EX-99.2 3 c324-20180307ex992efa906.htm EX-99.2

Report to:
7961 Shaffer Parkway, Suite 5 | Littleton, CO 80127 Phone: (720) 981-1185
NI 43-101 Technical Report Mt Todd Gold Project 50,000 tpd Preliminary Feasibility Study Northern Territory, Australia DOCUMENT NO.: 114910589-REP-R0001-01
PREPARED BY: | List of qualified persons and Designations
| Rex Clair Bryan, Ph.D.
| Anthony Clark, P.E.
| Thomas L. Dyer, P.E.
| Amy L. Hudson, Ph.D., CPG, REM
| Chris Johns, M.Sc., P.Eng
| Benjamin S. Johnson, P.E.
| Deepak Malhotra, Ph.D.
| Zvonimir Ponos, BE, MIEAust, CPeng, NER
| Guy Roemer, P.E.
| Vicki J. Scharnhorst, P.E., LEED AP
| D. Erik Spiller, QP
| Jessica I. Spriet, P.E.
| Keith Thompson, CPG, PG
350 Indiana Street, Suite 500 | Golden, CO 80401 Phone: 303-217-5700 | Fax: 303-217-5705

Report to:
7961 Shaffer Parkway, Suite 5 | Littleton, CO 80127 Phone: (720) 981-1185
NI 43-101 Technical Report Mt Todd Gold Project 50,000 tpd Preliminary Feasibility Study Northern Territory, Australia DOCUMENT: 114910589-REP-R0001-01
EFFECTIVE DATE: January 24, 2018 ISSUE DATE: March 2, 2018
350 Indiana Street, Suite 500 | Golden, CO 80401 Phone: 303-217-5700 | Fax: 303-217-5705

NI 43-101 Technical Report | Vista Gold Corp
50,000 tpd Preliminary Feasibility Study—Northern Territory, Austrailia | Mt Todd Gold Project

FORWARD-LOOKING STATEMENTS

This Technical Report contains forward-looking statements within the meaning of the U.S. Securities Act of 1933, as amended, and U.S. Securities Exchange Act of 1934, as amended, and forward-looking information within the meaning of Canadian securities laws. All statements, other than statements of historical facts, included in this Technical Report that address activities, events, plans or performance that is expected or anticipated to occur in the future are considered to be forward-looking statements. The forward-looking statements in this Technical Report includes such things as, plans and expectations for continued work on the Project, estimates of reserves and resources, estimates of capital, operating and decommissioning costs, projected project economics, including anticipated production, average cash costs, before and after-tax NPV, IRR, capital requirements and expenditures, gold recovery after-tax payback, operating costs, average tonne per day milling, mining methods procedures, recovery, project design, life of mine, average annual production overtime, commencement of commercial production, timing for construction and commissioning, results of the economic analysis, and other such matters. The material factors and assumptions used to develop the forward-looking statements and forward-looking information contained in this Technical Report include the following: the accuracy of the mineral resource and reserve estimates, and exploration and assay results, the terms and conditions of our agreements with contractors and our approved business plan, the anticipated timing and completion of a feasibility study on the Project; the anticipated receipt of required permits; no change in laws that materially impact mining development or operations of a mining business, the potential occurrence and timing of a production decision; the anticipated gold production at the Project; the life of any mine at the Project; all economic projections relating to the Project, including estimated cash cost, NPV, IRR, and initial capital requirements. When used in this press release, the words "optimistic," "potential," "indicate," "expect," "intend," "plans," "hopes," "believe," "may," "will," "if," "anticipate," and similar expressions are intended to identify forward-looking statements and forward-looking information. These statements involve known and unknown risks, uncertainties and other factors which may cause the actual results, performance or achievements to be materially different from any future results, performance or achievements expressed or implied by such statements. Such factors include, among others, uncertainty of resource estimates, estimates of results based on such resource estimates; risks relating to cost increases for capital and operating costs; risks related to the timing and the ability to obtain the necessary permits, risks of shortages and fluctuating costs of equipment or supplies; risks relating to fluctuations in the price of gold; the inherently hazardous nature of mining-related activities; potential effects on operations of environmental regulations in the countries in which it operates; risks due to legal proceedings; risks relating to political and economic instability in certain countries in which it operates; as well as those factors discussed under the headings "Note Regarding Forward-Looking Statements" and "Risk Factors" in Vista's Annual Report Form 10-K as filed on February 22, 2017 and other documents filed with the U.S. Securities and Exchange Commission and Canadian securities regulatory authorities. There may be other factors that cause results not to be as anticipated, estimated or intended. Except as required by law, Vista and the QPs assume no obligation to publicly update any forward-looking statements or forward-looking information; whether as a result of new information, future events or otherwise.

Cautionary Note to United States Investors

The United States Securities and Exchange Commission ("SEC") limits disclosure for U.S. reporting purposes to mineral deposits that a company can economically and legally extract or produce. This Technical Report uses the terms "proven reserves" and "probable reserves". Reserve estimates contained in this Technical Report are made pursuant to National Instrument (NI) 43-101 standards in Canada and do not represent reserves under SEC Industry Guide 7 standards. Under SEC Industry Guide 7 standards, a "final" or "bankable" feasibility study is required to report reserves, the three-year historical average price is used in any reserve

or cash flow analysis to designate reserves and all necessary permits and government approvals must be filed with the appropriate governmental authority. Additionally, this Technical Report uses the terms "measured resources", "indicated resources", and "measured & indicated resources". We advise U.S. investors that while these terms are Canadian mining terms as defined in accordance with NI 43-101, such terms are not recognized under SEC Industry Guide 7 and normally are not permitted to be used in reports and registration statements filed with the SEC. Mineral resources described in this press release have a great amount of uncertainty as to their economic and legal feasibility. The SEC normally only permits issuers to report mineralization that does not constitute SEC Industry Guide 7 compliant "reserves" as in-place tonnage and grade, without reference to unit measures. The term "contained gold ounces" used in this Technical Report is not permitted under the rules of the SEC. "Inferred resources" have a great amount of uncertainty as to their existence, and great uncertainty as to their economic and legal feasibility. It cannot be assumed that any or all part of an inferred resource will ever be upgraded to a higher category. U.S. Investors are cautioned not to assume that any part or all of mineral deposits in these categories will ever be converted into SEC Industry Guide 7 reserves.

TABLE OF CONTENTS

1.0 | SUMMARY | 1
1.1 | Introduction | 1
1.2 | Location | 2
1.3 | Property Description | 2
1.4 | Geology and Mineralization | 7
1.5 | Mineral Resource Estimate | 7
1.6 | Mineral Reserve Estimates | 9
1.6.1 | Heap Leach Reserve Estimate | 10
1.7 | Mining Methods | 11
1.8 | Metallurgy | 13
1.9 | Mineral Processing | 14
1.10 | Project Infrastructure | 14
1.11 | Market Studies and Contracts | 15
1.11.1 | Markets | 15
1.11.2 | Contracts | 15
1.12 | Social and Environmental Aspects | 16
1.12.1 | Existing Environmental and Social Information | 16
1.12.2 | Social or Community Requirements | 16
1.12.3 | Approvals, Permits and Licenses | 17
1.13 | Capital and Cost Estimates | 17
1.13.1 | Capital Cost Estimates – Base Case | 17
1.13.2 | Operating Cost Estimates – Base Case | 19
1.13.3 | Capital Cost Estimates – Alternate Case | 20
1.13.4 | Operating Cost Estimates — Alternate Case | 22
1.14 | Financial Analysis | 23
1.14.1 | Financial Analysis – Base Case | 23
1.14.2 | Financial Analysis – Alternate Case | 24
1.15 | Conclusions and Recommendations | 25
1.15.1 | Feasibility Study | 25
1.15.2 | Geology and Resources | 25
1.15.3 | Mineral Reserve and Mine Planning | 26
1.15.4 | Mineral Processing | 27
1.15.5 | Infrastructure | 27
1.15.6 | Environmental and Social Impacts | 28
1.15.7 | Results of the Site-wide Water Balance Model | 29
1.15.8 | Groundwater Hydrology and Mine Dewatering | 30
1.15.9 | Process Plant Geotechnical Investigation | 30
1.15.10 | TSF Design | 31
1.15.11 | Process | 31
1.15.12 | Geochemical Analyses | 31
2.0 | INTRODUCTION | 32
2.1 | Background Information | 32
2.2 | Terms of Reference and Purpose of the Report | 33
2.3 | Sources of Information | 34
2.4 | Units of Measure | 34
2.5 | Detailed Personal Inspections | 34
3.0 | RELIANCE ON OTHER EXPERTS | 35
4.0 | PROPERTY DESCRIPTION AND LOCATION | 36
4.1 | Location | 36
4.2 | Property Description | 36
4.3 | Lease and Royalty Structure | 36
4.4 | Risks | 41

5.0 | A CCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY | 42
5.1 | Accessibility | 42
5.2 | Climate and Physiography | 42
5.3 | Local Resources and Infrastructure | 42
5.4 | Topography, Elevation and Vegetation | 42
6.0 | HISTORY | 43
6.1 | History of Previous Exploration | 44
6.2 | Historic Drilling | 46
6.2.1 | Batman Deposit | 46
6.2.2 | Drillhole Density and Orientation | 46
6.2.3 | Quigleys | 47
6.3 | Historic Sampling Method and Approach | 50
6.4 | Historic Sample Preparation, Analysis and Security | 50
6.4.1 | Sample Analysis | 50
6.4.2 | Check Assays | 51
6.4.3 | Security | 51
6.5 | Historic Process Description | 51
6.6 | Technical Problems with Historical Process Flowsheet | 52
6.6.1 | Crushing | 52
6.6.2 | Flotation Circuit | 55
6.6.3 | CIL of Flotation Concentrate and Tailings | 55
7.0 | GEOLOGICAL SETTING AND MINERALIZATION | 56
7.1 | Geological and Structural Setting | 56
7.2 | Local Geology | 57
7.3 | Mineralization | 60
7.3.1 | Batman Deposit | 60
7.3.2 | Quigleys Deposit | 61
8.0 | DEPOSIT TYPES | 63
9.0 | EXPLORATION | 64
9.1 | Golden Eye Target | 66
9.2 | RKD Target | 66
9.3 | Silver Spray Target | 66
9.4 | Snowdrop Target | 67
10.0 | DRILLING | 68
10.1 | Drilling | 68
10.2 | Sampling | 70
11.0 | SAMPLE PREPARATION, ANALYSES AND SECURITY | 71
11.1 | Sample Preparation | 71
11.2 | Sample Analyses | 72
11.3 | Sample Security | 73
12.0 | DATA VERIFICATION | 74
12.1 | Drill Core and Geologic Logs | 74
12.2 | Topography | 74
12.3 | Verification of Analytical Data | 74
12.3.1 | Latest Drilling Data Verification | 79
13.0 | MINERAL PROCESSING AND METALLURGICAL TESTING | 82
13.1 | Summary | 82
13.2 | Historic Metallurgical Test Programs | 82
13.3 | 2017 Metallurgical Testwork | 84
13.3.1 | HPGR Testing at Thyssen-Krupp Industries (TKI) | 84
13.3.2 | Tomra/Outotec Ore Sorting Testwork | 86
13.3.3 | Preparation of Composites for Metallurgical Testwork | 89
13.3.4 | Mineralogical Study | 89
13.3.5 | Head Analyses | 90

13.3.6 | Abrasion Indices | 91
13.3.7 | Bond Ball Mill Work Indices | 92
13.3.8 | Leach Tests | 93
13.3.9 | Cyanide Destruction | 96
13.3.10 | Thickening Tests | 96
13.4 | Process Flowsheet | 97
14.0 | MINERAL RESOURCE ESTIMATES | 100
14.1 | Introduction | 100
14.2 | Geologic Modeling of the Batman Deposit (2017) | 102
14.2.1 | Batman Deposit Density Data | 107
14.2.2 | Grade Capping | 107
14.3 | Batman Block Model Parameters | 107
14.3.1 | Geostatistics of the Batman Deposit | 108
14.4 | Batman Estimation Quality | 120
14.5 | Modeling of the Quigleys Deposit | 123
14.5.1 | Quigleys Exploration Database | 124
14.5.2 | Quigleys Block Model Parameters | 124
14.6 | Existing Heap Leach Gold Resource | 130
14.7 | Relevant Factors Affecting Resource Estimates | 131
15.0 | MINERAL RESERVES | 132
15.1 | Pit Optimization | 132
15.1.1 | Economic Parameters | 132
15.1.2 | Slope Parameters | 133
15.1.3 | Pit-Optimization Results | 135
15.1.4 | Ultimate Pit Limit Selection | 136
15.2 | Pit Designs | 136
15.2.1 | Bench Height | 136
15.2.2 | Pit Design Slopes | 137
15.2.3 | Haulage Roads | 138
15.2.4 | Ultimate Pit | 138
15.2.5 | Pit Phasing | 140
15.3 | Cutoff Grade | 144
15.4 | Dilution | 144
15.5 | Reserves and Resources | 144
15.6 | In-Pit inferred Resources | 147
15.7 | Heap Leach Reserve Estimate | 148
16.0 | MINING METHODS | 149
16.1 | Methods | 149
16.2 | Site Landforms and Impoundments | 149
16.3 | Waste Material Definition | 150
16.4 | Mine-Waste Facilities | 150
16.5 | Mine-Production Schedule | 152
16.6 | Equipment Selection and Productivities | 157
16.7 | Mine Personnel | 161
17.0 | RECOVERY METHODS | 165
17.1 | Process Design Criteria | 165
17.2 | Flow Sheet Development | 166
17.2.1 | Crushing Modeling | 166
17.2.2 | Primary Crusher | 168
17.2.3 | Secondary Crushers | 168
17.2.4 | HPGR | 168
17.2.5 | Ore Sorting | 168
17.2.6 | Grinding Modeling | 168
17.2.7 | Thickener / Leach / CIP Design | 168
17.3 | Description of Process Areas | 169

