an infill drilling program is recommended to reduce the drill hole spacing to approximately 125 m. Topography, geohazards, geologic structure and the geometry of the mineralized system precludes drilling entirely from the surface. Underground access would provide the best way to achieve optimal drill hole intersections of the steeply dipping mineralized zones, It is recommended that an exploration adit be developed from near the Mitchell OPC and driven northeasterly beneath the Sulphurets Thrust Fault paralleling the hanging wall of the northwesterly plunging Iron Cap deposit. This adit would provide underground drill access for infill drilling the lower part of the 2020 PEA Iron Cap block cave design, allow for bulk sampling, geotechnical data collection and geologic mapping opportunities.

Development of a 3.6km long exploration adit would consist of surface facilities, including a year-round camp, portal and associated platform infrastructure, geosynthetic lined water rock storage facility, temporary water treatment plant and associated retention ponds and access roads connecting the camp to the portal and to infrastructure in between. The estimated total cost for the exploration adit and other facilities ranges between US$115.0 and US$130.0 million.

IRON CAP MINERAL RESOURCE DRILLING

The QP responsible for Mineral Resources recommends that Seabridge conduct an Iron Cap infill drilling program with the goal of upgrading the Inferred Mineral Resources that were used in the 2020 PEA mine plan. Most of the effort should occur from the adit to properly infill the lower elevations of the block cave design, supplemented by surface drilling to infill the near surface portions of the block cave design. It is estimated that approximately 29,500 m of drilling (51 holes) from underground and 2,400 m (5 holes) from surface may provide sufficient confidence to allow for the upgrade of the confidence classification of the Mineral Resources. This estimate is based on achieving a drill hole spacing of approximately 125 m. The estimated cost for this infill drilling campaign ranges between US$18.0 and US$21.0 million.

KERR MINERAL RESOURCE DRILLING

The 2020 PEA described in Section 24.0 indicates that Kerr could be mined using both open pit and underground mining methods. The QP responsible for Mineral Resources recommends that Seabridge conduct a Kerr infill drilling program with the goal of upgrading the confidence category of the Inferred Mineral Resources in the upper lift of the Kerr block cave design used in the 2020 PEA mine plan. The geometry of the mineralized eastern limb of the Kerr deposit within Lift 1 can be drilled from the surface through a combination of individual holes and mother/daughter wedge setups. It is estimated that approximately 16,000 m of drilling (17 holes) may provide sufficient confidence to upgrade the Inferred Mineral Resources to an Indicated Mineral Resource category. This estimate is based on achieving a drill hole spacing of approximately 125 m. The estimated cost for this surface infill drilling campaign ranges between US$9.0 and US$11.0 million.

Infill drilling of Kerr Lift 2 can only effectively be done from underground that will require the construction of the permitted Deep Kerr Exploration Adit. An estimate of infilling required to upgrade the Lift 2 Inferred Resource would require approximately 30,000 to 35,000 m of drilling (40 holes) from underground drill stations. This program would result in a drill hole spacing of approximately 125 and should provide sufficient confidence to allow for the upgrade of the Inferred Mineral Resources in Kerr Lift 2 to an Indicated category. No estimate of cost is provided since this activity would not occur in the next phase of study. Mining of this Mineral Resource would be anticipated to occur in the latter half of a multi-generational mine life.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

26.3.2 | Open Pit Mining

The following optimization studies are recommended for the PEA:

| ● | repeat the LG economic pit analysis with the updated slope design parameters and revise the pit designs as appropriate

| ● | undertake a detailed evaluation of the Mitchell dozer and truck shovel tramming pre-stripping plan including detailed dozing productivities, equipment scheduling and sequencing, and water flow through at the base of the short term RSFs

| ● | optimize the pre-production plan

| ● | study the potential to reduce pre-production rehandle costs by using cast blasting on the upper narrow benches of Mitchell pit

| ● | evaluate in detail the requirements for the pioneering roads, clearing and grubbing, and topsoil removal activities during pre-production

| ● | smooth out the production from Kerr and Sulphurets pits during the transition to underground

| ● | conduct trade off studies on options to reduce the haul fleet for the long Kerr and Sulphurets hauls

| ● | evaluate the potential to backfill waste rock into the depleted Mitchell pit

| ● | evaluate the basal and selenium drain design(s) to reduce the volume requirements, which could lead to reduced mining costs and greater flexibility in the timing of waste rock placement.

| ● | conduct operability studies including short range planning on selected key areas of the proposed mine plan

| ● | refine the startup pit selection and initial cutoff grades to reduce the capital payback time

| ● | analyze alternative mine equipment power technologies, such as liquefied natural gas (LNG), battery powered and fuel cells.

The total estimated cost for these optimization studies is between US$50,000 to US$75,000.

The open pit mine planning designs will need to be reworked to a feasibility level of detail. The total estimated cost for this open pit mine planning is between US$200,000 to US$250,000.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

Also, close-spaced drilling and a higher resolution Mineral Resource model is recommended for the material in the first several years of open pit operations to test for grade and metallurgical continuity. The estimated cost for the extra drilling, assaying, Mineral Resource modeling, and metallurgical testing, is between US$4 million and US$5 million.

Pit Slopes

Recommended activity to support future assessments of the KSM open pit slopes include completing geotechnical and hydrogeological drilling investigations with downhole measurements as necessary, supplemented by laboratory testing of core. This field work would be used to update the geotechnical database, hydrogeological model and geotechnical slope stability models to improve the reliability of slope stability performance and slope depressurization predictions in a Feasibility Study. The recommended work is estimated to cost US$3.8 million to US$5.0 million, inclusive of drilling, drill pads, field support (helicopters, camp), and engineering fees.

26.3.3 | Underground Mining and Geotechnical

The following optimization studies are recommended for underground block cave mining of the Iron Cap and Deep Kerr deposits for the next level of study.

| ● | undertake a geotechnical drilling and data collection program. The most significant gap in advancing Iron Cap and Deep Kerr to the next level of study is that little to no geotechnical drilling and logging has been undertaken specifically directed to characterizing the response of the rock mass to block cave mining for these deposits. In the current study, the geotechnical characterization has been done primarily by comparison and inference with the Mitchell deposit, limited geotechnical drilling for pit stability assessments in the mineralized rock, and limited geotechnical logging of the core from exploration drilling. It is recommended that a program of geotechnical drilling and logging appropriate for block cave mining assessments be undertaken for the Iron Cap and Deep Kerr deposits

| ● | undertake a detailed study of the optimum pit profile at Deep Kerr to transition from open pit to underground mining

| ● | review the geotechnical logging information that is collected within the mineralized rock and geotechnically log additional available core to upgrade assessments of fragmentation, drawpoint hang ups, oversize, and drawpoint layout and spacing

| ● | undertake additional laboratory rock strength tests on core to characterize the rock mass strength of the various lithologies and alteration types and intensities

| ● | evaluate the impact of fragmentation, oversize, and hang ups on productivity

| ● | develop a hydraulic fracturing plan to enhance fragmentation and mitigate rock stress impacts

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| ● | assess the optimum drawpoint layout and spacing to maintain favourable interaction between draw cones without experiencing adverse dilution and premature reporting of sub-economic rock to the drawpoints

| ● | undertake computer analyses of the stress conditions controlling the extent to which stress-induced fracturing enhances primary fragmentation in the back of the caves, and the potential to experience associated rock bursting as a result of this stress-induced fracturing

| ● | undertake computer analyses of the abutment and cave front stress concentrations and analyze the stability and support requirements for the drawpoints and development excavations on each of the production horizons. Assess whether an advanced undercut approach, rather than the presently proposed concurrent undercut approach, needs to be adopted to control the stress concentrations on the drawpoints and maintain adequate drawpoint stability

| ● | undertake additional studies of the progressive formation of the crater and resulting destabilizing influence of this on natural slopes. Conduct more detailed assessments of the overall surface disturbance and the influence of cave disturbance on infrastructure adjacent to the block cave operation

| ● | undertake additional studies to assess surface water inflows and possible mitigative measures to reduce such inflows into the block cave crater and surrounding disturbed rock mass. Undertake a detailed review the underground water management system and conduct optimization studies to meet water conveyance and temporary storage requirements for surface water inflows

| ● | develop and undertake a geotechnical drilling and logging program to characterize the rock that the peripheral infrastructure will be excavated in, including access ramps, ventilation shafts, water conveyance and storage tunnels, crusher stations, underground excavations to support production activities, etc.

| ● | undertake detailed GEOVIA PCBC™ scheduling studies to optimize caving front advance and drawpoint sequencing and draw control requirements

| ● | assess alternative development techniques to improve cycle times, productivity, and unit costs. The current mine designs are based on conventional drill and blast tunneling methods. Alternative techniques could include the use of navigating drill rigs, comprehensive control systems with drill plan handling, auto generated drill plans, and drilling in automatic mode. In addition, cost savings could be achieved by improving cycle times through maximizing the length of drill rounds, precise drilling, optimizing explosives and blasting for smooth contours that minimize overbreak and scaling, and reducing the ground support cycle time.

| ● | conduct mine productivity simulations to confirm operability and identify potential restraints and restrictions in achieving the proposed production schedules

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

| ● | develop designs of underground infrastructure including crushers, conveyor systems, shops, and ventilation

| ● | evaluate the use of electric vehicles for the Deep Kerr Mine to reduce ventilation demand and airway development

| ● | assess the use of battery electric vehicles for all support and mine development equipment to reduce the ventilation demand

| ● | optimize the train haulage level and train operations at Iron Cap to reduce level development

| ● | evaluate the use of battery electric haulage trucks in place of train haulage once such trucks are available in the 65-tonne class

| ● | further evaluate the ventilation design for an all-electric mine including optimizing drift sizes and surface fan specifications.

The cost of these design and optimization studies is estimated between US$8.0 million and US$12.0 million.

26.3.4 | Iron Cap and Deep Kerr Metallurgical Testwork

Wood recommends that Seabridge perform further testwork for the Iron Cap and Deep Kerr mineralization, investigating the response of various zones to both flotation and cyanidation. This will require composite sampling to assess:

| ● | specific energy, pressing force, and abrasion work to confirm suitability of the HPGR approach to comminution. This testwork is to be done through the use of HPGR testing, piston load testing and pilot plant work on samples

| ● | optimization of primary and regrind milling will be necessary. Bond rod and ball mill work index tests together with regrind jar test work will be necessary for the sizing of ball and regrind mills

| ● | flotation optimization which will include rougher flotation optimization, cleaning work and confirmation of performance through rougher and cleaner batch flotation work followed by locked cycle flotation testwork. The effect of low slurry solid density on copper and gold in bulk rougher flotation should be further investigated

| ● | cyanidation amenability including direct cyanidation kinetics, reagent consumption optimization, and retention time optimization. This testwork will involve both direct cyanidation and CIL batch tests. The processes should also investigate dissolved gold and silver recovery from cyanide leach solution and alternative gold and silver extraction processes, such as thiosulphate leaching, ultrafine regrinding treatment and bacterial oxidation. Additional value that may contribute from associated copper due to co-dissolution during gold and silver extraction should be investigated

| ● | cyanide recovery through the SART approach will be tested through a combination of bench scale batch and bench scale continuous work.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

This work would confirm the applicably of the process flowsheet used for processing the Deep Kerr mineralization in the PEA.

Testing can be performed using drill core from past and future work. It is expected that the comminution testwork would involve approximately 120 point samples requiring approximately 1.5 t of material, 3 small scale continuous HPGR tests requiring another 1 t of material, and approximately 4 t of material for the pilot plant level HPGR work. The smaller samples would be produced through diamond drill hole drilling or current core inventory while the large quantities would be produced through bulk underground sampling. Subsequently, comminution samples would undergo flotation and cyanidation testwork, both batch and locked cycle.

This testing would confirm comminution, flotation and cyanidation viability. With confirmation of the process parameters, the testing of domain composites and point samples would address geometallurgical variability to a PFS level. The overall cost of the program using both available and new sample collection from underground, including for assaying and mineralogical work would be between US$3.0 million and US$4.0 million to provide information to the PFS level for Iron Cap and Deep Kerr materials. This cost may be lower if drilling is also performed in conjunction with other requirements, but it may be also higher if metallurgical issues are indicated by new testing.

26.3.5 | Rock Storage Facilities

KCB recommends that a geotechnical site investigation program be carried out for further delineation and characterization of foundation soils in Mitchell Valley, accompanied by stability assessment and design reviews with particular attention to compressible soils. Cost of this program and associated engineering assessments is between US$750,000 and US $825,000.

26.3.6 | Water Management

KCB recommends additional hydrogeological and geotechnical site investigations in the WSD footprint area including up to 2700 m of hydrogeological/ geotechnical drilling and up to three large diameter pumping tests. Cost of this program range from US$4.0 million to US$4.5 Million.

Several of the IGRB recommendations for design updates to the WSD and mine site water management have been addressed within design updates included in the following technical reports:

| ● | design of WSD grouting gallery, within an initial concrete valley plug structure at base of WSD (KCB, 2020b)

| ● | completion of a preliminary geotechnical and geochemical testing program (KCB 2019). KCB has assessed the use of Mitchell Monzonites for WSD rockfill and has found that the Mitchell Monzonites can replace the Sulphurets Monzonites resulting in a shorter haul distance.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

KCB also recommends WSD design studies and design updates following up on various IGRB recommendations and design improvements, hydrology/climate and water balance assessments and laboratory testing on rockfill. The cost of these studies is from US$2.0 million to US$2.5 million.