17.3.1 | Area 3100 – Crushing Circuit Availabilities | 169
17.3.2 | Area 3200 – Coarse Ore Stockpile, Reclaim, HPGR and Ore Sorting | 170
17.3.3 | Area 3300 – Grinding and Classification | 170
17.3.4 | Area 3400 – Pre-Leach Thickening, Leach Conditioning, Leach and CIP | 171
17.3.5 | Area 3500 – Desorption, Goldroom and Carbon Regeneration | 171
17.3.6 | Area 3600 – Detoxification and Tailings | 172
17.3.7 | Area 3700 – Reagents | 172
17.3.8 | Area 3800 – Process Plant Services | 173
17.4 | Process Water | 173
17.4.1 | Process Compressed Air | 173
17.5 | Plant Mobile Equipment | 174
18.0 | PROJECT INFRASTRUCTURE | 175
18.1 | Facility 2000 – Mine | 175
18.1.1 | Area 2300 – Mine Support Facilities | 175
18.1.2 | Area 2400 – Mine Support Services | 177
18.2 | Facility 4000 – Project Services | 178
18.2.1 | Area 4100 – Water Supply | 178
18.2.2 | Area 4200 – Power Supply | 179
18.2.3 | Area 4300 – Communications | 180
18.2.4 | Area 4400 – Tailings Dam | 181
18.2.5 | Area 4500 – Waste Disposal | 182
18.2.6 | Area 4600 – Plant Mobile Equipment | 182
18.3 | Facility 5000 – Project Infrastructure | 183
18.3.1 | Area 5100 – Site Preparation | 183
18.3.2 | Area 5200 – Support Buildings | 183
18.3.3 | Area 5400 – Heavy Lift Cranage | 186
18.3.4 | Area 5600 – Bulk Transport | 186
18.3.5 | Area 5800 – Communications | 186
18.4 | Facility 6000 – Permanent Accommodation | 187
18.4.1 | Area 6100 – Personnel Transport | 187
18.5 | Facility 7000 – Site Establishment and Early Works | 187
18.5.1 | Area 7300 – Construction Camp | 187
18.6 | Facility 8000 – Management, Engineering, EPCM Services | 188
18.6.1 | Area 8100 – EPCM Services | 188
18.6.2 | Area 8200 – External Consultants/Testing | 188
18.6.3 | Area 8300 – Commissioning | 188
18.6.4 | Area 8400 – Owners Engineering/Management | 188
18.6.5 | Area 8800 – License, Fees and Legal Costs | 188
18.6.6 | Area 8900 – Project Insurances | 189
18.7 | Facility 9000 – Preproduction Costs | 189
18.7.1 | Area 9100 – Preproduction Labor | 189
18.7.2 | Area 9200 – Commissioning Expenses | 189
18.7.3 | Area 9300 – Capital Spares | 189
18.7.4 | Area 9400 – Stores and Inventories | 189
18.7.5 | Area 9600 – Working Capital and Finance | 189
18.7.6 | Area 9700 – Escalation and Foreign Currency Exchange | 189
18.7.7 | Area 9800 – Contingency Provision | 190
18.7.8 | Area 9900 – Management Reserve Provision | 190
18.8 | Electric Power Plant | 190
18.8.1 | Generation Option Selection | 191
18.8.2 | Mt Todd Electrical | 192
19.0 | MARKET STUDIES AND CONTRACTS | 195
19.1 | Markets | 195
19.2 | Contracts | 195

20.0 | ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT | 196
20.1 | Environmental Studies | 196
20.1.1 | Flora and Vegetation | 197
20.1.2 | Nationally Threatened Species | 197
20.1.3 | Migratory and / or Marine Species | 197
20.1.4 | National Heritage Places | 197
20.2 | Waste and Tailings Disposal, Site Monitoring and Water Management | 198
20.2.1 | Waste Rock Disposal | 198
20.2.2 | Tailings Disposal | 198
20.2.3 | Site Monitoring | 198
20.2.4 | Water Management | 198
20.3 | Permitting and Authorizations | 200
20.4 | Social or Community Requirements | 201
20.5 | Mine Reclamation and Closure | 202
20.5.1 | Batman Pit | 203
20.5.2 | Waste Rock Dump | 203
20.5.3 | Tailings Disposal Facility | 203
20.5.4 | Processing Plant and Pad Area | 204
20.5.5 | Heap Leach Pad and Pond | 204
20.5.6 | Low Grade Ore Stockpile | 205
20.5.7 | Mine Roads | 205
20.5.8 | Water Storage Ponds | 205
20.5.9 | Low Permeability Borrow Area | 205
20.5.10 | Closure Cost Estimate | 206
21.0 | CAPITAL AND OPERATING COSTS | 207
21.1 | Capital Cost | 207
21.1.1 | Mining (MDA) | 209
21.1.2 | CIP Process and Infrastructure (TTP) | 214
21.1.3 | Power Plant (POWER Engineers) | 223
21.1.4 | Mine Dewatering (Tetra Tech) | 224
21.1.5 | Reclamation and Closure (Tetra Tech) | 224
21.1.6 | Water Treatment Plant (Tetra Tech) | 225
21.1.7 | Raw Water Dam (Tetra Tech) | 225
21.1.8 | Tailings Storage Facilities (Tetra Tech) | 226
21.2 | Operating Costs | 227
21.2.1 | Mining (MDA) | 228
21.2.2 | Mine Dewatering (Tetra Tech) | 232
21.2.3 | CIP Process and G&A (TTP) | 232
21.2.4 | Power Plant (POWER Engineers) | 237
21.2.5 | Water Treatment Plant (Tetra Tech) | 239
21.2.6 | Tailings Storage Facilities (Tetra Tech) | 240
21.2.7 | General & Administrative | 240
22.0 | ECONOMIC ANALYSIS | 241
22.1 | Principal Assumptions | 241
22.2 | LoM Production | 242
22.3 | Capital Costs | 244
22.3.1 | 2000 Mining | 245
22.3.2 | 3000 Process Plant | 246
22.3.3 | 4000 Project Services | 247
22.3.4 | 5000 Project Infrastructure | 248
22.3.5 | 6000 Permanent Accommodation | 248
22.3.6 | 7000 Site Establishment & Early Works | 249
22.3.7 | 8000 Management, Engineering, EPCM Services | 249
22.3.8 | 9000 Pre-Production Costs | 250

22.3.9 | 1000 Asset Sale | 250
22.4 | Operating Costs | 252
22.4.1 | Open Pit Mining | 253
22.4.2 | CIP Process Plant | 254
22.4.3 | Water Treatment Plant | 254
22.4.4 | Tailings | 255
22.4.5 | General & Administrative | 255
22.4.6 | JAAC Royalty | 255
22.4.7 | Refining Costs | 256
22.4.8 | Operating Cost Inputs | 256
22.5 | Economic Results | 269
22.5.1 | Taxes, Royalties | 272
22.5.2 | Sensitivity | 273
23.0 | ADJACENT PROPERTIES | 276
24.0 | OTHER RELEVANT DATA AND INFORMATION | 277
24.1 | Process Plant Geotechnical | 277
24.2 | Water Management | 278
24.2.1 | Site-wide Water Balance | 278
24.2.2 | Wet Infrastructure | 281
24.3 | Geochemistry | 288
24.4 | Surface Water Hydrology | 290
24.5 | Regional Groundwater Model and Mine Dewatering | 290
24.5.1 | Regional and Site Hydrogeology | 291
24.5.2 | Regional Numerical Groundwater Flow Model | 292
24.5.3 | Inflow Estimates | 293
24.5.4 | Mine Dewatering | 293
24.6 | Project Implementation | 297
24.6.1 | Project Implementation Strategy | 297
24.6.2 | Project Organization | 297
24.6.3 | EPCM Management | 300
24.6.4 | Engineering | 300
24.6.5 | EPCM Controls | 301
24.6.6 | Procurement | 302
24.6.7 | Construction Management | 307
24.6.8 | Commissioning | 308
24.6.9 | Temporary Construction Facilities | 310
24.6.10 | Industrial Relations | 310
24.6.11 | Health and Safety | 310
24.6.12 | Environment | 312
24.6.13 | Schedule | 312
24.7 | Alternate Case | 316
24.7.1 | Mineral Resource | 316
24.7.2 | Mining | 316
24.7.3 | Process Facility | 331
24.7.4 | Infrastructure | 332
24.7.5 | Site-wide Water Balance Model | 333
24.7.6 | Capital Costs, Alternate Case | 334
24.7.7 | Operating Costs, Alternate Case | 336
24.7.8 | Economic Results, Alternate Case | 337
25.0 | INTERPRETATION AND CONCLUSIONS | 340
25.1 | Project Risks | 340
25.2 | Geology and Resources | 341
25.3 | Mineral Reserve and Mine Planning | 341
25.4 | Mineral Processing | 342
25.5 | Infrastructure | 342

25.5.1 | Site Preparation | 342
25.5.2 | Support Buildings | 342
25.5.3 | Access Roads Parking and Laydown | 343
25.5.4 | Heavy Lifts | 343
25.5.5 | Bulk Transport | 343
25.5.6 | Communications | 343
25.6 | Environmental and Social Conclusions | 343
25.6.1 | Existing Body of Work | 343
25.6.2 | Environmental Impact Study and Approvals | 343
25.6.3 | Social or Community Impacts | 343
25.7 | Results of the Site-wide Water Balance Model | 344
26.0 | R ECOMMENDATIONS | 345
26.1 | Feasibility Study | 345
26.2 | Resource and Exploration | 345
26.3 | Mining Risks and Opportunities | 345
26.3.1 | Opportunities | 345
26.3.2 | Risks | 345
26.4 | Environmental Studies | 346
26.5 | Site-wide Water Balance | 346
26.6 | Groundwater Hydrology and Mine Dewatering | 347
26.7 | Process Plant Geotechnical Investigation Recommendations | 347
26.7.1 | Grinding | 347
26.7.2 | Leach/CIP | 347
26.7.3 | Stockpile & Reclaim | 348
26.7.4 | Coarse Screening | 348
26.7.5 | HPGR | 348
26.8 | Tailings Facility Design | 348
26.8.1 | TSF Construction Schedule | 348
26.8.2 | Investigation of TSF1 Drainage Features | 348
26.8.3 | Geotechnical Investigation and Assessment | 348
26.8.4 | Waste Rock Testing | 349
26.8.5 | Consolidation/Seepage Modeling | 349
26.8.6 | Water Balance | 349
26.8.7 | TSF Consequence Classification | 349
26.9 | Process Operating Costs | 349
26.10 | Geochemical Analyses | 350
26.11 | Process Parameter Optimization | 350
27.0 | REFERENCES | 351
28.0 | CERTIFICATE OF QUALIFIED PERSON | 356
28.1 | Qualifications of Consultants | 356
28.2 | Table of Responsibility | 357

LIST OF TABLES

Table 1 1: PFS Highlights | 2
Table 1 2: Description of Landforms and Impoundments | 4
Table 1 3: Statement of Mineral Resources Estimates | 8
Table 1 4: Statement of Mineral Reserve Estimate | 10
Table 1 5: Initial Economic Parameters | 11
Table 1 6: WhittleTM Pit Optimization Results – Base Case Using 0.40 g-Au/t Cutoff | 12
Table 1 7: WhittleTM Pit Optimization Results – Alternate Case Using 0.40 g-Au/t Cutoff | 12
Table 1 8: Headline Design Criteria | 14
Table 1 9: Mt Todd Permit Status | 17
Table 1 10: Estimated Capital Cost Summary (US$000s) | 18
Table 1 11: Estimated LoM Operating Costs (US$) | 19
Table 1 12: Estimated Capital Cost Summary, Alternate Case (US$000s) | 21
Table 1 13: Estimated LoM Operating Costs, Alternate Case (US$) | 22
Table 1 14: Estimated Technical-Economic Results (US$000s) | 23
Table 1 15: Estimated Economic Results, Alternate Case (US$000s) | 24
Table 6 1: Heap Leach – Historic Actual Production | 43
Table 6 2: Property History | 45
Table 6 3: Summary of Quigleys Exploration Database | 47
Table 7 1: Geologic Codes and Lithologic Units | 57
Table 9 1: Exploration Sampling | 64
Table 9 2: Exploration Prospects | 65
Table 10 1: Batman Deposit Drillholes Added for Resource Update | 68
Table 11 1: Assay and Preparation Laboratories | 72
Table 13 1: Material Balance for HPGR Tests | 85
Table 13 2: Tomra Sorting Test Results | 87
Table 13 3: Head Analyses of Composite Samples | 90
Table 13 4: Whole Rock Analyses of Composite Samples | 90
Table 13 5: Assayed vs. Projected Head Analyses | 91
Table 13 6: Abrasion Indices for the Various Composite Samples | 91
Table 13 7: Bond Ball Mill Work Indices for Composite Samples | 92
Table 13 8: Bond Ball Mill Work Indices for Ore Sorting Products and Wastes | 92
Table 13 9: Gold Extraction vs. Grind Size for the Four Composites | 93
Table 13 10: Gold Extraction at P80 of 270 mesh (53µm) with Two-stage Grind for the Four Composites | 94
Table 13 11: Effect of Pulp Density and NaCN Concentration on Gold Extraction for Composite No. 1 at P80 of 270 mesh (53µm) with Two-stage Grinding | 94
Table 13 12: Effect of Pulp Density and NaCN Concentration on Gold Extraction for Composite No. 3 at P80 of 270 mesh (53µm) with Two-stage Grinding | 95
Table 13 13: Effect of Pulp Density and NaCN Concentration on Gold Extraction for Composite No. 4 at P80 of 270 mesh (53µm) with Two-stage Grinding | 95
Table 13 14: Cyanide Destruction Test Results | 96
Table 13 15: Unit Area Requirements for Thickener for Composite Samples | 96
Table 14 1: Summary of the Batman, Heap Leach Pad and Quiqleys Deposits | 102
Table 14 2: Summary of Batman Bulk Density Data by Oxidation State | 107
Table 14 3: Block Model* Physical Parameters – Batman Deposit | 107
Table 14 4: Batman Resource Classification Criteria and Variogram | 110
Table 14 5: Batman Deposit Measured and Indicated Gold Resource Estimate | 118
Table 14 6: Batman Deposit Inferred Gold Resource Estimate | 119
Table 14 7: Quigleys Deposit Specific Gravity Data | 123
Table 14 8: Summary of Quigleys Exploration Database | 124
Table 14 9: Block Model Physical Parameters – Quigleys Deposit | 124
Table 14 10: Search Parameters for each Domain | 127
Table 14 11: Search Parameters and Sample Restrictions | 127
Table 14 12: WhittleTM Pit Shell Parameters | 128