Changes in the PEA RSF layout have reduced contact and non-contact water catchment sizes at closure. Additional reduction in contact water treatment for the PEA configuration could be achieved by optimizing alignments and stagings of diversions and routings during construction and operations. Design updates of the closure channels to optimize routings and further examine potential reductions in closure costs is recommended. The total cost of engineering and hydrology studies and design updates to complete this optimization for the PEA configuration is estimated to range between US$50,000 and US$100,000.

26.3.7 | TMF Area

KCB recommends additional hydrogeological, geotechnical and geophysical site investigations in the TMF footprint area. These include up to 26 drillholes, 11 pump tests and 7 Cone Penetration Tests for dam foundation assessments along with a testing program for borrow materials. The cost of these programs is between US$9.0 million and US$11.0 million.

Design updates are also recommended. The costs of these updated are US$1.0 million to US$1.5 million.

The PEA TMF configuration requires relocation of the Splitter Dam to accommodate the increased proportion of pyritic tailings.

A program of up to 900 m of geotechnical/hydrogeological drilling including three CPT test sites and a revised foundation and stability assessment would be required to characterize foundation conditions. The total cost for drilling and testing programs and engineering studies to complete this work is estimated to range between US$0.8 and US$1.0 million.

26.3.8 | Tunnels

Updates to the MDT and MTDT alignments for the PEA have resulted in new outlet locations that have not previously been subject to geotechnical investigations. The following programs are recommended to characterize conditions at these locations:

| ● | examine geotechnical conditions in the area of the relocated Mitchell Diversion Tunnel outlet portal with a mapping program and several shallow drillholes and an updated assessment of geotechnical tunneling conditions along the revised alignment. Total cost of this work is between US$300,000 and US$400,000

| ● | examine geotechnical conditions in the area of the relocated McTagg Diversion Tunnel outlet portal with a mapping program and several shallow drillholes and an updated assessment of geotechnical tunneling conditions along the revised alignment. Total cost of this work is between US$300,000 and US$400,000.

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report |  |

26.3.9 | Power Supply and Energy Recovery Studies

Brazier recommends the following studies relative to KSM power supply and energy recovery:

| ● | study the installation of an energy recovery turbine on the decant line from the TMF to the diffusers in Treaty Creek. An energy recovery turbine on the decant line would recover significant energy from the decant water flows, even after mine closure

| ● | study adding a second pump turbine in the tailing line upper energy recovery building (that is currently in the design), to recover energy during the initial operating years from the second tailing line running to the Saddle Dam area. The additional turbine would increase the amount of energy recovered from the tailing flow

| ● | study energy recovery on the two tailing lines to the South Cell, after Year 16. This must include a study on the south cell filling schedule. The installation would be very similar to the energy recovery plant for the North Cell for Years 1 to 16, except it would be on both lines. A preliminary line route direct to the South Cell must also be selected as part of this study. The two turbines would provide tailing energy recovery after tailing flow is switched to the South Cell in Year 16

| ● | study processing plant design to facilitate demand control so that loads can be deferred for short periods of time in order to shave the peaks off of the KSM electrical demand. Demand control would allow for a larger average load be supplied while keeping within the current BC Hydro 245 MW contract demand.

The estimated cost of these studies is between US$30,000 and US$40,000.

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| 27.0 | MIReferences

1985 Fisheries Act, RSC. C. F-14. (1996a).

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ALS Canada, Ltd., ALS Metallurgy Kamloops (2013). Preliminary Metallurgical Testing – Kerr Deep and Camp Zone KSM Project – (KM 3735). May 9, 2013.

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ALS Metallurgy (2018). Metallurgical Testing of a Mitchell Bulk Cleaner Scavenger Tailings Sample (KM 5455). June 19, 2018.

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Cloos, M. (2001). Bubbling magma chambers, cupolas, and porphyry copper deposits. International Geology Review, 43(4), 285-311.

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Febbo, G. E., Kennedy, L. A., Savell, M., Creaser, R. A., and Friedman, R. M. (2015). Geology of the Mitchell Au-Cu-Ag-Mo porphyry deposit, northwestern British Columbia, Canada. Geological Fieldwork 2014, British Columbia Ministry of Energy and Mines, British Columbia Geological Survey Paper 2015-1, 59-86.

Febbo, G.E., Friedman, R.M., Kennedy, L.A., and Nelson, J.L. (2019a). U-Pb geochronology of the Mitchell deposit, northwestern British Columbia. British Columbia Ministry of Energy, Mines and Petroleum Resources, British Columbia Geological Survey, GeoFile 2019-03, 8 p.

Febbo, G.E., Kennedy, L.A., Nelson, J.L., Savell, M.J., Campbell, M.E., Creaser, R.A., Friedman, R.M., van Straaten, B.I., and Stein, H.J. (2019b). The evolution and structural modification of the supergiant Mitchell Au-Cu porphyry, northwestern British Columbia. Economic Geology, 114(2), 303-324.

Fowler B., and Wells, R. (1995). The Sulphurets Gold zone, northwestern British Colombia. In Schroeter, T.G., ed., Porphyry deposits of the northwestern Cordillera of North America: Canadian Institute of Mining, Metallurgy and Petroleum, Special Volume 46, PART B - Porphyry Copper (±Au±Mo) Deposits of the Calc-Alkalic Suite, 484-492.

G&T (2007). Preliminary Assessment on Mitchell Zone Samples (KM 1909). June 2007.

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G&T (2009). Metallurgical and Pilot Plant Testing on Samples from the Kerr-Sulphurets-Mitchell (KSM) Project (KM 2344). December 2009.

G&T (2010). Ancillary Testing Kerr Sulphurets-Mitchell-KSM Project (KM 2755). December 2010.

G&T (2010). Bench Scale and Pilot Plant Testing Kerr-Sulphurets-Mitchell Project (KM 2670), August 2010.

G&T (2010). Miscellaneous Metallurgical Testing on Samples From the KSM Project (KM 2535). March 2010.

G&T (2011). Data Reports (KM 2897). February – March 2011.

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G&T (2011). Flotation and Cyanidation Testing – Mitchell Composites Years 0-10/0-20 – (KM 3081). November 28, 2011.

G&T (2011). Preliminary Metallurgical Testing on Samples from the Iron Cap Zone – KSM Project (KM 2748). January 2011.

G&T (2012). Flotation and CIL Cyanidation Testing – Kerr-Sulphurets-Mitchell Project – (KM 3080). January 10, 2012.

G&T (2012). Ore Hardness and Flotation Testing – Kerr-Sulphurets-Mitchell (KSM) Project – (KM 3174). January 6, 2012.

Golder (2011). Block Cave Mine Study. Report Number REP 0516_11. Submitted May17, 2011.

Golder (2012). Pre-Feasibility Block Cave Mine Design – Mitchell Deposit. Prepared for Seabridge Gold Inc. Submitted May 2012.

Gustafson, L.B., and Hunt, J.P. (1975). The porphyry copper deposit at El Salvador, Chile. Economic Geology, v.70, 857-912.

Hazen (2008). Comminution Testing (Project # 10724). February 2008.

IGRB (2016). Review of Water Dam, Water Management, and Tailings Storage Systems, KSM Project, British Columbia, Canada. Rev C.1. April 2016.

KCB (2009). 2008 KSM Site Investigation Report.

KCB (2010). 2009 KSM Site Investigation Report.

KCB (2011). 2010 KSM Site Investigation Report.

KCB (2012), 2012 Engineering Design Update of Tailing Management Facility, December 21, 2012.

KCB (2012a). 2012 Site Investigation Report for the Mine Area.

KCB (2012b). 2012 TMF Site Investigations.

KCB (2012c). Engineering Design Update of Tailing Management Facility.

KCB (2012d). KSM Water Storage Dam - Value Engineering Study Report.

KCB (2013a). 2012 Geotechnical Design of Rock Storage Facilities and Design of Associated Water Management Facilities (Technical Report in Support of 2012 PFS), January 29, 2013.

KCB (2013b). KSM Project. Rock Storage Facilities Design Report. July 8, 2013.

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KCB (2014). KSM Project Mitchell/McTagg RSF Water Collection for Selenium Treatment. March 4, 2014.

KCB (2016a). Best Available Technology (BAT) Study for Tailing Management at the KSM Project. June 22, 2016.

KCB (2016a), 2016 Pre-Feasibility Study Update, Addendum Report, Tailing Management Facility Design – Rev. 1, September 22, 2016

KCB (2016b), 2016 Pre-Feasibility Study Update, Addendum Report – Mine Area Water Management – Rev. 1, September 27, 2016

KCB (2016c), 2016 Review of Regional Climate, 2016 Pre-Feasibility Study Update Addendum Report Tailing Management Facility Design – Rev. 1, September 22, 2016

KCB (2019a), September 2018 Site Visit TMF West Till Borrow Area Sampling and Geotechnical Laboratory Testing - Rev. 1, January 23, 2019.

KCB (2019b), Evaluation of Results from Batch 1 and 2 Monzonite Geochemical and Geotechnical Laboratory Testing Programs, March 20, 2019.

KCB (2020a) KSM 2020 Preliminary Economic Assessment, Review of PEA Mine Area Catchments – Rev. 1, February 12, 2020

KCB (2020b) KSM 2020 Preliminary Economic Assessment, Water and Waste Management - Basis of Quantities – Rev. 3, March 13, 2020

KCB (2020c) KSM 2020 Preliminary Economic Assessment, Diversion Tunnel Modifications, January 28, 2020

Kirkham, R.V. (1963). The geology and mineral deposits in the vicinity of the Mitchell and Sulphurets glaciers, Northwest British Columbia. Unpublished M.Sc. thesis: Vancouver, Canada, University of British Columbia, 142 p.

Kirkham, R.V., and Margolis, J. (1995). Overview of the Sulphurets area, northwestern British Columbia. In Schroeter, T.G., ed., Porphyry deposits of the northwestern Cordillera of North America: Canadian Institute of Mining, Metallurgy and Petroleum, Special Volume 46, PART B - Porphyry Copper (±Au±Mo) Deposits of the Calc-Alkalic Suite, pp. 473-483.

Köeppern Machinery Australia Pty Ltd. (2010). High Pressure Comminution Test Work on Processing of Mitchell Zone Ore. February 2010.

Lechner, M.J. (2007). Mitchell Creek Technical Report, Northern British Columbia, NI 43-101 Technical Report prepared for Seabridge Gold.

Lechner, M.J. (2008a). Kerr-Sulphurets Technical Report, Northern British Columbia, NI 43-101 Technical Report prepared for Seabridge Gold.

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Lechner, M.J. (2008b). Updated Mitchell Creek Technical Report, Northern British Columbia, NI 43-101 Technical Report prepared for Seabridge Gold.

Lechner, M.J. (2009). Updated KSM Mineral Resources, NI 43-101 Technical Report prepared for Seabridge Gold.

Lechner, M.J. (2010). January 2010 Updated KSM Mineral Resources, NI 43-101 Technical Report prepared for Seabridge.

Lechner, M.J. (2011). March 2011 Updated KSM Mineral Resources, NI 43-101 Technical Report prepared for Seabridge.

Lechner, M.J. (2014). Initial Deep Kerr Resource Estimate, British Columbia, Canada, NI 43-101 Technical Report.

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Margolis, J. (1993). Geology and intrusion-related copper-gold mineralization, Sulphurets, British Columbia. Unpublished Ph.D. thesis, Eugene, USA, University of Oregon, 289 p.

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Nelson, J. and Colpron, M. (2007). Tectonics and metallogeny of the British Columbia, Yukon and Alaskan Cordillera, 1.8 Ga to the present. Mineral Deposits of Canada: A Synthesis of Major Deposit-Types, District Metallogeny, the Evolution of Geological Provinces, and Exploration Methods. Edited by W.D. Goodfellow. Geological Association of Canada, Mineral Deposits Division, Special Publication, 5, 755-791.

Nordic Minesteel Technologies, 2016, KSM – MTT Automated Rail Ore Transportation System, May 2016.

Palacios, C., Hérail, G., Townley, B., Maksaev, V., Sepúlveda, F., de Parseval, P., Rivas, P., Lahsen, A. and Parada, M.A. (2001). The composition of gold in the Cerro Casale gold-rich porphyry deposit, Maricunga belt, northern Chile. The Canadian Mineralogist, 39(3), 907-915.

Piteau Associates (1991). Mine Rock and Overburden Piles Investigation and Design Manual: Interim Guidelines. Prepared for the British Columbia Mines Waste Rock Pile Research Committee by Piteau Associates Engineering Ltd. http://www.empr.gov.bc.ca/Mining/Permitting-Reclamation/​Geotech/Documents/MinedRock+OverburdenPiles/MinedRockOverburdenPile_Investigation+DesignManual.pdf (accessed January 2013).

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Placer Dome Inc. (1990). Metallurgical Research Centre, Kerr Project Report No.1. October 1990.

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SGS Minerals Services (2010). An Investigation into the Grindability and Flotation Characteristics of Two Samples From the KSM Deposit. February 2010.

SGS Minerals Services (2010). An Investigation into the Recovery of Cyanide and Detoxification of Leach Tailing From Cyanidation of KSM Project Samples. February 2010.

Sillitoe, R.H. (2010). Porphyry copper systems. Economic geology, 105(1), 3-41.