Table 14 13: Quigleys Deposit Measured and Indicated Gold Resource Estimate within M&I WhittleTM Shell (August 2017) | 128
Table 14 14: Quigleys Deposit Inferred Gold Resource Estimate within M&I WhittleTM Shell (August 2017) | 129
Table 14 15: Existing Heap Leach Indicated Gold Resource Estimate (May 2013) | 130
Table 15 1: Initial Economic Parameters | 132
Table 15 2: Slope Angles for Pit Optimization | 134
Table 15 3: WhittleTM Pit Optimization Results – Base Case using 0.40 g-Au/t Cutoff | 135
Table 15 4: Pit Design Slope Parameters | 137
Table 15 5: Interim Pit Slope Parameters (Sectors 1 & 2) | 137
Table 15 6: US$1,250 Gold Price Cutoff Grades (g-Au/t) | 144
Table 15 7: Base Case Proven and Probable Reserves by Pit Phase | 146
Table 15 8: Total Batman Project Reserves (Base Case plus Heap Leach) | 146
Table 15 9: In-Pit Inferred Resources Inside Base Case Pits | 147
Table 16 1: Description of Landforms and Impoundments | 149
Table 16 2: Base Case Construction and Reclamation Requirements | 151
Table 16 3: Annual Mine Production Schedule – Base Case | 154
Table 16 4: Annual Stockpile Balance – Base Case | 155
Table 16 5: Annual Ore Delivery to the Mill Crusher – Base Case | 156
Table 16 6: Maximum Loader Productivity Estimate | 158
Table 16 7: Annual Load and Haul Equipment Requirements – Base Case | 160
Table 16 8: Mine Personnel Requirements – Base Case | 163
Table 17 1: Headline Design Criteria | 165
Table 17 2: Mobile Equipment for Process Plant | 174
Table 18 1: 50 ktpd TSF 1 and TSF 2 Parameters | 181
Table 18 2: Mobile Equipment for Process Plant | 182
Table 18 3: Heavy Lift Cranage Requirements | 186
Table 18 4: Power Station Location Budgetary Comparison | 191
Table 20 1: Mt Todd Permit Status | 201
Table 20 2: Reclamation Approach | 202
Table 21 1: Operating Periods | 207
Table 21 2: Estimated Capital Cost Summary (US$000s) | 208
Table 21 3: Estimated Mine Annual Capital Costs (US$000s) – Base Case | 209
Table 21 4: Estimated Mine Light Vehicle Capital (US$) | 213
Table 21 5: Estimated Capital Cost Summary (AUD000s) | 214
Table 21 6: CCE Methodology for Facility 3000 – Process Plant | 215
Table 21 7: Methodology for Other Areas of the Capital Cost Estimate | 216
Table 21 8: Estimated Construction Gang Rate Development (AUD) | 219
Table 21 9: Estimated Power Station Installed Capital Cost Summary (AUD) | 223
Table 21 10: Estimated Mine Dewatering Capital Cost Summary (US$000s) | 224
Table 21 11: Estimated Reclamation Capital Cost Summary (US$000s) | 224
Table 21 12: Estimated Water Treatment Plant Capital Cost Summary (US$000s) | 225
Table 21 13: Estimated Raw Water Dam Capital Cost Summary (US$000s) | 225
Table 21 14: Estimated Tailings Storage Facility Capital Cost Summary (US$000s) | 226
Table 21 15: Estimated LoM Operating Costs (US$) | 227
Table 21 16: Estimated Annual Mine Operating Costs (US$) | 230
Table 21 17: Estimated Plant Operating Costs (@ Steady State) (AUD) | 232
Table 21 18: Estimated Fuel Cost Summary (AUD) | 238
Table 21 19: Estimated Personnel Costs, Power Plant (AUD) | 238
Table 21 20: Estimated Gas Turbine Maintenance Cost Schedule – 70MW (AUD) | 239
Table 22 1: TEM Principal Assumptions | 242
Table 22 2: Estimated Refining Costs (US$) | 242
Table 22 3: LoM Ore Production | 242
Table 22 4: Estimated LoM Capital Costs (US$000s) | 244
Table 22 5: Estimated Mining Costs (US$000s) | 245

Table 22 6: Estimated CIP Process Plant Capital Costs (US$000s) | 246
Table 22 7: Estimated Project Services Capital Costs (US$000s) | 247
Table 22 8: Estimated Project Infrastructure Capital Costs (US$000s) | 248
Table 22 9: Estimated Permanent Accommodation Costs (US$000s) | 248
Table 22 10: Estimated Site Establishment & Early Works (US$000s) | 249
Table 22 11: Estimated Management, Engineering, EPCM Services (US$000s) | 249
Table 22 12: Estimated Pre-Production Costs (US$000s) | 250
Table 22 13: Estimated Asset Sale (US$000s) | 250
Table 22 14: Estimated LoM Operating Costs (US$) | 252
Table 22 15: Estimated Open Pit Operating Costs (US$) | 253
Table 22 16: Estimated CIP Process Plant Operating Costs (US$) | 254
Table 22 17: Estimated Water Treatment Plant Operating Costs (US$) | 254
Table 22 18: Estimated Tailings Operating Costs (US$) | 255
Table 22 19: Estimated G&A Operating Costs (US$) | 255
Table 22 20: Estimated JAAC Royalty Costs (US$) | 255
Table 22 21: Estimated Refining Costs (US$) | 256
Table 22 22: Estimated Labor Rates & Costs (AUD) | 257
Table 22 23: Position & Salary Matrix (AUD) | 265
Table 22 24: Process Reagents (AUD) | 268
Table 22 25: Process Consumables (AUD) | 268
Table 22 26: Technical-Economic Results (US$000s) | 270
Table 22 27: Annual Cash Flow | 271
Table 22 28: Project Sensitivity | 273
Table 22 29: Base Case Sensitivity to Gold Price versus Foreign Exchange Rate (US$:AUD) | 275
Table 22 30 Base Case Sensitivity to Gold Prices versus NPV Discount Rate | 275
Table 24 1: Mean Monthly Precipitation | 280
Table 24 2: Site-specific Trigger Values, Edith River Downstream of WTP Discharge | 282
Table 24 3: Edith River Flow at SW4 (m3/h), February 2013 – September 2017 | 283
Table 24 4: Water Quality Data at Sampling Site SW2, Edith River Upstream of WTP Discharge, January 2015 – April 2017 | 284
Table 24 5: Mt Todd WTP Effluent Goals | 284
Table 24 6: Anticipated Influent Water Quality at the WTP | 285
Table 24 7: Opinion of Probable Capital Costs | 286
Table 24 8: Opinion of Probable Annual Chemical Consumption | 287
Table 24 9: Catchment and Pit Areas, Inflow Volumes, and Dewatering Times for Mine Dewatering Design | 293
Table 24 10: Construction Packages | 303
Table 24 11: Supply Packages | 304
Table 24 12: Supply Packages with Significant Lead Times | 313
Table 24 13: Initial Economic Parameters, Alternate Case | 316
Table 24 14: WhittleTM Pit Optimization Results – Alternate Case Using 0.40 g-Au/t Cutoff | 317
Table 24 15: US$1,250 Calculated Gold Price Cutoff Grades (g-Au/t) | 323
Table 24 16: Alternate Case Proven and Probable Reserves by Phase | 324
Table 24 17: Annual Mine Production Schedule – Alternate Case | 326
Table 24 18: Annual Stockpile Balance – Alternate Case | 327
Table 24 19: Annual Ore Delivery to the Mill Crusher – Alternate Case | 328
Table 24 20: Annual Load and Haul Equipment Requirements – Alternate Case | 330
Table 24 21: Headline Design Criteria | 331
Table 24 22: Estimated Alternate Case Capital Cost Summary (US$000s) | 335
Table 24 23: Estimated Operating Cost Summary, Alternate Case (US$) | 336
Table 24 24: Economic Results, Alternate Case (US$000s) | 337
Table 24 25: Project Sensitivity | 338
Table 24 26: Alternate Case Sensitivity to Gold Price versus Foreign Exchange Rate (US$:AUD) | 339
Table 24 27: Alternate Case Sensitivity to Gold Prices versus NPV Discount Rate | 339
Table 25 1: Project Risks | 340

LIST OF FIGURES

Figure 1‑1: General Project Location Map | 3
Figure 1‑2: Concessions | 5
Figure 1‑3: General Arrangement | 6
Figure 1‑4: Measured & Indicated Resource Estimates Grade Tonnage Curves – Batman Deposit | 9
Figure 1‑5: Mt Todd Flowsheet | 13
Figure 4‑1: General Project Location Map | 38
Figure 4‑2: Concessions | 39
Figure 4‑3: General Arrangement | 40
Figure 6‑1: Drillhole Location Map – Batman and Quigleys Deposits | 49
Figure 6‑2: Plant Process Flowsheet for Project as Designed | 53
Figure 6‑3: Modified Plant Process Flowsheet for Project | 54
Figure 7‑1: General Geologic Map | 59
Figure 7‑2: Concessions | 62
Figure 10‑1: Drillhole Location Map Batman Deposit to VB17-003 | 69
Figure 12‑1: NAL Resplit Analyses | 76
Figure 12‑2: NAL Pulp Repeats | 77
Figure 12‑3: Original Pulp Cross Lab Checks | 78
Figure 12‑4: Scatterplot of Relative Au Value to Certified Standard Reference Material Value | 80
Figure 12‑5: Scatterplots (Log Scale) of Replicates by Drillhole | 81
Figure 13‑1: Protocol for HPGR/Ore Sorting | 85
Figure 13‑2: Conceptual Process Flowsheet for Mt Todd Ore (1/2) | 98
Figure 13‑3: Conceptual Process Flowsheet for Mt Todd Ore (2/2) | 99
Figure 14‑1: Drillhole Location Map Batman & Quigleys Deposits and Heap Leach Pad | 101
Figure 14‑2: Schematic of Codes and Surface Designations (Looking North) | 104
Figure 14‑3: Sectional View of Drillhole Data 8,434,803 mN (Looking North) | 106
Figure 14‑4: 2012 Example Log Variograms of Gold within the Core Complex | 109
Figure 14‑5: 2013 Study – Blocks Kriged Au – Cross-section 8,434,900 mN Looking North, Batman Deposit | 112
Figure 14‑6: 2013 Study – Classified Blocks Measured, Indicated, and Inferred – Cross-section 8,434,900 mN Looking North, Batman Deposit | 113
Figure 14‑7: 2013 Study -- Blocks Kriged Au – Level Plan -100m msl Batman Deposit | 114
Figure 14‑8: 2013 Study – Classified Blocks Measured, Indicated, and Inferred – Level Plan -100m msl Batman Deposit | 115
Figure 14‑9: 2013 Study -- Blocks Kriged Au – Long Section of the Core Complex Looking West | 116
Figure 14‑10: 2013 Study – Classified Blocks Measured, Indicated, and Inferred – Long Section of the Core Complex Looking West | 117
Figure 14‑11: Grade Tonnage Curve of Measured and Indicated Resource for the Batman Deposit | 119
Figure 14‑12: Jackknife Correlation Plot for Measured Blocks | 121
Figure 14‑13: Jackknife Correlation Plot for Inferred Blocks | 122
Figure 14‑14: 3-D Visualization of the Quigleys Deposit Mineralized Zone Positions with Wireframe Codes | 125

Figure 14‑15: Quiqleys Median Indicator Variogram | 126
Figure 14‑16: Measured and Indicated Grade Tonnage Curve for the Quigleys Deposit | 129
Figure 14‑17: Inferred Resource Grade Tonnage Curve for the Quigleys Deposit | 131
Figure 15‑1: Mt Todd Geotechnical Sectors | 134
Figure 15‑2: Measured and Indicated Graph of Whittle TM Results – Base Case using 0.40 g-Au/t Cutoff | 136
Figure 15‑3: Mt Todd Ultimate Pit Design – Base Case (February 12, 2018) | 139
Figure 15‑4: Base Case Phase 1 Design (February 12, 2018) | 141
Figure 15‑5: Base Case Phase 2 Design (February 12, 2018) | 142
Figure 15‑6: Base Case Phase 3 Design (February 12, 2018) | 143
Figure 16‑1: Mine Organizational Chart | 154
Figure 17‑1: Simplified Process Flow Diagram | 154
Figure 18‑1: Power Station Location | 154
Figure 18‑2: Conceptual Electrical Line Diagram | 154
Figure 18‑3: General Plant Arrangement | 154
Figure 22‑1: Project Sensitivity | 154
Figure 24‑1: Open Pit Dewatering System Conceptual Design | 154
Figure 24‑2: Conceptual Layout of Dewatering System | 154
Figure 24‑3: EPCM Stage 1 – Design & Procure. Refer Diagram 1 | 154
Figure 24‑4: EPCM Stage 2 – Construct & Commission. Refer Diagram 2 | 154
Figure 24‑5: Commissioning Phases | 154
Figure 24‑6: EPCM Summary Schedule | 154
Figure 24‑7: Measured and Indicated Graph of Whittle TM Results – Alternate Case Using 0.40 g-Au/t Cutoff | 154
Figure 24‑8: Mt Todd Ultimate Pit Design – Alternate Case (February 12, 2018) | 154
Figure 24‑9: Phase I Pit Design – Alternate Case (February 12, 2018) | 154
Figure 24‑10: Phase II Pit Design – Alternate Case (February 12, 2018) | 154
Figure 24‑11: Project Sensitivity | 154

ACRONYMS, ABBREVIATIONS AND SYMBOLS

" | second (plane angle)
% | percent
‘ | minute (plane angle)
< | less than
> | greater than
° | degree
°C | degrees Celsius
°F | degrees Fahrenheit
µm | microns
3D | three-dimensional
A | ampere
a | annum (year)
ABA | acid base accounting
AD | annual deduction
ADWG | Australian Drinking Water Guidelines
AGR | Australian Gold Reagents Pty. Ltd.
ALS | Australian Laboratory Services
Alternate Case | 33,000 tpd
AN | Ammonium nitrate
ANE | Ammonium nitrate emulsion
ANFO | Ammonium nitrate fuel oil
ANZECC | Australian and New Zealand Environment Conservation Council
ANZMARC | Australian and New Zealand Marketing Academy
AOM | Australian Ores and Minerals Limited
AP | aeration/settling ponds
APW | Aerobic Polishing Wetlands
ARD/ML | acid rock drainage and metal laden leachates
ARMCANZ | Agriculture and Resource Management Council of Australia and New Zealand
AStrk | Along Strike
Au | gold
AUD | dollar (Australian)
Ausenco | Ausenco Limited
B | billion
Base Case | 50,000 tpd Case
BCR | biochemical reactor
BFA | Bench face angle
bgs | below ground surface