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Surface Science Western, the University Of Western Ontario (2016). Sub-microscopic Gold Determination of Leach Residues. August 30, 2016.

Tetra Tech (2012). 2012 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study. Document No. 1252880100-REP-R0001-05. Prepared for Seabridge Gold Inc. June 22, 2012.

Tetra Tech (2014). Feasibility Study and Technical Report Update on the Brucejack Project, Stewart, BC. Document No. 1491990100-REP-R0001-01. June 19, 2014.

Tetra Tech (2016). 2016 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update and Preliminary Economic Assessment. Document No. 735-1552880100-REP-R0002-02. Prepared for Seabridge Gold Inc. October 6, 2016.

Wardrop (2008). Kerr-Sulphurets-Mitchell Preliminary Economic Assessment 2008, report written for Seabridge and later filed as NI 43-101 Technical Report, December 19, 2008.

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Wardrop (2011). Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study Update 2011, report written for Seabridge and later filed as NI 43-101 Technical Report, June 15, 2011.

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28.0 Certificates of Qualified Persons

| 28-1 |

Certificate of Qualified Person

I, Hassan Ghaffari, P.Eng., do hereby certify:

| ● | I am a Director of Metallurgy with Tetra Tech Inc. with a business address at Suite 1000, 10th Floor, 885 Dunsmuir Street, Vancouver, BC, V6C 1N5.

| ● | This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report”, with an effective date of April 30, 2020 (the “Technical Report”).

| ● | I am a graduate of the University of Tehran (M.A.Sc., Mining Engineering, 1990) and the University of British Columbia (M.A.Sc., Mineral Process Engineering, 2004).

| ● | I am a member in good standing of the Engineers and Geoscientists British Columbia (#30408).

| ● | My relevant experience includes 29 years of experience in mining and mineral processing plant operation, engineering, project studies and management of various types of mineral processing, including hydrometallurgical mineral processing for porphyry mineral deposits.

| ● | I am a “Qualified Person” for the purposes of National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

| ● | I conducted a personal inspection of the KSM property on September 20, 2014.

| ● | I am responsible for Sections 1.1, 1.2.1, 1.12.5, 1.12.6, 1.14, 1.16, 1.18.1, 2.0, 3.1, 12.3.2 (financial analysis), 18.1, 18.5, 18.6, 18.7, 18.8, 18.9, 18.12.2, 18.13, 18.14, 18.15, 21.1 and 21.2 (all other capital costs except for open pit mining, underground mining, permanent electrical power supply and distribution, energy recovery, NTL contribution and MTDT mini hydro generation station costs), 22.0, 25.1, 25.2.5, 25.2.6, 25.4.4, 26.1 and 27.0 (only references from sections for which I am responsible) of the Technical Report.

| ● | I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

| ● | I have had involvement with the KSM property that is the subject of the Technical Report, in acting as a Qualified Person for the “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study” dated June 22, 2012 and “2016 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update and Preliminary Economic Assessment” dated October 6, 2016.

| ● | I have read NI 43-101 and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

| ● | As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Signed and dated this 30th day of April, 2020

“signed and stamped” |
Hassan Ghaffari, P.Eng. Director of Metallurgy Tetra Tech Inc. |

Certificate of Qualified Person

I, Jianhui (John) Huang, Ph.D., P.Eng., do hereby certify:

| ● | I am a Senior Metallurgist with Tetra Tech Inc. with a business address at Suite 1000, 10th Floor, 885 Dunsmuir Street, Vancouver, British Columbia, V6C 1N5.

| ● | This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report”, with an effective date of April 30, 2020 (the “Technical Report”).

| ● | I am a graduate of North-East University, China (B.Eng., 1982), Beijing General Research Institute for Non-ferrous Metals, China (M.Eng., 1988), and Birmingham University, United Kingdom (Ph.D., 2000).

| ● | I am a member in good standing of the Engineers and Geoscientists British Columbia (#30898).

| ● | My relevant experience includes over 34 years involvement in mineral processing for base metal ores, gold and silver ores, and rare metal ores, and mineral processing plant operation and engineering including hydrometallurgical mineral processing for porphyry mineral deposits.

| ● | I am a “Qualified Person” for purposes of National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

| ● | My most recent personal inspection of the KSM property was from June 21, 2017 to June 22, 2017.

| ● | I am responsible for Sections 1.10, 1.11, 1.15, 13.0, 17.0, 19.0, 21.3 (excluding open pit and underground mining costs), 24.19, 25.2.4, 25.4.5, 26.2.5 and 27.0 (only references from sections for which I am responsible) of the Technical Report.

| ● | I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

| ● | I have had involvement with the KSM property that is the subject of the Technical Report, in acting as a Qualified Person for the “Kerr-Sulphurets-Mitchell Preliminary Economic Assessment 2008”, dated December 22, 2008, the “Kerr-Sulphurets-Mitchell Preliminary Economic Assessment Addendum – 2009” dated September 8, 2009, the “Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study” dated March 31, 2010, the “Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study Update” dated June 15, 2011, “2012 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study” dated June 22, 2012 and “2016 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update and Preliminary Economic Assessment” dated October 6, 2016.

| ● | I have read NI 43-101 and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

| ● | As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Signed and dated this 30th day of April, 2020

“signed and stamped” |
Jianhui (John) Huang, Ph.D., P.Eng. Senior Metallurgist Tetra Tech Inc. |

Certificate of Qualified Person

I, Michael J. Lechner, P.Geo., RPG, CPG, do hereby certify:

| ● | I am an independent consultant and owner of Resource Modeling Inc., an Arizona Corporation, with a business address at 124 Lazy J Drive, PO Box 295, Stites, ID, 83552.

| ● | This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report”, with an effective date of April 30, 2020 (the “Technical Report”).

| ● | I am a graduate of the University of Montana, (B.A. Geology, 1979).

| ● | I am a registered professional geologist in the State of Arizona (#37753), a Certified Professional Geologist with the American Institute of Professional Geologists (#10690), a Registered Member of the Society of Mining, Metallurgy and Exploration (#4124987) and Professional Geoscientist with the Engineers and Geoscientists British Columbia (#155344).

| ● | My relevant experience includes over 40 years of work as an exploration geologist, mine geologist, engineering superintendent, and resource estimator for a variety of precious metal and base metal deposits located around the world.

| ● | I am a “Qualified Person” for purposes of National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

| ● | My most recent personal inspection of the KSM property was from October 5, 2019 to October 7, 2019.

| ● | I am responsible for Sections 1.4, 1.5, 1.6, 1.7, 5.0, 6.0, 7.0, 8.0, 9.0, 10.0, 11.0, 12.1, 12.2, 12.4.1, 14.0, 23.0, 26.3.1, and 27.0 (only references from sections for which I am responsible) of the Technical Report.