BH | Bench height
BKK | Bateman Kinhill and Kilborne
BP | Batman pit
Bt | billion tonnes
BWi | Bond Ball Mill work index
CAPEX | capital expenditure or capital expense
CCE | Capital Cost Estimate
CCI | Chamber of Commerce and Industry
CCTV | closed circuit television
CDN | Canadian dollar
CIL | carbon-in-leach
CIM | Canadian Institute of Mining, Metallurgy, and Petroleum Definition Standards
CIP | carbon-in-pulp
cm | centimeters
cm 2 | square centimeter
cm 3 | cubic centimeter
CoA | chart of accounts
CRD | capital recognition deduction
CV | Construction Verification
CWi | Crusher work index
d | day
d/a | days per year (annum)
D&C | Design and Construct
d/wk | days per week
DDH | Diamond drillhole core
DH | drillhole
dmt | dry metric ton
DO | Dissolved oxygen
DoR | Department of Resources
DRDPIFR | Department of Regional Development, Primary Industry, Fisheries and Resources
DC | Dry Commissioning
DUST | dust suppression
DWi | Drop Weight index
E&I | Electrical and Instrumentation
EEE | eligible exploration expenditure
EFCE | Enhanced Factored Cost Estimate
EHS | Environment, Health and Safety
EIS | Environmental Impact Statement
EL | exploration licenses

EMP | Environmental Management Plan
EPCM | Engineering procurement construction management
EQP | equalization pond
F80 | 80% feed passing size
FIS | Free In Store
FS | Feasibility Study
ft | foot
ft 2 | square foot
ft 3 | cubic foot
ft 3/s | cubic feet per second
g | gram
g/L | grams per liter
g/m 3 | gram per cubic meter
g/t | grams per tonne
G&A | general and administrative
Ga | billion years ago
GCL | geosynthetic clay liner
General Gold | General Gold Resources Pty. Ltd.
GHD | GHD Pty Ltd.
GJ | Gigajoule
gpm | gallons per minute (US)
GR | gross realization
GW | gigawatt
h/a | hours per year
h/d | hours per day
h/wk | hours per week
ha | hectare (10,000 m2)
HAZOP | Hazard and Operability
HCL | Hydrochloric Acid
HHV | Higher Heating Value
HLP | heap leach pad
HNO3 | nitric acid
HPGR | high pressure grinding rolls
HQ | 88.9 mm drill rod (outer diameter)
hr | hour
HSEC | Health, Safety, Environment and Community
HV | Heavy vehicles
HW | hanging wall
Hz | hertz

IBC | Intermediate bulk containers
ICP | Inductively Coupled Plasma Atomic Emission Spectroscopy
ICP-OES | Inductively Coupled Plasma Optical Emission Spectroscopy
in | inch
in 2 | square inch
in 3 | cubic inch
IP | Internet Protocol
IRA | Inner-ramp angles
IRR | Internal Rate of Return
IR | Industrial Relations
IT | Information Technology
ITV | interim trigger values
JAAC | Jawoyn Association Aboriginal Corporation
k | kilo (thousand)
kg | kilogram
kg/h | kilograms per hour
kg/m 2 | kilograms per square meter
kg/m 3 | kilograms per cubic meter
km | kilometer
km/h | kilometers per hour
km 2 | square kilometer
koz | kilo-ounce
kPa | kilopascal
kt | kilotonne
KV | Kriging variance
kV | kilovolts
kVA | kilovolt-ampere
kW | kilowatt
kWh | kilowatt hour
kWh/a | kilowatt hours per year
kWh/t | kilowatt hours per tonne
kW/sec | Kilowatts per second
L | liter
L/m | liters per minute
lb | pound(s)
LGOS | low grade ore stockpile
LIMS | Laboratory information system
LLDPE | linear low-density polyethylene
LoM | life of mine

LPM | low-permeability material
m | meter(s)
M | million
m bgs | meters below ground surface
m/min | meters per minute
m/s | meters per second
m 2 | square meter
m 3 | cubic meter
m 3/hr | cubic meter(s) per hour
MARC | maintenance and repair contract
masl | meters above mean sea level
Mb/s | megabytes per second
Mbm 3 | million bank cubic meters
Mbm 3/a | million bank cubic meters per annum
mbsl | meters below sea level
MCC | Motor Control Center
MDA | Mine Development Associates
mg | milligram
MGA | Map Grid of Australia
mg/L | milligrams per liter
MIF | Measured, Indicated, inferred
min | minute (time)
mL | milliliter
MLN | Mineral License Number
mm | millimeter
MMP | Mining Management Plan
mo | month
Moz | million ounces
Mpa | megapascal
mPa∙s | centipoise
MPU | Mobile processing unit
MRT | Mining & Resource Technology Pty Ltd
Mt | million tonnes
Mt/a | million tonnes per annum
MTO | material take-off
Mtpy | million tonnes per year
MVA | megavolt-ampere
MW | megawatt
MWH | Montgomery Watson Harza (now Stantec)

N/mm 2 | Newtons per square millimeter
NAG | Net Acid Generation
NAL | Northern Australian Laboratories
NaOH | Sodium Hydroxide
NAPP | net acid production potential
NHMRC | National Health and Medical Research Council
NI | National Instrument
NOI | Notice of Intent
NP | neutralization potential
NPI | Non Process Infrastructure
NPR | neutralizing potential ratio
NPV | Net Present Value
NQ | 69.9 mm drill rod (outer diameter)
NRETAS | Natural Resources, Environment, the Arts and Sport
NRMMC | Natural Resource Management Ministerial Council
NT | Northern Territory
NTEL | NT Environmental Laboratories
NTEPA | Northern Territory Environmental Protection Authority
ø | diameter
OC | operating costs
OH&S | Occupational Health and Safety
OP | open rotary holes
OPEX | operating expenditure or operating expense
OPGW | optical ground wire
oz | ounce
oz/a | ounces/annum
oz/d | ounces/day
P80 | 80% product passing size, in micrometers
P&ID | piping and instrumentation diagram
Pa | Pascal
Pacific Gold Mines | Pacific Gold Mines NL
PAG | potentially acid generating
PAH | Pincock Allen and Holt
PbS | galena
PC | Prime Cost
PCG | Pine Creek Geosyncline
pcg | Porphyry copper gold
PER | Public Environmental Report
PFS | Preliminary Feasibility Study

PGM | plant growth medium
POWER | POWER Engineers, Inc.
PP | Process Plant
ppb | parts per billion
ppm | parts per million
Project | Mt Todd Gold Project
PRP | Process Plant Retention Pond
PWC | Power and Water Corporation
PWP | Process Water Pond
QA/QP | Quality Assurance/Quality Control
QP | Qualified Person
R&R | Rest and recreation
RDi | Resource Development Inc.
RKD | RKD (Company Name)
RL | Sample name
RO | runoff pond
RoM | Run of Mine
RP | retention pond
rpm | revolutions per minute
RVC | reverse circulation drilling method
RWD | raw water dam
s | second (time)
SABC | semiautogeneous ball crusher
SAPS | Successive alkalinity producing systems
SG | specific gravity
SMBS | sodium metabisulfite
SMC | SAG mill comminution
SME | Society for Mining, Metallurgy, and Exploration, Inc.
SMP | Structural, Mechanical and Piping
SOCS | Site of Conservation Significance
SoW | Scope of Work
SPX | SPX company name
SRE | Soil and Rock Engineering
st | short ton (2,000 lb)
st/d | short tons per day
st/y | short tons per year
SWWB | Site-wide water balance
t | tonne (1,000 kg) (metric ton)
t/a | tonnes per year

t/d | tonnes per day
t/m 3 | tonnes per cubic meter
Technical Report | this Preliminary Feasibility Study
TEM | technical economic model
Tetra Tech | Tetra Tech, Inc.
TKI | Thyssen-Krupp Industries
tpd | tonnes per day
tph | tonnes per hour
TSF | tailings storage facility
TTP | Coffey Services Australia Pty Ltd (trading as Tetra Tech Proteus)
TUNRA | The University of Newcastle Research Associates
TV | Trigger value
TWC | The Winters Company
UCS | Unconfined compressive strength
US$ | U.S. dollar
V | volt
Vista | Vista Gold Corp.
Vista Australia | Vista Gold Australia Pty Ltd
VoIP | voice over Internet protocol
w/w | weight/weight
WA | Western Australia
WAD | weak acid dissociable
WC | Wet Commissioning
WDL | Waste Discharge License
wk | week
WRD | waste rock dump
WTP | water treatment plant
WWTP | waste water treatment plant
XRD | x-ray diffraction
yd 3 | cubic yard
ZnS | Sphalerite

UNITS OF MEASURE

All dollars are presented in U.S. dollars (US$) unless otherwise noted. Common units of measure and conversion factors used in this report include:

Weight:

1 oz (troy)= 31.1035 g

Analytical Values:

| percent | grams per metric tonne
1% | 1% | 10,000
1 g/t | 0.0001% | 1.0
10 ppb |  |
100 ppm |  |

Linear Measure:

1 inch (in) = 2.54 centimeters (cm)

1 foot (ft) = 0.3048 meters (m)

1 year (yd) = 0.9144 meters (m)

1 mile (mi) = 1.6093 kilometers (km)

Area Measure:

1 acre = 0.4047 hectare

1 square mile = 640 acres = 259 hectares

ABBREVIATIONS OF THE PERIODIC TABLE

actinium = Ac | aluminum = Al | amercium = Am | antimony = Sb | argon = Ar
arsenic = As | astatine = At | barium = Ba | berkelium = Bk | beryllium = Be
bismuth = Bi | bohrium = Bh | boron = B | bromine = Br | cadmium = Cd
calcium = Ca | californium = Cf | carbon = C | cerium = Ce | cesium = Cs
chlorine = Cl | chromium = Cr | cobalt = Co | copper = Cu | curium = Cm
dubnium = Db | dysprosium = Dy | einsteinum = Es | erbium = Er | europium = Eu
fermium = Fm | fluorine = F | francium = Fr | gadolinium = Gd | gallium = Ga
germanium = Ge | gold = Au | hafnium = Hf | hahnium = Hn | helium = He
holmium = Ho | hydrogen = H | indium = In | iodine = I | iridium = Ir
iron = Fe | juliotium = Jl | krypton = Kr | lanthanum = La | lawrencium = Lr
lead = Pb | lithium = Li | lutetium = Lu | magnesium = Mg | manganese = Mn
meltnerium = Mt | mendelevium = Md | mercury = Hg | molybdenum = Mo | neodymium = Nd
neon = Ne | neptunium = Np | nickel = Ni | niobium = Nb | nitrogen = N
nobelium = No | osmium = Os | oxygen = O | palladium = Pd | phosphorus = P
platinum = Pt | plutonium = Pu | polonium = Po | potassium = K | prasodymium = Pr
promethium = Pm | protactinium = Pa | radium = Ra | radon = Rn | rhodium = Rh
rubidium = Rb | ruthenium = Ru | rutherfordium = Rf | rhenium = Re | samarium = Sm
scandium = Sc | selenium = Se | silicon = Si | silver = Ag | sodium = Na
strontium = Sr | sulfur = S | technetium = Tc | tantalum = Ta | tellurium = Te
terbium = Tb | thallium = Tl | thorium = Th | thulium = Tm | tin = Sn
titanium = Ti | tungsten = W | uranium = U | vanadium = V | xenon = Xe
ytterbium = Yb | yttrium = Y | zinc = Zn | zirconium = Zr |

1.0SUMMARY

1.1Introduction

Vista Gold Corp. (Vista) retained Tetra Tech, Inc., along with Mine Development Associates (MDA), Resource Development Inc. (RDi), Tetra Tech Proteus (TTP), and POWER Engineers, Inc. (POWER) to prepare this preliminary feasibility study (PFS) for its Mt Todd Gold Project (the Project) in Northern Territory (NT), Australia. The PFS (Technical Report) evaluates the Base Case, a development scenario of a 50,000 tonne per day (tpd) processing facility. In addition, an Alternate Case was considered at 33,000 tpd with higher grades presented under Section 24.0 – Other Relevant Data and Information.

Vista and its subsidiary, Vista Gold Australia Pty Ltd (Vista Australia) entered into an agreement to acquire an interest in the Project located in NT, Australia on March 1, 2006. The acquisition was completed on June 16, 2006 when the mineral leases comprising the Project were transferred to Vista Australia and funds held in escrow were released. Vista Australia is the operator of the Mt Todd property.

The Mt Todd property contains a number of known occurrences of gold, which have been explored and/or exploited to various degrees. The largest and best-known deposits are the Batman and Quigleys deposits, both of which have had historic mining by prior operators. The Batman deposit has produced and been explored more extensively than the Quigley deposit. Vista has reported mineral resource estimates in accordance with National Instrument (NI) 43-101 Standards of Disclosure for Mineral Projects and Canadian Institute of Mining, Metallurgy and Petroleum (CIM Definition Standards) for Mineral Resources and Mineral Reserves (CIM) Standards for the Batman and Quigley deposits and a mineral reserve estimate in accordance with NI 43-101 and CIM Standards for only the Batman deposit.

The primary purpose of this Technical Report is to provide updated material, scientific, and technical information based on additional data obtained from drilling conducted from 2015 to 2017 and the newly implemented sorting method of removing waste rock from the process circuit; the redesign of the grinding circuit; and describe the Base Case mine plan of 50,000 tpd and an Alternative Case of 33,000 tpd which contains ore mined from the Batman open pit plus the ore from the existing heap leach pad.

The highlights of the PFS are presented in Table 1‑1.

Table 1‑1: PFS Highlights

Description | Years 1-5 | Life of Mine (LoM) (13 years)
Annual Average | Total | Annual Average | Total
Average Milled Grade (g-Au/t) | 0.98 | 0.82
Payable Gold (koz) | 479 | 2,397 | 382 | 4,956
Gold Recovery (%) | 86.4% | 85.8%
Cash Costs (US$/oz) | $571 | $645
Strip Ratio (waste:ore) | 2.83 | 2.52
Initial Capital (US$ millions) | $839
After-tax Payback (Production Years) | 3.2
Pre-tax NPV5% (US$ millions) | $1,178
After-tax NPV5% (US$ millions) | $679
IRR (Pre-tax / After-tax) (%) | 27.3% / 20.5%

NOTE: Economics presented using US$1,300/oz gold and a flat US$0.80:AUD1.00 exchange rate and assumes deferral of certain NT tax obligations as well as realization of equipment salvage values at the end of the mine life.

This Technical Report notes material items which may, or may not, affect the market price of Vista’s securities. Updated information contained herein includes scientific, technical and economic information deemed material to the Project.

1.2Location

The Project is located 56 kilometers (km) by road northwest of Katherine, and approximately 290 km southeast of Darwin in NT, Australia (Figure 1‑1). Access to the property is via high quality, two-lane paved roads from the Stuart Highway, the main arterial within the territory.

1.3Property Description

Vista Australia is the holder of four mineral licenses (ML) MLN 1070, MLN 1071, MLN 1127, and MLN 31525 comprising approximately 5,544 hectares (ha). In addition, Vista Australia controls exploration licenses (EL) EL 29882, EL 29886, EL 30898, and EL 28321 comprising approximately 153,700 ha. Figure 1‑1 illustrates the general location of the tenements and the position of the Batman deposit.

The general arrangement for the Project is shown on Figure 1-3. and landforms and impoundments are described in Table 1‑2.

Figure 1‑1: General Project Location Map

Table 1‑2: Description of Landforms and Impoundments

Landform/Impoundment | Abbreviated Name
Tailings Storage Facility 1 | TSF 1
Tailings Storage Facility 2 | TSF 2
Raw Water Dam | RWD
Low Grade Ore Stockpile | LGOS
Low Grade Ore Stockpile Retention Pond | LGRP
Heap Leach Pad | HLP
Batman Pit | RP3
Process Plant Retention Pond | PRP
Waste Rock Dump | WRD
Waste Rock Dump Retention Pond | RP1
Process Water Pond | PWP
Water Treatment Plant | WTP
Process Plant | PP

NOTE: Prepared by Vista Gold Corp.; updated on February 23, 2018

Figure 1‑2: Concessions

Figure 1‑3: General Arrangement

1.4Geology and Mineralization

The Project is situated within the southeastern portion of the Early Proterozoic Pine Creek Geosyncline (PCG). Meta-sediments, granitoids, basic intrusives, acid and intermediate volcanic rocks occur within this geological province.

The Batman deposit geology consists of a sequence of hornfelsed interbedded greywackes, and shales with minor thin beds of felsic tuff. Bedding is striking consistently at 325°, dipping at 40° to 60° to the southwest. Minor lamprophyre dykes trending north-south pinch and swell, crosscutting the bedding.

The deposits are similar to other gold deposits of the PCG and are classified as orogenic gold deposits in the subdivision of thermal aureole gold style. The Batman deposit shares some characteristics with intrusion-related gold systems, especially in terms of the association of gold with bismuth and reduced ore mineralogies. This makes the Batman deposit unique in the PCG. The mineralization within the Batman deposit is directly related to the intensity of the north-south trending quartz sulfide veining. The lithological units impact on the orientation and intensity of mineralization.

Sulfide minerals associated with the gold mineralization are pyrite, pyrrhotite and lesser amounts of chalcopyrite, bismuthinite and arsenopyrite. Galena and sphalerite are also present but appear to be post-gold mineralization and are related to calcite veining, bedding and the east-west trending faults and joints.

A variety of mineralization styles occur within the Project area. Of greatest known economic significance are auriferous quartz-sulfide vein systems. These vein systems include the Batman, Jones, Golf, Quigleys and Horseshoe prospects, which occur within a north-northeast trending corridor, and are hosted by the Burrell Creek Formation. Tin occurs in a north-northwest trending corridor. The tin mineralization comprises cassiterite, quartz, tourmaline, kaolin, and hematite bearing assemblages, which occur as bedding parallel to breccia zones and pipes. Polymetallic Au, W, Mo, and Cu mineralization occurs in quartz-greisen veins within the Yinberrie Leucogranite; a late stage highly fractionated phase of the Cullen Batholith. The Batman deposit extends approximately 2,200 meters (m) along strike, 400 m across dip and drill tested to a depth of 800 m. Drilling indicates the Batman mineralization to be open along-strike and down-dip.

To date, with regard to the exploration licenses (ELs), they represent an early-stage exploration program which has not produced an announceable discovery. While the work is promising and will be ongoing, there are no quantifiable resources or reserves on the ELs. Once an announceable discovery is made, Vista will detail that discovery according to all applicable disclosure regulations.

1.5Mineral Resource Estimate

The following sections summarize the process, procedures, and results of Tetra Tech’s independent estimate of the contained gold resources of the:

1)Batman deposit;

2)Existing heap leach pad; and

3)Quigleys deposit.

The resource estimate for the Batman deposit is updated from the July 7, 2014 Amended & Restated NI 43-101 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia, Technical Report prepared by Tetra Tech. This report includes an estimate of gold contained in a historic heap leach pad adjacent to the Batman deposit. Additionally, this report contains the resource estimation of the Quigleys deposit. The updated

Project resource estimates are shown in Table 1‑3, grade tonnage curve for the measured and indicated resource for the Batman deposit is presented in Figure 1‑4.

Table 1‑3: Statement of Mineral Resources Estimates

| Batman Deposit (August 2017) | Heap Leach Pad (May 2013) | Quigleys Deposit (August 2017)
Tonnes (000s) | Grade (g/t) | Contained Ounces (000s) | Tonnes (000s) | Grade (g/t) | Contained Ounces (000s) | Tonnes (000s) | Grade (g/t) | Contained Ounces (000s)
Measured (M) | 77,725 | 0.88 | 2,191 | - | - | - | 457 | 1.27 | 19
Indicated (I) | 200,112 | 0.80 | 5,169 | 13,354 | 0.54 | 232 | 5743 | 1.12 | 207
Measured & Indicated | 277,837 | 0.82 | 7,360 | 13,354 | 0.54 | 232 | 6,200 | 1.13 | 225
inferred (F) | 61,323 | 0.72 | 1,421 | - | - | - | 1,600 | 0.84 | 43

NOTES:

(1)Measured & indicated resources include proven and probable reserves.

(2)Batman and Quigleys resources are quoted at a 0.40g-Au/t cut-off grade. Heap Leach resources are the average grade of the heap, no cut-off applied.

(3)Batman: Resources constrained within a US$1,300/oz gold Whittle TM pit shell. Pit parameters: Mining Cost US$1.50/tonne, Milling Cost US$7.80/tonne processed, G&A Cost US$0.46/tonne processed, 50K TPD Ore, 355 Days/Yr., TPY Ore 17,750,000 TPY, G&A/Year 8,201 K US4, Au Recovery, Sulfide 85%, Transition 80%, Oxide 80%, 0.2g-Au/t minimum for resource shell. Selling Cost: US$/oz recovered US412.00.

(4)Quigleys: Resources constrained within a US$1200/oz gold Whittle TM pit shell. Pit parameters: Mining cost US$2.07/tonne, Milling Cost US$9.623/tonne processed, Sale Cost US$/oz US$15.18, Royalty 1% NPR (Jawoyn), Gold Recovery All Types, 70%.

(5)Differences in the table due to rounding are not considered material

(6)Rex Bryan of Tetra Tech is the qualified person responsible for the Statement of Mineral Resources for the Batman, Heap Leach Pad and Quiqleys deposits.

(7)Thomas Dyer of Mine Development Associates is the qualified person responsible for developing the resource Whittle TM pit shell for the Batman Deposit.

(8)The effective date of the Batman and Quigleys resource estimate is August 2017, the effective date of the Heap Leach resource is May 2013.

(9)Mineral resources that are not mineral reserves have no demonstrated economic viability and do not meet all relevant modifying factors.

Mineral resources that are not mineral reserves have no demonstrated economic viability and do not meet all relevant modifying factors.

Figure 1‑4: Measured & Indicated Resource Estimates Grade Tonnage Curves – Batman Deposit

1.6Mineral Reserve Estimates

Mine Development Associates (MDA) has used measured and indicated resources provided by Tetra Tech on August 2017 to estimate mineral reserves. Pit optimization was done using Geovia’s Whittle TM software to define pit limits with input for economic and slope parameters.

Optimization used only measured and indicated resources for processing. All inferred resource was considered as waste.

Varying gold prices were used to evaluate the sensitivity of the deposit to the price of gold as well as to develop a strategy for optimizing Project cash flow. To achieve cash-flow optimization, mining phases or push backs were developed using the guidance of Whittle TM pit shells at lower gold prices.

The statement of mineral reserve estimates is shown in Table 1‑4.

Table 1‑4: Statement of Mineral Reserve Estimate

| Batman Deposit (January 2018) | Heap Leach Pad (May 2013) | Total P&P Reserves (January 2018)
| Tonnes (000s) | Grade (g/t) | Contained Ounces (000s) | Tonnes (000s) | Grade (g/t) | Contained Ounces (000s) | Tonnes (000s) | Grade (g/t) | Contained Ounces (000s)
Proven | 72,672 | 0.88 | 2,057 | - | - | - | 72,672 | 0.88 | 2,057
Probable | 135,015 | 0.82 | 3,559 | 13,354 | 0.54 | 232 | 148,369 | 0.79 | 3,791
Proven & Probable | 207,687 | 0.84 | 5,616 | 13,354 | 0.54 | 232 | 221,041 | 0.82 | 5,848

NOTES:

(1)Thomas L. Dyer, P.E., is the QP responsible for reporting the Batman Deposit proven and probable reserves.

(2)Batman deposit reserves are reported using a 0.40 g-Au/t cutoff grade.

(3)Deepak Malhotra is the QP responsible for reporting the heap-leach pad reserves.

(4)Because all of the heap-leach pad reserves are to be fed through the mill, these reserves are reported without a cutoff grade applied.

(5)The reserves point of reference is the point where material is fed into the mill.

1.6.1Heap Leach Reserve Estimate

Existing heap leach pad (HLP) reserves are provided in Table 1‑4, which are estimated to be 13.4 million tonnes (Mt). These reserves will be processed through the mill at the end of the mine life.

Testwork indicated the following:

Cyanidation leach tests on “as is” material on the heap will extract ± 30% of the gold.

CIP cyanidation tests at a grind size of P80 of 90 microns will extract on average 72% of gold (range: 64.14% to 80.37%) in 24 hours of leach time. The average lime and cyanide consumptions were 1.75 kg/t and 0.78 kg/t, respectively.

1.7Mining Methods

The Project is designed to be a conventional, owner-operated, large open-pit mining operation that will use large- scale mining equipment in a drill/blast/load/haul operation. All dollar values in Section 1.7 are reported in US$.

A base gold price of US$1,250 per ounce was used for scenario analysis. However, various gold prices from US$300 to US$2,000 per ounce, in increments of US$20 per ounce, were used to determine different optimized pit shells. Economic parameters are provided in Table 1‑5.

Table 1‑5: Initial Economic Parameters

Parameter | Base Case | Alternate Case
Gold Recovery | 85% Sulfide 80% Transition 80% Oxide | 85% Sulfide 80% Transition 80% Oxide
Payable Gold | 99.9% | 99.9%
Overall Mining Cost | US$1.90 per tonne mined | US$2.16 per tonne mined
Processing Cost | US$7.80 per tonne processed | US$8.65 per tonne processed
Tailings | US$0.90 per tonne processed | US$0.90 per tonne processed
General & Administrative | US$0.46 per tonne processed | US$0.77 per tonne processed
Water Treatment | US$0.09 per tonne processed | US$0.10 per tonne processed
Royalty | 1% gross proceeds | 1% gross proceeds

The mining costs used were varied by bench. An incremental cost of US$0.010 was added for each 6 meter bench below the 145 meter elevation. This represents the incremental increase in cost of haulage for both waste and ore for each bench that is to be mined. Reference mining costs of US$1.64 and US$1.86 were used for the Base and Alternate Cases, respectively. Additionally, an incremental cost was determined based on truck operating costs, truck cycle time to haul and return through a 6 meter gain in elevation, and truck capacity. The reference mining cost was determined using first principles from previous studies. The total mining cost (reference plus incremental) is US$1.90 and US$2.16 for the Base and Alternate Cases, respectively.

Processing, tailings construction, tailings reclamation, and water treatment costs were provided by Vista based on previous studies. Calculated cutoff grades based on the economic parameters are 0.38 and 0.33 g‑Au/t for the Alternate and Base cases respectively. At Vista’s request, MDA used a minimum cutoff grade of 0.40 g-Au/t for both the Alternate and Base cases. This was done to maintain higher grades with respect to material allowed to be processed.

Several iterations of pit optimizations were reviewed to determine the final pit limits. Ultimately, the pit limits were chosen to reflect the approximate total pit size from the previous PFS. For the Base Case a US$1,000/oz‑Au pit shell was used to guide the ultimate pit design. For the Alternate Case the US$800/oz‑Au pit shell was used to guide the ultimate pit design. Table 1‑6 shows the Whittle TM optimization results for the Base Case and Table 1‑7 shows the pit optimization results for the Alternate Case. Note that pit results for the base Au price used for pit optimization of US$1,250/oz‑Au is highlighted in light green and the ultimate pit used for pit design is highlighted in orange.

Table 1‑6: Whittle TM Pit Optimization Results – Base Case Using 0.40 g-Au/t Cutoff

Pit | Gold Price (US$) | Material Processed | Waste Tonnes | Total Tonnes | Strip Ratio
K Tonnes
1 | $ 300 | 3,282 | 1.77 | 186 | 2,797 | 6,078 | 0.85
6 | $ 400 | 8,578 | 1.54 | 425 | 7,507 | 16,085 | 0.88
11 | $ 500 | 15,988 | 1.34 | 686 | 15,740 | 31,728 | 0.98
16 | $ 600 | 37,253 | 1.12 | 1,340 | 53,757 | 91,010 | 1.44
21 | $ 700 | 89,301 | 0.99 | 2,855 | 171,617 | 260,918 | 1.92
26 | $ 800 | 121,187 | 0.92 | 3,585 | 222,919 | 344,106 | 1.84
31 | $ 900 | 159,485 | 0.87 | 4,442 | 316,889 | 476,374 | 1.99
36 | $ 1,000 | 185,915 | 0.85 | 5,093 | 429,208 | 615,123 | 2.31
41 | $ 1,100 | 212,340 | 0.84 | 5,741 | 566,907 | 779,247 | 2.67
46 | $ 1,200 | 230,587 | 0.83 | 6,184 | 675,714 | 906,302 | 2.93
49 | $ 1,250 | 234,858 | 0.83 | 6,278 | 700,519 | 935,376 | 2.98
51 | $ 1,300 | 240,195 | 0.83 | 6,416 | 742,833 | 983,029 | 3.09
56 | $ 1,400 | 243,306 | 0.83 | 6,498 | 771,190 | 1,014,497 | 3.17
61 | $ 1,500 | 249,389 | 0.83 | 6,658 | 829,933 | 1,079,321 | 3.33
66 | $ 1,600 | 254,050 | 0.83 | 6,779 | 880,583 | 1,134,633 | 3.47
70 | $ 1,700 | 254,348 | 0.83 | 6,785 | 883,222 | 1,137,571 | 3.47
74 | $ 1,800 | 259,140 | 0.83 | 6,908 | 943,012 | 1,202,152 | 3.64
78 | $ 1,900 | 259,964 | 0.83 | 6,927 | 952,872 | 1,212,836 | 3.67
81 | $ 2,000 | 260,099 | 0.83 | 6,929 | 953,985 | 1,214,083 | 3.67

Pit 36 was used for design purposes and Pit 49 illustrates the potential floating cone using a US$1,250/oz-Au price.