| ● | I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

| ● | I have had involvement with the KSM property that is the subject of the Technical Report, in acting as a Qualified Person for the following reports:

| – | “Mitchell Creek Technical Report, Northern British Columbia, NI 43-101 Technical Report”, April 6, 2007

| – | “Kerr-Sulphurets Technical Report, Northern British Columbia, NI 43-101 Technical Report” February 29, 2008

| – | “Updated Mitchell Creek Technical Report, Northern British Columbia” dated March 27, 2008.

| – | “Kerr-Sulphurets-Mitchell Preliminary Economic Assessment 2008”, dated December 22, 2008.

| – | “Kerr-Sulphurets-Mitchell Preliminary Economic Assessment Addendum – 2009” dated September 8, 2009

| – | “January 2010 Updated KSM Mineral Resources” dated January 25, 2010

| – | “Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study” dated March 31, 2010

| – | “Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study Update” dated June 15, 2011

| – | “NI 43-101 Technical Report of Initial Deep Kerr Resource, British Columbia, Canada”, dated March 31, 2014.

| – | “2016 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update and Preliminary Economic Assessment” dated October 6, 2016.

| ● | I have read NI 43-101 and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

| ● | As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Signed and dated this 30th day of April, 2020

“signed and stamped” |
Michael J. Lechner, P.Geo., RPG, CPG President Resource Modeling Inc. |

Certificate of Qualified Person

I, James H. Gray, P.Eng., do hereby certify:

| ● | I am a Mining Engineer with Moose Mountain Technical Services with a business address at #210 1510 2nd Street North, Cranbrook, British Columbia, V1C 3L2.

| ● | This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report”, with an effective date of April 30, 2020 (the “Technical Report”).

| ● | I am a graduate of the University of British Columbia, (Bachelor of Applied Science – Mineral Engineering, 1975).

| ● | I am a member in good standing of Engineers and Geoscientists British Columbia (#11919).

| ● | My relevant experience includes mine operation, supervision, and mining engineering in North America, South America, Australia, Eastern Europe, and Greenland.

| ● | I am a “Qualified Person” for purposes of National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

| ● | My most recent personal inspection of the KSM property was from June 21 to 22, 2017.

| ● | I am responsible for open pit mining for Sections 1.9, 1.12, 1.14, 1.15, 1.17, 1.18, 12.3, 15.0, 16.0, 18.0, 21.0, 24.0, 25.0,and 26.0 as well as MTT and rail systems for Sections 18.0, 21.0, 24.0, 25.0 and 26.0 and infrastructure tunnels for Sections 18.0, 21.0, 24.0, 25.0 and 26.0.

| ● | I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

| ● | I have had previous involvement with the KSM property that is the subject of the Technical Report, in acting as a Qualified Person for the “Kerr-Sulphurets-Mitchell Preliminary Economic Assessment 2008”, dated December 22, 2008, the “Kerr-Sulphurets-Mitchell Preliminary Economic Assessment Addendum – 2009” dated September 8, 2009, the “Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study” dated March 31, 2010, the “Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study Update” dated June 15, 2011, the “2012 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study” dated June 22, 2012 and amended November 11, 2014, and the “2016 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update and Preliminary Economic Assessment” dated October 6, 2016.

| ● | I have read the NI 43-101 and the sections of the Technical Report that I am responsible for have been prepared in compliance with the NI 43-101.

| ● | As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Signed and dated this 30th day of April, 2020

“signed and stamped” |
James H. Gray, P.Eng. Principal Mining Engineer Moose Mountain Technical Services |

Certificate of Qualified Person

I, Ross David Hammett, Ph.D., P.Eng., do hereby certify:

| ● | I am a Senior Engineer and Principal with Golder Associates Ltd. with a business address at Suite 200, 2920 Virtual Way, Vancouver, British Columbia, V5M 0C4.

| ● | This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report”, with an effective date of April 30, 2020 (the “Technical Report”).

| ● | I am a graduate of James Cook University of North Queensland (Ph.D., 1976; M.Eng.Sc., 1972; B.E Civil, 1970).

| ● | I am a member in good standing of the Engineers and Geoscientists British Columbia (# 11020).

| ● | I have 44 years of experience in mining and civil engineering. I have provided consulting services for more than 150 underground mining projects and has provided services related to mine planning, mining method selection, mine design, geotechnical studies, support designs, blasting, backfill, caving mechanics, rock stress control, geohydrology, mine dewatering, mining systems, mining automation, and environmental aspects of mining.

| ● | I am a “Qualified Person” for purposes of National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

| ● | My most recent personal inspection of the KSM property was on September 12, 2017 to September 15, 2017.

| ● | I am responsible for underground mining aspects of Sections 1.9, 1.17.2, 1.17.4, 1.17.5, 12.3.1, 15.3, 15.4, 15.5, 16.1, 16.3, 16.4, 21.2 (underground mining costs), 21.3 (underground mining costs except for underground crushers and conveyors and underground electrical system), 24.16.1, 24.16.3, 24.16.4, 24.16.5, 24.18.3 (except underground crushers and conveyors and underground electrical reticulation system), and 24.18.4 (except underground crushers and conveyors and underground electrical reticulation system), 25.2.3.2, 25.3.1.2, 25.4.2, 25.5.2, 26.2.3, 26.3.3 and 27.0 (only references from sections for which I am responsible) of the Technical Report.

| ● | I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

| ● | I have had involvement with the KSM property that is the subject of the Technical Report, in acting as a Qualified Person for the “2016 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update and Preliminary Economic Assessment” dated October 6, 2016.

| ● | I have read NI 43-101 and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

| ● | As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Signed and dated this 30th day of April, 2020

“signed and stamped” |
Ross D. Hammett Ph.D, P.Eng Senior Engineer & Principal Golder Associates Ltd. |

Certificate of Qualified Person

I, Derek Kinakin, P.Geo., P.G., do hereby certify:

| ● | I am a Senior Engineering Geologist with BGC Engineering Inc. with a business address at 234 St. Paul Street, Kamloops, British Columbia, V2C 6G4.

| ● | This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report”, with an effective date of April 30, 2020 (the “Technical Report”).

| ● | I am a graduate of Simon Fraser University (B.Sc. (Hons), 2002; M.Sc., 2005).

| ● | I am a member in good standing of Engineers and Geoscientists British Columbia (#32720).

| ● | My relevant experience includes 18 years of rock mechanics research, slope stability assessments, and slope designs for open pit mines in Canada, US, and Africa.

| ● | I am a “Qualified Person” for purposes of National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

| ● | My most recent personal inspection of the KSM property was August 13, 2018 to August 17, 2018.

| ● | I am responsible for Section 1.12.1, 16.2.5, 24.16.2 (pit slopes) and 26.2.2 and 27.0 (only references from sections for which I am responsible) of the Technical Report.