Table 1‑7: Whittle TM Pit Optimization Results – Alternate Case Using 0.40 g-Au/t Cutoff

Pit | Gold Price (US$) | Material Processed | Waste Tonnes | Total Tonnes | Strip Ratio
K Tonnes
1 | $ 300 | 966 | 1.88 | 59 | 731 | 1,697 | 0.76
6 | $ 400 | 6,467 | 1.63 | 338 | 5,647 | 12,114 | 0.87
11 | $ 500 | 11,133 | 1.46 | 523 | 10,299 | 21,432 | 0.93
16 | $ 600 | 21,103 | 1.25 | 850 | 22,828 | 43,931 | 1.08
21 | $ 700 | 60,702 | 1.06 | 2,078 | 115,883 | 176,585 | 1.91
26 | $ 800 | 95,474 | 0.98 | 3,019 | 185,830 | 281,304 | 1.95
31 | $ 900 | 121,301 | 0.92 | 3,586 | 222,805 | 344,106 | 1.84
36 | $ 1,000 | 154,092 | 0.87 | 4,307 | 295,849 | 449,941 | 1.92
41 | $ 1,100 | 185,525 | 0.85 | 5,085 | 427,833 | 613,358 | 2.31
46 | $ 1,200 | 202,606 | 0.84 | 5,504 | 512,606 | 715,212 | 2.53
49 | $ 1,250 | 212,904 | 0.84 | 5,753 | 569,384 | 782,289 | 2.67
52 | $ 1,300 | 223,045 | 0.84 | 5,996 | 626,265 | 849,310 | 2.81

57 | $ 1,400 | 234,858 | 0.83 | 6,278 | 700,519 | 935,376 | 2.98
62 | $ 1,500 | 241,417 | 0.83 | 6,454 | 756,553 | 997,970 | 3.13
67 | $ 1,600 | 247,350 | 0.83 | 6,605 | 809,868 | 1,057,218 | 3.27
71 | $ 1,700 | 250,425 | 0.83 | 6,684 | 841,140 | 1,091,565 | 3.36
75 | $ 1,800 | 254,050 | 0.83 | 6,779 | 880,583 | 1,134,633 | 3.47
80 | $ 1,900 | 254,353 | 0.83 | 6,785 | 883,275 | 1,137,628 | 3.47
84 | $ 2,000 | 259,140 | 0.83 | 6,908 | 943,012 | 1,202,152 | 3.64

Pit 26 was used for design purposes and Pit 49 illustrates the potential floating cone using a US$1,250/oz-Au price.

1.8Metallurgy

The flowsheet consists of primary crushing, closed circuit secondary crushing, closed circuit tertiary crushing using high pressure grinding rolls (HPGRs), ore sorting, two-stage ball milling, cyclone classification, pre-leach thickening, leach and adsorption, elution electrowinning and smelting, carbon regeneration, tailings detoxification and disposal to conventional tailings storage facility (TSF). Figure 1‑5 provides the schematic diagram of the flowsheet.

Figure 1‑5: Mt Todd Flowsheet

1.9Mineral Processing

Detailed design criteria have been developed for the process plant. The nominal headline design criteria are listed in Table 1‑8 below.

Table 1‑8: Headline Design Criteria

| Unit | Base Case | Alternate Case
Annual Ore Feed Rate | Mt/a | 17.75 | 11.72
Operating Days per Year | d/a | 355 | 355
Daily Ore Feed Rate | t/d | 50,000 | 33,000
Crushing Rate (6,637 hours per year availability) | tph | 2,674 | 1,765
HPGR Rate (7,838 hours per year) | tph | 2,264 | 1,495
Ore Sorting Rate (7,838 hours per year) | tph | 408 | 318
Milling Rate (7,838 hours per year) | tph | 2,055 | 1,332
Gold Head Grade | g/t | 0.82 | 0.82
Copper Head Grade | % | 0.055 | 0.055
Cyanide Soluble Copper | % | 0.0024 | 0.0024
Bulk Density | t/m 3 | 2.76 | 2.76
Primary Grind P80 to Secondary Grind | µm | 250 | 250
Grind P80 to Leach | µm | 60 | 60
Gold Recovery | % | 85 | 85
Gold Production (average) | oz/d | 1,120 | 740
Gold Production (average) | oz/a | 397,761 | 262,634

1.10Project Infrastructure

Access to local resources and infrastructure is excellent. The Project is located sufficiently close to the city of Katherine to allow for an easy commute for workers. The area has both historic and current mining activity and therefore a portion of the skilled workforce will be sourced locally. In addition, Katherine offers the necessary support functions that are found in a medium-sized city with regard to supplies, accommodations, communications, etc.

The property has an existing high-pressure gas line and an electric power line that were used by previous operators. In addition, wells for potable water and a dam for process water are also located on or adjacent to the site. Finally, a side hill-type TSF is present on site.

Planned infrastructure for the site includes the following:

Ammonium Nitrate and Fuel Oil (ANFO) Facility;

Mine Support Facilities (Heavy Vehicle (HV) Workshop, Lube Farm, Washdown and Tire Change, Warehouse, Fuel Farm, Mining Offices, Core Storage Facility);

Heap Leach Pad (existing);

Accommodation Camp;

Water Treatment Plant (WTP);

Power Supply;

Pit Dewatering;

Mine Services;

Communications;

Gatehouse; and

Expanded existing and additional TSF.

1.11Market Studies and Contracts

1.11.1Markets

Gold metal markets are mature, with many reputable refiners and brokers located throughout the world. The advantage of gold, like other precious metals, is that virtually all production can be sold in the market. As such, market studies, and entry strategies are not required.

Metallurgical process studies confirm that the Project will produce doré of a specification comparable with existing operating mines.

Demand is presently high with prices showing remarkable increases during recent times. The 36-month average London PM gold price fix through December 31, 2017 is US$1,222/oz.

1.11.2Contracts

Currently there are no contracts for development. However, Vista has obtained budgetary quotes, as is common for PFS level studies, for future service needs. The following contracts are expected to be in place upon project commencement:

Secure doré transportation to market;

Doré refining;

Supplier and service contracts including;

Diesel and fuel oil;

Natural gas for the power plant;

Process reagents;

Equipment preventive maintenance and repair (MARC) services;

Site security services; and

Camp management, catering and support services.

1.12Social and Environmental Aspects

1.12.1Existing Environmental and Social Information

A number of environmental studies have been conducted at the Project in support of development of EISs and as required for environmental and operational permits. Studies conducted have investigated soils, climate and meteorology, geology, geochemistry, biological resources, cultural and anthropological sites, socio-economics, hydrogeology, and water quality.

In January 2018, the “authorization of a controlled activity” was received for the Project as required under the Australian Environmental Protection and Biodiversity Conservation Act of 1999 (EBPC) as it relates to the Gouldian Finch, and as such has received approval from the Australian Commonwealth Department of Environment and Energy.

The Mt Todd Project Environmental Impact Statement (EIS) submitted June 28, 2013 to the Northern Territory Environment Protection Authority (NTEPA), approved in September 2014, provides an understanding of the existing environmental conditions and an assessment of the environmental impact of the Project.

1.12.2Social or Community Requirements

The Jawoyn people have strong involvement in the planning for the future of the Project. Vista has a good relationship with the Jawoyn. Areas of aboriginal significance have been designated, and the mine plan has avoided development in these restricted works areas.

1.12.3Approvals, Permits and Licenses

The Project will require approvals, permits and licenses for various components of the Project. Table 1‑9 includes a list of approvals, permits, and licenses required for the Project and their current status.

Table 1‑9: Mt Todd Permit Status

Approval/ Permit/ License | Current Status | Approval/ Permit License Date | Expiration Date
Environmental Impact Statement | The NT Environmental Protection Authority provided its final assessment of the Project in June 2014. | Approved Sep. 2014 | NA
Mining Management Act (or Plan) Approval from NT Department of Mines & Energy | Mine operating permit request has not been submitted. Interim MMP submitted in November 2017 regarding current operations. | Prior to commencing mine operations | NA
Heritage Act permit to destroy or damage archeological sites and scatters/ Aboriginal Areas Protection Authority Clearances | Authority Certificate Number 2011/15538 issued. This certificate defined restricted works areas and granted select clearances to allow for initial investigations. Additional clearances will be required for further investigations as well as prior to disturbance associated with mine development. | Aboriginal Areas Protection Authority dated Jul. 31, 2012 | NA
Dangerous Goods Act (1988) permit for blasting activities | Waiting on final mine plan | NA | NA
Extractive Permit (under DME Guidelines) for development of borrow pits outside of approved mining areas | Would be required for PGM or LPM borrow areas. Permit application not yet in progress pending final selection of borrow areas | NA | NA
Waste Discharge License (under Section 74 of the Water Act 1992) for management of water discharge from the site | WDL 178-5 licensing discharge of waste water into the Edith River from the Mt Todd mine site, granted with conditions | Jan. 16, 2017 | Aug. 31, 2018
Waste water treatment system permits under Public Health Act 1987 and Regulations | May be required for the waste water treatment system for the construction and operations accommodation village. Permit application not yet in progress pending design and siting of accommodation village. | NA | NA
Approval to Disturb Site of Conservation Significance (SOCS) | Batman pit expansion will disturb SOCS as breeding / foraging habitat for the Gouldian finch, pending determination on EIS. | Jan. 22, 2018 | NA

1.13Capital and Cost Estimates

1.13.1Capital Cost Estimates – Base Case

LoM capital cost requirements are estimated at US$1,191 million as summarized in Table 1‑10. Initial capital of US$839 million is required to commence operations. At the end of operations, the Project will receive an estimated US$142 million credit for asset sales and salvage.

Table 1‑10: Estimated Capital Cost Summary (US$000s)

Area | Description | Cont. (%) | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
Estimate | Contingency | Total | Estimate | Contingency | Total | Estimate | Contingency | Total
SUMMARY
2000 | Mining | 0.3% | $144,278 | $1,204 | $145,482 | $385,207 | $237 | $385,445 | $529,485 | $1,441 | $530,927
3000 | Process Plant | 20.7% | $340,768 | $72,386 | $413,154 | $22,992 | $2,951 | $25,943 | $363,760 | $75,337 | $439,097
4000 | Project Services | 3.5% | $117,676 | $7,085 | $124,760 | $82,857 | $0 | $82,857 | $200,533 | $7,085 | $207,617
5000 | Project Infrastructure | 12.9% | $23,263 | $3,001 | $26,264 | $0 | $0 | $0 | $23,263 | $3,001 | $26,264
6000 | Permanent Accommodation | 10.0% | $68 | $7 | $75 | $0 | $0 | $0 | $68 | $7 | $75
7000 | Site Establishment & Early Works | 11.4% | $19,723 | $2,248 | $21,972 | $0 | $0 | $0 | $19,723 | $2,248 | $21,972
8000 | Management, Engineering, EPCM Services | 9.5% | $85,941 | $8,139 | $94,080 | $0 | $0 | $0 | $85,941 | $8,139 | $94,080
9000 | Pre-Production Costs | 17.2% | $11,127 | $1,918 | $13,045 | $0 | $0 | $0 | $11,127 | $1,918 | $13,045
10000 | Asset Sale | 0.0% | $0 | $0 | $0 | ($141,788) | $0 | ($141,788) | ($141,788) | $0 | ($141,788)
| Capital Cost | 9.1% | $742,844 | $95,988 | $838,832 | $349,269 | $3,189 | $352,458 | $1,092,113 | $99,177 | $1,191,290

1.13.2Operating Cost Estimates – Base Case

LoM operating costs requirements are estimated to be US$14.18/t-milled as summarized in Table 1‑11.

Table 1‑11: Estimated LoM Operating Costs (US$)

Description | US$/t-milled
OPEN PIT MINE |
Mine General Service | $0.07
Mine Maintenance | $0.12
Engineering | $0.05
Geology | $0.03
Drilling | $0.73
Blasting | $1.20
Loading | $0.62
Hauling | $2.89
Mine Support | $0.37
Mine Dewatering | $0.01
Open Pit Mine | $6.08
CIP PROCESS PLANT |
Labor | $0.89
3100-Crush/Screen/Stockpile | $0.19
3200-Reclaim & HPGR | $0.49
3300-Classification & Grinding | $2.66
3400-Pre-Leach,Thick/Aeration/CIP | $0.15
3500-Desorption, Gold Room | $0.02
3600-Detox & Tailings Pumping | $0.08
3700-Reagents | $2.11
3800-Plant Services | $0.04
Mining, Infrastructure & Misc | $0.05
General Consumables | $0.01
Plant Mobile Equipment | $0.01
Plant Gas Consumption | $0.02
CIP Process Plant | $6.71
Project Services | $0.16
G&A | $1.22
Operating Costs | $14.18

1.13.3Capital Cost Estimates – Alternate Case

LoM capital cost requirements are estimated at US$870 million as summarized in Table 1‑12. Initial capital of approximately US$641 million is required to commence operations. At the end of operations, the Project will receive an estimated US$96 million credit for asset sales and salvage.

Table 1‑12: Estimated Capital Cost Summary, Alternate Case (US$000s)

Area | Description | Cont. (%) | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
Estimate | Contingency | Total | Estimate | Contingency | Total | Estimate | Contingency | Total
2000 | Mining | 0.4% | $75,599 | $1,194 | $76,793 | $264,265 | $168 | $264,433 | $339,863 | $1,363 | $341,226
3000 | Process Plant | 21.6% | $264,474 | $57,888 | $322,363 | $11,209 | $1,537 | $12,746 | $275,683 | $59,425 | $335,109
4000 | Project Services | 4.0% | $97,835 | $5,757 | $103,592 | $47,687 | $0 | $47,687 | $145,522 | $5,757 | $151,279
5000 | Project Infrastructure | 12.9% | $22,881 | $2,958 | $25,838 | $0 | $0 | $0 | $22,881 | $2,958 | $25,838
6000 | Permanent Accommodation | 10.0% | $68 | $7 | $75 | $0 | $0 | $0 | $68 | $7 | $75
7000 | Site Establishment & Early Works | 11.4% | $18,579 | $2,127 | $20,706 | $0 | $0 | $0 | $18,579 | $2,127 | $20,706
8000 | Management, Engineering, EPCM Services | 9.0% | $74,977 | $6,720 | $81,697 | $0 | $0 | $0 | $74,977 | $6,720 | $81,697
9000 | Pre-Production Costs | 17.3% | $8,734 | $1,509 | $10,243 | $0 | $0 | $0 | $8,734 | $1,509 | $10,243
10000 | Asset Sale | 0.0% | $0 | $0 | $0 | ($96,170) | $0 | ($96,170) | ($96,170) | $0 | ($96,170)
| Capital Cost | 10.1% | $563,147 | $78,160 | $641,307 | $226,991 | $1,706 | $228,697 | $790,138 | $79,865 | $870,004

1.13.4Operating Cost Estimates — Alternate Case

LoM operating cost estimates are summarized in Table 1‑13. The operating costs will average US$13.61 over the LoM.