| ● | I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

| ● | I have had involvement with the KSM property that is the subject of the Technical Report, in acting as a Qualified Person for the “2016 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update and Preliminary Economic Assessment” dated October 6, 2016.

| ● | I have read NI 43-101 and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

| ● | As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Signed and dated this 30th day of April, 2020

“signed and stamped” |
Derek Kinakin, P.Geo., P.G. Senior Engineering Geologist BGC Engineering Inc. |

Certificate of Qualified Person

I, J. Graham Parkinson, P.Geo., do hereby certify:

| ● | I am a Geoscientist with Klohn Crippen Berger Ltd. with a business address at Suite 200, 2920 Virtual Way, Vancouver, British Columbia, V5M 0C4.

| ● | This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report”, with an effective date of April 30, 2020 (the “Technical Report”).

| ● | I am a graduate of the University of British Columbia, (Bachelor’s Degree – Physics, 1978), and the University of Alberta (Special Certificate – Geophysics, 1984).

| ● | I am a member in good standing of the Engineers and Geoscientists British Columbia (#19008).

| ● | My relevant experience includes 24 years with Klohn Crippen and Klohn Crippen Berger engaged in the evaluation and development of mine waste and water management facilities; involvement in over 20 mine waste facilities; involvement in a number of Environmental Impact Assessments, Baseline Studies, Mine Waste Facility Site Investigations, and the design of several major mine tailings dams; and 12 years of experience in engineering and exploration geophysics for the mining, engineering, and petroleum industries.

| ● | I am a “Qualified Person” for purposes of National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

| ● | My most recent personal inspection of the KSM property was from September 10, 2018 to September 14, 2018 as well as during 2007, 2008, 2009, 2012, 2014 and 2016.

| ● | I am responsible for Section 1.12.2, 1.12.3, 1.12.7 (water management tunnels), 18.2, 24.18.1 (except MTT rail system), 25.2.5.3, 25.3.3, 25.4.3 and 25.5.3 (water management tunnels), 26.2.4, 26.2.6, 26.2.7, 26.2.8, 26.3.5, 26.3.6, 26.3.7, 26.3.8 and 27.0 (only references from sections for which I am responsible) of the Technical Report.

| ● | I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

| ● | I have had involvement with the KSM property that is the subject of the Technical Report, in acting as a Qualified Person for the “Kerr-Sulphurets-Mitchell Preliminary Economic Assessment 2008”, dated December 22, 2008, the “Kerr-Sulphurets-Mitchell Preliminary Economic Assessment Addendum – 2009” dated September 8, 2009, the “Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study” dated March 31, 2010, the “Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study Update” dated June 15, 2011, the “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study” dated June 22, 2012 and the “2016 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update and Preliminary Economic Assessment” dated October 6, 2016.

| ● | I have read NI 43-101 and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

| ● | As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Signed and dated this 30th day of April, 2020

“signed and stamped” |
J. Graham Parkinson, P.Geo. Senior Geoscientist Klohn Crippen Berger Ltd. |

Certificate of Qualified Person

I, Neil Brazier, P.Eng., do hereby certify:

| ● | I am a Principal with WN Brazier Associates Inc. with a business address at #8–3471 Regina Ave., Richmond, BC. V6X 2K8

| ● | This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report”, with an effective date of April 30, 2020 (the “Technical Report”).

| ● | I am a graduate of the University of Saskatchewan (B.Sc. Electrical Engineering, 1969).

| ● | I am a member in good standing of the Engineers and Geoscientists British Columbia (#8337).

| ● | My relevant experience includes engineering, construction supervision, and commissioning of a large number of diesel and combustion turbine power plants, high-voltage transmission lines and substations for mining applications.

| ● | I am a “Qualified Person” for purposes of National Instrument 43-101 (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

| ● | My most recent personal inspection of the KSM property was September 1, 2013 to September 4, 2013.

| ● | I am responsible for Sections 1.12.8, 18.10, 18.11, portions of Section 21.1 and 21.2 related to permanent electrical power supply and distribution, energy recovery, NTL contribution, and MTDT mini hydro generation station costs of the Technical Report, portions of Section 24.18.2, 24.18.3 and 24.18.4 related to permanent electrical power supply and distribution, and 26.3.9 and 27.0 (only references from sections for which I am responsible) of the Technical Report.

| ● | I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

| ● | I have had involvement with the KSM property that is the subject of the Technical Report, in acting as a Qualified Person for the “Kerr-Sulphurets-Mitchell Preliminary Economic Assessment 2008”, dated December 22, 2008, the “Kerr-Sulphurets-Mitchell Preliminary Economic Assessment Addendum – 2009” dated September 8, 2009, the “Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study” dated March 31, 2010, the “Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study Update” dated June 15, 2011, the “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study” dated June 22, 2012 and the “2016 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update and Preliminary Economic Assessment” dated October 6, 2016.

| ● | I have read NI 43-101 Standards of Disclosure for Mineral Projects (NI 43–101) and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

| ● | As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Signed and dated this 30th day of April, 2020

“signed and stamped” |
Neil Brazier, P.Eng. Principal WN Brazier Associates Inc. |

Certificate of Qualified Person

I, Brendon Masson, do hereby certify:

| ● | I am a Senior Project Engineer of Engineering with McElhanney Consulting Services Ltd. with a business address at Suite 100, 402 – 11th Ave SE, Calgary, Alberta T2G 0Y4.

| ● | This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report”, with an effective date of April 30, 2020 (the “Technical Report”).

| ● | I am a graduate of the University of University of New Brunswick (B.Sc., Civil Engineering, 2007).

| ● | I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia (#36610).

| ● | My relevant experience is 15 years of location, survey, design, and construction of roads in the Forestry, Mining, and Oil & Gas sectors.

| ● | I am a “Qualified Person” for purposes of National Instrument 43-101 (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

| ● | My most recent personal inspection of the KSM property was on September 13, 2012.

| ● | I am responsible for Sections 1.12.4, 18.12 (except 18.12.2), 24.18.2.2, 24.18.2.3 and 27 (only references from sections for which I am responsible) the Technical Report.

| ● | I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

| ● | I have no prior involvement with the KSM property that is the subject of the Technical Report.

| ● | I have read NI 43-101 Standards of Disclosure for Mineral Projects (NI 43–101) and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

| ● | As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Signed and dated this 30th day of April, 2020

“signed and stamped” |
Brendon Masson, P.Eng. Senior Project Engineer McElhanney Consulting Services Ltd. |

Certificate of Qualified Person

I, Rolf Schmitt, P.Geo., do hereby certify:

| ● | I am a Technical Director with ERM Consultants Canada Ltd. with a business address at 1500-1111 West Hastings St., Vancouver, British Columbia, V6E 2J3.

| ● | This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report”, with an effective date of April 30, 2020 (the “Technical Report”).