Table 1‑13: Estimated LoM Operating Costs, Alternate Case (US$)

Description | US$/t-milled
OPEN PIT MINE |
Mine General Service | $0.10
Mine Maintenance | $0.16
Engineering | $0.07
Geology | $0.05
Drilling | $0.52
Blasting | $0.84
Loading | $0.45
Hauling | $1.69
Mine Support | $0.52
Mine Dewatering | $0.01
Open Pit Mine | $4.40
CIP PROCESS PLANT |
Labor | $1.22
3100-Crush/Screen/Stockpile | $0.24
3200-Reclaim & HPGR | $0.55
3300-Classification & Grinding | $2.69
3400-Pre-Leach,Thick/Aeration/CIP | $0.17
3500-Desorption, Gold Room | $0.02
3600-Detox & Tailings Pumping | $0.07
3700-Reagents | $2.12
3800-Plant Services | $0.04
Mining, Infrastructure & Misc | $0.06
General Consumables | $0.01
Plant Mobile Equipment | $0.01
Plant Gas Consumption | $0.02
CIP Process Plant | $7.23
Project Services | $0.18
G&A | $1.80
Operating Costs | $13.61

1.14Financial Analysis

1.14.1Financial Analysis – Base Case

Estimated economic results are summarized in Table 1‑14. The analysis suggests the following conclusions, assuming a 100% equity project at a gold price of US$1,300:

...........Mine Life: 13 years;

Pre-Tax NPV5%: US$1,178 million, IRR: 27.3%;

After-tax NPV5%: US$679 million, IRR: 20.5%;

Payback (After-tax): 3.2 years;

NT Royalty Taxes Paid: US$357 million;

Australian Income Taxes Paid: US$466 million; and

Cash costs (including Jawoyn Royalty): US$645.33/oz-Au.

Table 1‑14: Estimated Technical-Economic Results (US$000s)

Cash Flow Summary | LoM Cost (US$000s) | Unit Cost US$/t-milled | Unit Cost US$/oz-Au
Gold Produced (koz) | 4,956 | - | -
Gold Price (US$/oz) | 1,300 | - | -
Gold Sales | 6,442,462 | 29.15 | 1,300
Refinery Costs | (15,763) | (0.071) | (3.18)
Net Smelting Return | 6,426,699 | 29.075 | 1,297
Jawoyn Royalty | (64,425) | (0.291) | (13.00)
Gross Income from Mining | 6,362,274 | 28.783 | 1,284
Open Pit Mine | (1,343,917) | (6.08) | (271)
CIP Process Plant | (1,484,260) | (6.71) | (300)
Project Services | (35,169) | (0.16) | (7.10)
G&A | (270,324) | (1.22) | (54.55)
Operating Costs | (3,133,670) | (14.18) | (632.33)
Power Sales Credit | 13,715 | 0.062 | 2.77
Cash Cost of Goods Sold (COGS) | (3,135,719) | (14.48) | (645.33)
Operating Margin | 3,242,319 | 14.67 | 654.26
Mining | 530,927 |  |
Process Plant | 439,097 |  |
Project Services | 207,617 |  |
Project Infrastructure | 26,264 |  |
Permanent Accommodation | 75 |  |
Site Establishment & Early Works | 21,972 |  |
Management, Engineering, EPCM Services | 94,080 |  |
Pre-Production Costs | 13,045 |  |
Asset Sale | (141,788) |  |

Capital Costs | 1,191,290 |  |
Pre-Tax Cash Flow | 2,051,029 |  |
NPV5% | 1,178,110 |  |
IRR (%) | 27.3% |  |
After-tax Cash Flow | 1,245,596 |  |
NPV5% | 679,021 |  |
IRR (%) | 20.5% |  |
After-tax Payback (years) | 3.2 |  |

1.14.2Financial Analysis – Alternate Case

Economic results are summarized in Table 1‑15. The analysis suggests the following conclusions, assuming a 100% equity project at a gold price of US$1,300:

Mine Life: 11 years;

Pre-Tax NPV5%: US$722 million, IRR: 23.4%;

After-tax NPV5%: US$418 million, IRR: 17.8%;

Payback (After-tax): 3.6 years;

NT Royalty Taxes Paid: US$209 million;

Australian Income Taxes Paid: US$282 million; and

Cash costs (including Jawoyn Royalty): US$593.38/oz-Au.

Table 1‑15: Estimated Economic Results, Alternate Case (US$000s)

Cash Flow Summary | LoM Cost (US$000s) | Unit Cost US$/t-milled | Unit Cost US$/oz-Au
Gold Produced (koz) | 3,003 | - | -
Gold Price (US$/oz) | 1,300 | - | -
Gold Sales | 3,903,895 | 30.50 | 1,300
Refinery Costs | (9,737) | (0.076) | (3.242)
Net Smelting Return | 3,894,158 | 30.420 | 1,297
Jawoyn Royalty | (39,039) | (0.305) | (13.00)
Gross Income from Mining | 3,855,119 | (30.115) | 1,284
Open Pit Mine | (563,491) | (4.40) | (188)
CIP Process Plant | (925,686) | (7.23) | (308)
Project Services | (23,326) | (0.18) | (8)
G&A | (230,369) | (1.80) | (77)
Operating Costs | (1,742,873) | (13.61) | (580)
Power Sales Credit | 13,056 | 0.102 | 4
Cash Cost of Goods Sold (COGS) | (1,739,554) | (13.59) | (579)

Operating Margin | 2,125,302 | 16.60 | 708
Mining | 341,226 |  |
Process Plant | 335,109 |  |
Project Services | 151,279 |  |
Project Infrastructure | 25,838 |  |
Permanent Accommodation | 75 |  |
Site Establishment & Early Works | 20,706 |  |
Management, Engineering, EPCM Services | 81,697 |  |
Pre-Production Costs | 10,243 |  |
Asset Sale | (96,170) |  |
Capital Costs | 870,004 |  |
Pre-Tax Cash Flow | 1,255,298 |  |
NPV5% | 721,977 |  |
IRR (%) | 23.4% |  |
After-tax Cash Flow | 778,521 |  |
NPV5% | 417,938 |  |
IRR (%) | 17.8% |  |
Post- Tax Payback (years) | 3.6 |  |

1.15Conclusions and Recommendations

1.15.1Feasibility Study

A Feasibility Study (FS) should be completed to advance the Project and provide the detailed information necessary to support capital and operating cost estimates for a potential project development decision.

The estimated budget for the FS is approximately US$2.5M – 4.0M.

1.15.2Geology and Resources

1.15.2.1Conclusions

The Project is situated within the southeastern portion of the Early Proterozoic Pine Creek Geosyncline which is comprised of the Burrell Creek Formation, the Tollis Formation, and the Kombolgie Formation.

Gold mineralization in this area is constrained to a single mineralization event and the deposits are classified as orogenic gold deposits in the subdivision of thermal aureole gold style. The Batman deposit has characteristics of an intrusion related gold system making it the primary resource.

The Batman deposit is defined by approximately 7.4 million ounces (Moz) of gold within 278 Mt of measured and indicated resource at an average grade of 0.82 g-Au/t and a cutoff grade of 0.4 g-Au/t. as provided in Table 1‑3.

1.15.2.2Recommendations

The Batman deposit potentially extends along strike both to the north and south. Step out drilling should be evaluated.

Additional deep infill drilling should be used to help define potential deep mineralization not detected by early historical drillholes.

Infill drilling within and exploration drillholes along the trend of the Quigleys deposit is recommended.

Exploration of the exploration licenses, including work on geophysical and geochemical anomalies, should continue in a systematic manner.

The estimated budget for drilling within the MLs is US$500,000-1,000,000 and US$500,000-1,000,000 for initial drilling on the ELs.

1.15.3Mineral Reserve and Mine Planning

Pit designs were completed based on Whittle TM pit optimizations and are appropriate for metal prices of approximately US$800 per ounce Au for the Alternate Case and US$1,000 per ounce Au for the Base Case. The Mt Todd proven and probable reserve estimates have been defined using economics based on a gold price of US$1,250 per ounce and an elevated cutoff grade of 0.40 g‑Au/t. The proven and probable reserve estimates were used to create a production schedule for mining, and a positive cash-flow analysis has been done based on the production schedule by Tetra Tech. The reserve estimates have reasonable economics with respect to the statement of reserves under NI 43-101 regulations.

Mine production constraints were imposed to ensure that mining wasn’t overly aggressive with respect to the equipment anticipated for use at Mt Todd. The schedule has been produced using mill targets and stockpiling strategies to enhance the project economics. The constraints and limits are reasonable to support the project economics which are used to justify the statement of reserve estimates.

Pit designs use six-meter benches for mining. This corresponds to the resource model block heights, and MDA believes this to be reasonable with respect to dilution and equipment anticipated to be used in mining. In areas where the material is consistently ore or waste so that dilution is not an issue, benches may be mined in 12 m heights.

1.15.4Mineral Processing

1.15.4.1Conclusions

The substantial quantity and quality of metallurgical test work data developed from Mt Todd drill core samples has led to the development of a robust energy efficient comminution circuit followed by a standard gold recovery process. Key conclusions drawn from the metallurgy studies are:

Mt Todd (Batman) ore is among the hardest and most competent ore types processed for mineral recovery. The most energy efficient comminution circuit has been determined to be the sequence of primary crushing, closed circuit secondary crushing, and closed circuit HPGR tertiary crushing and ore sorting followed by two stages of grinding.

The ore is free-milling, is not preg-robbing, and is amenable to gold extraction by conventional cyanidation processes.

The ore has moderate cyanide consumption, determined to be 0.4 kg of sodium cyanide per tonne of ore. This is largely due to the presence of sulfides and cyanide consuming copper.

The use of sorting has helped to decrease operational costs and remove portions of the uneconomic material mined.

The equipment selection criteria for the Base Case operation has received considerable interaction with specialist vendors to the point where there is a reasonably high-degree of confidence in selected technology and process units at this preliminary feasibility study stage. The recommended flowsheet for a FS consists of primary crushing, closed circuit secondary crushing, closed circuit tertiary crushing using HPGRs and ore sorting, two-stage ball milling, cyclone classification, secondary grinding, cyclone classification, pre-leach thickening, leach and adsorption, elution electrowinning and smelting, carbon regeneration, tailings detox and disposal to conventional tailings storage facility.

1.15.4.2Recommendations

The on-going testwork is directed towards optimization of the process parameters. The study would potentially lead to reduction of reagent usage thereby reducing operating costs. In addition, the potential of increasing leach pulp density is being evaluated which could result in reducing size of leach tanks and hence capital cost.

1.15.5Infrastructure

Bulk earthworks are designed to minimize the import of fill materials.

Administration offices, gatehouse/security facilities, cribs/ablutions are planned to be transportable buildings.

The process plant offices, workshop and warehouse are located inside the existing Flotation Building.

Sample preparation and laboratory will have a purpose-built steel shed.

The access road is based on the repaired existing road.

Heavy cranage is allowed for all lifts greater than 50 t.

All bulk transport will be weighed.

Site-wide communication is based on a 50 m tall communication tower that will support eight channels.

1.15.6Environmental and Social Impacts

1.15.6.1Conclusions

A number of environmental studies have been conducted at the Project Site in support of Environmental Impact Statements (EIS) and as required for environmental and operational permits. Studies conducted have investigated soils, climate and meteorology, geology, geochemistry, biological resources, cultural and anthropological sites, socio-economics, hydrogeology, and water quality.

The draft EIS for the Project was submitted in June 2013. The document was prepared by independent consultants GHD Pty Ltd to identify potential environmental, social, transport, cultural and economic impacts associated with reopening and operating the mine. The Northern Territory Environmental Protection Authority (NTEPA) provided its final assessment of the Project in June 2014. Final approval was given in September 2014.

In January 2018, the “authorization of a controlled activity” was received for the Project as required under the Australian Environmental Protection and Biodiversity Conservation Act of 1999 (EBPC) as it relates to the Gouldian Finch, and as such has received approval from the Australian Commonwealth Department of Environment and Energy.

The Jawoyn people have strong involvement in the planning of the Project. Areas of aboriginal significance have been designated, and the mine plan has avoided development in these restricted works areas.

1.15.6.2Recommendations

Additional studies will be needed to further assess environmental baseline conditions to support feasibility level design, permitting, and closure planning for the Project, including:

Erosion analyses;

Waste and cover material hydraulic properties characterization and analysis;

Ongoing aquatic, benthic, and wildlife studies;

Comprehensive vegetation survey;

Archaeological and historical assessments for all areas to be disturbed;

Further hydrogeologic investigations and site-wide hydrogeologic characterization; and

Continued precipitation, stream flow, and watershed data.

The estimated budget for this work is US$350,000.

1.15.7Results of the Site-wide Water Balance Model

1.15.7.1Conclusions

The WTP rate of 500 m 3/hr and process water pond (PWP) sizing for 6 days of storage was determined to be appropriate for the 50,000 tpd production process water requirements.

The greatest amount of make-up water required from the raw water dam (RWD) was quantified as 11,955 m 3/day. RWD requirements were found to be the most dependent upon TSF decant volumes.

The Waste Rock Dump (WRD) retention pond (RP1), low grade ore stockpile retention pond (LGRP), process plant retention pond (PRP), and heap leach pad (HLP) were typically observed to overtop less than 1% of the time during the 13 year simulation.[1] LGRP storage may be optimized.

1.15.7.2Recommendations

Recommended model improvements include:

The site-wide water balance model is dependent upon the TSF water balance model, which provides decant water to the process facility, and the vadose and seepage models, which characterize seepage through the various rock piles on site (WRD, low grade ore stockpile (LGOS), and HLP). As such, completion of these models to the greatest detail practicable affects the overall quality of the site-wide water balance model results.

Incorporate results of future Batman Pit potential groundwater inflow investigation.

Optimize management of Batman Pit dewatering effluent and other contact water. This water may be of sufficient quality to be used as make-up water to the process circuit offsetting the water coming from the RWD. This water would also reduce the amount of water ultimately requiring treatment.