| ● | I graduated with a Master of Science in Geology from the University of Ottawa in 1993. In addition, I have obtained a Master of Science, Regional Planning from the University of British Columbia, 1985, and Honours Bachelor of Science, Geology, from the University of British Columbia in 1977.

| ● | I am a member in good standing of the Engineers and Geoscientists British Columbia (#19824).

| ● | I have worked as a geologist, project manager, technical director and senior policy advisor for a total of 42 years since my graduation from university. Key areas of experience of relevance to the KSM project include: Senior Exploration Geologist, Kidd Creek Mines (northwest BC), Research Exploration Geochemist (Geological Survey of Canada), Senior Land Use Geologist and Mineral Policy Specialist (BC Ministry of Energy, Mines and Petroleum Resources), Technical Director and Project Manager (Rescan and ERM), delivering mine Environmental Assessment Projects in BC, permitting major mines in BC and undertaking mine ESG Due Diligence assignments across Canada and internationally as NI 43-101 QP. I have worked on multiple mine assignments through all phases of life-of-mine as part of owners’ exploration/ engineering/ environmental teams and as a senior strategic advisor.

| ● | I am a “Qualified Person” for purposes of National Instrument 43-101 (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

| ● | My most recent personal inspection of the KSM property was on July 12, 2019.

| ● | I am responsible for Sections 1.13, 20.0, 24.20, 26.2.6 (water balance, temporary water treatment plants, and geochemistry database) and 27.0 (only references from sections for which I am responsible) of the Technical Report.

| ● | I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

| ● | I have no prior involvement with the KSM property that is the subject of the Technical Report.

| ● | I have read NI 43-101 Standards of Disclosure for Mineral Projects (NI 43–101) and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

| ● | As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Signed and dated this 30th day of April, 2020

“signed and stamped” |
Rolf Schmidt, P.Geo. Division Managing Director, Canada ERM Consultants Canada Ltd. |

Certificate of Qualified Person

I, Kirk Hanson, P.E. do hereby certify:

| ● | I, Kirk Hanson, P.E., am employed as a Technical Director Open Pit Mining with Wood Group USA Mining Consulting SLC engineering at the company office at 10876 S river Front Pkwy #250, South Jordan, UT 84095, United States.

| ● | This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report”, with an effective date of April 30, 2020 (the “Technical Report”).

| ● | I am registered as a Professional Engineer in the State of Nevada (#10640) and in the State of Alaska (#12126). I graduated with a B.Sc. degree from Montana Tech of the University of Montana, Butte, Montana in 1989 and from Boise State University, Boise, Idaho with a MBA degree in 2004.

| ● | I have practiced my profession for 31 years. I was Chief Engineer at Barrick’s Goldstrike operation, where I was responsible for all aspects of open-pit mining, mine designs, mine expansions and strategic planning. After earning an MBA in 2004, I was assistant manager of operations and maintenance for the largest road department in Idaho. In 2007, I joined AMEC (now Wood) as a principal mining consultant. Over the past 13 years, I have been the mining lead for multiple scoping, pre-feasibility, and feasibility studies. I have also done financial modelling for multiple mines as part of completing the scoping, pre-feasibility and feasibility studies.

| ● | As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101 Standards of Disclosure for Mineral Projects (NI 43–101) for those sections of the Technical Report that I am responsible for preparing.

| ● | I have not completed a personal inspection of the KSM property.

| ● | I am responsible for Section 1.2.2, 1.8, 1.17.1, 1.17.7, 1.17.8, 1.17.9, 1.18.2, 3.2, 12.3.2 (except financial analysis), 12.4.2, 12.4.3, 24.21, 24.22, 25.2.1 and 25.3.4 of the Technical Report.

| ● | I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

| ● | I have no prior involvement with the KSM property that is the subject of the Technical Report.

| ● | I have read NI 43-101 and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

| ● | As of the effective date of the Technical Report, to the best of my knowledge, information, and belief, the sections of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Signed and dated this 30th day of April, 2020.

“signed and stamped” |
Kirk Hanson, P.E. Technical Director, Open Pit Mining Wood Canada Limited |

Certificate of Qualified Person

I, Alan Keylock, P.Eng. do hereby certify:

| ● | I, Alan Keylock, am employed as a Principal Process Engineer with Wood Canada Limited, at the company office at 111 Dunsmuir St #400, Vancouver, BC V6B 5W3

| ● | This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report”, with an effective date of April 30, 2020 (the “Technical Report”).

| ● | I am registered as a Professional Engineer with Engineers and Geoscientists British Columbia (#14684).

| ● | I graduated from the University of Manchester with a Chemical Engineering degree in 1972.

| ● | My relevant experience includes over 30 years of mineral processing including gold extraction by flotation concentration, gravity separation, carbon-in-leach, intensive cyanide leach and pressure oxidation.

| ● | I am a “Qualified Person” for purposes of National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43–101) for those sections of the Technical Report that I am responsible for preparing.

| ● | I have not completed a personal inspection of the KSM property.

| ● | I am responsible for Sections 1.17.6, 24.17, 25.3.2, 25.5.4, 25.5.5, 26.3.4 and 27 (only references from sections for which I am responsible) of the Technical Report.

| ● | I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

| ● | I have no prior involvement with the KSM property that is the subject of the Technical Report.

| ● | I have read NI 43-101 and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

| ● | As of the effective date of the Technical Report, to the best of my knowledge, information, and belief, the sections of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Signed and dated this 30th day of April, 2020.

“signed and stamped” |
Alan Keylock, P.Eng. Principal Process Engineer Wood Canada Limited |

Certificate of Qualified Person

I, William E. Threlkeld, P.Geo., PG, do hereby certify:

| ● | I am the Senior Vice President, Exploration of Seabridge Gold Inc. with a business address at Suite 400, 106 Front Street East, Toronto, Ontario M5A 1E1

| ● | This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment Update, NI 43-101 Technical Report”, with an effective date of April 30, 2020 (the “Technical Report”).

| ● | I am a graduate of the University of Western Ontario, (M.Sc. Economic Geology, 1982).

| ● | I am a registered professional geoscientist with the Engineers and Geoscientists British Columbia (#155516) and a professional geologist in the State of Washington (#790).

| ● | My relevant experience includes over 40 years of professional experience.

| ● | I am a “Qualified Person” for purposes of National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43–101) for those sections of the Technical Report that I am responsible for preparing.

| ● | My most recent personal inspection of the Property was from September 3, 2019 to September 8, 2019.

| ● | I am responsible for Sections 1.3 and 4.0 of the Technical Report.

| ● | I am not independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

| ● | I have read NI 43-101 and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

| ● | As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Signed and dated this 30th day of April, 2020

“signed and stamped” |
William E. Threlkeld, P.Geo., PG Senior Vice President, Exploration Seabridge Gold Inc |