In this iteration of the site-wide water balance model, the entirety of the WTP effluent is being used as dust suppression around the mine site during the dry season. Further investigation of other uses of the WTP effluent should be conducted.

Further investigation of the adequacy of RP1 storage capacity is recommended, particularly within the early stages of the LoM when a larger fraction of the catchment reports to this pond.

Incorporate RWD stage-storage relationship and catchment area into the site-wide water balance model such that it may be modeled as a reservoir, as opposed to an infinite source.

Inclusion of process, fire, potable and raw water tanks. At present, only the dust suppression tank is modeled. The tanks above are currently modeled as drawing water directly from the RWD, rather than demands on discrete tanks.

Review and update dust suppression requirements.

Incorporate evolving Batman Pit shell geometry to more accurately model that facility.

The estimated budget for this work is US$200,000.

[1]A typical value is given. Separate model runs provide a range of overtopping events, due to the stochastic nature of the model.

1.15.8Groundwater Hydrology and Mine Dewatering

The following work is recommended with respect to groundwater hydrology and mine dewatering:

Additional hydrogeologic study should be completed in the vicinity of the Batman Pit to provide more detailed information on which to base calibration of the regional groundwater flow model and subsequent prediction of groundwater inflows to the pit and post-mining recovery of the groundwater system. The study should include permeability testing by packer test methods in three to five borings within or immediately adjacent to the pit and drilled to the proposed ultimate depth of the pit. The packer testing in those borings should be conducted over relatively short (10 m to 30 m) intervals, beginning at the water table and extending to the total depth of the borings. This testing could be combined with exploration drilling by performing the testing in exploration core holes. The study should also include measurement of depth to water in any accessible existing borings or core holes within or immediately adjacent to the pit.

Calibration of the regional groundwater flow model should be completed with the additional data, and the calibrated model should be used to refine the estimates of groundwater inflow to the pit and predictions of the hydrogeologic effects of pit dewatering.

The post-mining version of the groundwater flow model should be updated with the calibrated model used as its basis. Output from the post-mining model should be incorporated into any geochemical modeling of post-mining pit lake formation and geochemistry.

The estimated budget for this work is US$400,000.

1.15.9Process Plant Geotechnical Investigation

1.15.9.1Conclusions

There are no conclusions with regard to the Process Plant geotechnical investigation.

1.15.9.2Recommendations

Future geotechnical work is suggested in the following areas for foundation design:

GRINDING

This is a large vibrating structure which should be founded in rock rather than on fill to reduce dynamic effects. An accurate rock level is required to confirm foundation design and accurately estimate required concrete quantities and rock excavation. Existing structure concrete slabs are located directly over the new mill location, so it is recommended that additional new test pits around all four sides be undertaken to allow interpolation.

LEACH/CIP

Large tanks to be constructed in this area, with high foundation bearing pressure. Variation in rock level will impact on differential settlement which needs to be considered. Recommend at least four additional test pits to evaluate variance in rock levels in north-south and east-west directions.

STOCKPILE & RECLAIM

No test pit data anywhere near this area (nearest test pit is more than 200 m away). Steeply sloping ground surface exists (in excess of 5 m variation in ground level across the reclaim tunnel) so there may be considerable variation in rock levels which will effect potential settlement due to stockpile surcharge and required excavation for the concrete vault and tunnel. Recommend a new borehole on the high side (uncertain whether this high side bench is fill material, and therefore whether rock would be encountered within the limits of an excavator for a test pit) and a new test pit on low side, to determine rock levels.

COARSE SCREENING

Vibrating structure which should be founded on rock rather than on fill to reduce dynamic effect. An accurate rock level is required to confirm foundation design and accurately estimate required concrete quantities and rock excavation. Existing test pits are on one side only so unable to interpolate between existing pits, the nearest existing test pit is approximately 100 m to the north so recommend a new test pit at this location.

HPGR

Vibrating structure which should be founded on rock rather than on fill to reduce dynamic effect. Therefore need to know rock level to confirm foundation design and accurately estimate required concrete quantities and rock excavation. Nearest existing test pit is approximately 70 m away so recommend new test pit at this location.

The estimated budget for this work is US$150,000.

1.15.10TSF Design

As part of advancing the TSF design in the FS, work should include optimization of the TSF construction schedule, geotechnical investigation and assessment of TSF 1 and TSF 2, TSF water balance update, and TSF consequence classification. The estimated budget for this work is included in the FS budget estimate.

1.15.11Process

Two major items incurring operating costs are grinding media and reagents. Together these items make up 71% of the plant consumables operating costs. The FS should investigate options for reducing the consumption rate and the unit costs for these consumables.

The estimated budget for this work is included in the FS budget estimate.

1.15.12Geochemical Analyses

Geochemical characterization will be updated to reflect the designations of Potentially Acid Forming, Potentially Acid Forming-Low Capacity, Non Acid Forming, Acid Consuming and Uncertain in accordance with DITR (2007) guidelines. Additionally, Tetra Tech recommends performing geochemical testing on the sorter reject material.

The estimated budget for this work is included in the Feasibility Study.

2.0INTRODUCTION

Vista Gold Corp. and its subsidiaries (collectively, “Vista,” the “Company,” “we,” “our,” or “us”) operate in the gold mining industry. We are focused on the evaluation, acquisition, exploration and advancement of gold exploration and potential development projects, which may lead to gold production or value adding strategic transactions such as earn-in right agreements, option agreements, leases to third parties, joint venture arrangements with other mining companies, or outright sales of assets for cash and/or other consideration. We look for opportunities to improve the value of our gold projects through exploration drilling and/or technical studies focused on optimizing previous engineering work. We do not currently generate cash flows from mining operations.

The Company’s flagship asset is its 100% owned Mt Todd gold Project (Mt Todd) in the Northern Territory (NT) Australia. Mt Todd is the largest undeveloped gold project in Australia. The Company recently received authorization for the last major environmental permit and completed an updated Preliminary Feasibility Study (PFS) for Mt Todd, which confirms the project’s robust economics at today’s gold price. With these important milestones complete, Vista is in a position to actively pursue those strategic alternatives that provide the best opportunity to maximize value for the Company.

Vista Gold Corp. was originally incorporated on November 28, 1983 under the name “Granges Exploration Ltd.” It amalgamated with Pecos Resources Ltd. during June 1985 and continued as Granges Exploration Ltd. In June 1989, Granges Exploration Ltd. changed its name to Granges Inc. Granges Inc. amalgamated with Hycroft Resources & Development Corporation during May 1995 and continued as Granges Inc. Effective November 1996, Da Capo Resources Ltd. and Granges, Inc. amalgamated under the name “Vista Gold Corp.” and, effective December 1997, Vista continued from British Columbia to the Yukon Territory, Canada under the Business Corporations Act (Yukon Territory). On June 11, 2013, Vista continued from the Yukon Territory, Canada to the Province of British Columbia, Canada under the Business Corporations Act (British Columbia).

2.1Background Information

Vista Gold Corp. (Vista) retained Tetra Tech, Inc. (Tetra Tech) along with Mine Development Associates (MDA), Resource Development Inc. (RDi), Tetra Tech Proteus (TTP) and POWER Engineers, Inc. (POWER) to prepare this preliminary feasibility study (PFS) for its Project (the Project) in Northern Territory, Australia. The PFS evaluates the Base Case, a development scenario of a 50,000 tonnes per day (tpd) processing facility. In addition, an Alternate Case was considered for a 33,000 tpd processing facility with higher grades. This case is presented under Section 24.0 – Other Relevant Data and Information.

Key differences between the Base Case and the Alternate Case include:

A 33,000 tpd processing facility as compared to a 50,000 tpd facility with associated lower mining rates and a smaller mining fleet;

Pit design is based on a pit shell calculated using a US$1,000/oz-Au and a US$800/oz-Au for the Base and Alternate Cases, respectively. The same cut-off grade of 0.40 g-Au/t was used; and

Shorter operating life for the Alternate Case.

The Base Case includes:

Estimated proven and probable reserves of 5.848 Moz of gold (221 Mt at 0.82 g-Au/t) at a cut-off grade of 0.40 g-Au/t;

Average annual production of 381,211 ounces of gold per year over the mine life, including average annual production of 479,450 ounces of gold per year during the first five years of operations;

LoM average cash costs of US$645 per ounce, including average cash costs of US$571 per ounce during the first five years of operations;

A 13 year operating life;

After-tax NPV5% of US$679 million and internal rate of return (IRR) of 20.5% at US$1,300 per ounce gold prices and US$0.80:AUD1.00 exchange rate, and

Initial capital requirements of US$839 million.

The Alternate Case discussed in Section 24.0 – Other Relevant Data and Information includes:

Estimated proven and probable reserves of 3.557 Moz of gold (121 Mt at 0.86 g-Au/t) at a cut-off grade of 0.40 g-Au/t;

Average annual production of 273,000 ounces of gold per year over the mine life, including average annual production of 301,778 ounces of gold per year during the first five years of operations;

LoM average cash costs of US$593 per ounce, including average cash costs of US$581 per ounce during the first five years of operations;

An 11 year operating life;

After-tax NPV5% of US$418 million and IRR of 17.8% at US$1,300 per ounce gold prices and US$0.80:AUD1.00 exchange rate; and

Initial capital requirements of US$641 million.

2.2Terms of Reference and Purpose of the Report

This Technical Report was prepared as a NI 43-101 Technical Report for Vista by Tetra Tech. The quality of information, conclusions, and estimates contained herein are consistent with the level of effort involved in Tetra Tech’s 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 provides mineral resource and mineral reserve estimates, and a classification of resources and reserves in accordance with the CIM Standards. The CIM Standards requires the completion of a PFS as the minimum prerequisite for the conversion of mineral resources to mineral reserves.

A preliminary feasibility study is a comprehensive study of a range of options for the technical and economic viability of a mineral project that has advanced to a stage where a preferred mining method and the open pit configuration is established and an effective method of mineral processing is determined. It includes a financial analysis based on reasonable assumptions on the modifying factors and the evaluation of any other relevant factors which are sufficient for a qualified person, acting reasonably, to determine if all or part of the mineral resource may be converted to a mineral reserve at the time of reporting. Modifying factors are considerations used to convert mineral resources to mineral reserves. These include, but are not restricted

to, mining, processing, metallurgical, infrastructure, economic, marketing, legal, environmental, social and governmental factors.

A PFS is at a lower confidence level than a Feasibility Study (FS).

2.3Sources of Information

The primary technical documents and files relating to the Project that were used in the preparation of this report are listed in Section 27.0 — References.

2.4Units of Measure

The metric system has been used throughout this report. Tonnes are metric of 1,000 kilograms (kg), or 2,204.6 pounds (lb). Gold is reported in troy ounces (oz), equivalent to 31.1035 grams (g). Currency is in Q4 2017 U.S. dollars (US$) unless otherwise stated.

2.5Detailed Personal Inspections

1)Rex Bryan last visited and inspected the property from June 28 th through June 29 th, 2017. Dr. Bryan spent time on site and reviewed the current database and archived supporting material, core logging, sampling procedures, handling and security measures, QA/QC procedures and inspected modern and historically collected core.

2)Anthony Clark last visited and inspected the property from June 28 th through June 29 th, 2017. Mr. Clark inspected the existing power infrastructure at the site, natural gas pipeline, and power rights-of-way.

3)Thomas Dyer last visited and inspected the property from June 28 th through June 29 th, 2017. Mr. Dyer toured the site along with geotechnical consultants and reviewed the pit, waste dump, tailings facility, and resource drilling sites. Previous mine production records held on site were also reviewed.

4)Chris Johns visited and inspected the property from June 28 th through June 29 th, 2017. Mr. Johns inspected the existing Tailings Storage Facility 1 (TSF 1) and the proposed site for Tailings Storage Facility 2 (TSF 2).

5)Zvonimir Ponos last visited and inspected the property from June 28 th through June 29 th, 2017. Mr. Ponos inspected the existing site infrastructure and process facility.

6)Vicki Scharnhorst visited and inspected the property from June 28th through June 29th, 2017. Ms. Scharnhorst inspected the infrastructure at site and reviewed the status of environmental permitting with site staff.

QPs not listed above have not visited or inspected the property. Personal inspections by these QPs are not required to complete their responsibilities.

3.0RELIANCE ON OTHER EXPERTS

The QPs used their experience to determine if the information from previous reports was suitable for inclusion in this Technical Report. This report includes technical information which required subsequent calculations to derive subtotals, totals and weighted averages. Such calculations inherently involve a degree of rounding and consequently introduce a margin of error. Where these occur, the QPs do not consider them to be material.

The QPs relied upon the following experts:

Environmental Impact Statement for the Project prepared by GHD (June 2013) to describe environmental matters (Tetra Tech, Section 20.0 – Environmental Studies, Permitting, and Social or Community Impact).

4.0PROPERTY DESCRIPTION AND LOCATION

4.1Location

The Project is located 56 kilometers (km) by road northwest of Katherine, and approximately 290 km southeast of Darwin in NT, Australia (Figure 4‑1). Access to the property is via high quality, two‑lane paved roads from the Stuart Highway, the main arterial within the territory.

4.2Property Description

Vista Australia is the holder of four mineral licenses (ML) MLN 1070, MLN 1071, MLN 1127, and MLN 31525 comprising approximately 5,544 hectares (ha). In addition, Vista Australia controls exploration licenses (EL) EL 29882, EL 29886, EL 30898, and EL 28321 comprising approximately 153,700 ha. Figure 4‑2 illustrates the general location of the tenements and the position of the Batman deposit. A general arrangement is provided in Figure 4‑3.

4.3Lease and Royalty Structure

Vista Australia entered into a lease agreement (the Lease Agreement) with the NT government for an initial term of five years commencing January 1, 2006, with an extension of five years at Vista Australia’s option and three additional years upon the application of Vista Australia and with the approval of the NT government. Pursuant to the conditions of the first five-year term of the Lease Agreement, Vista Australia undertook a comprehensive technical and environmental review of the Project to evaluate site environmental conditions and developed a program to stabilize the environmental conditions and minimize offsite contamination. Vista also reviewed the water management plan and made recommendations and developed a Technical Report for the re-starting of operations. During the term of the Lease Agreement, Vista Australia was also required to examine all technical, economic, and environmental issues, estimate the cost to rehabilitate the site, explore and evaluate the potential of the Project, and prepare a technical and economic feasibility study for the potential development of the Project site.

Vista provided notice to the NT government in June 2010 that it wished to extend the Lease Agreement. In November 2010, the NT government granted the renewal and the Lease Agreement was extended for an additional five years to December 31, 2015. The NT government renewed the Lease Agreement by deed of variation in 2014 and again in May 2017, extending it to December 31, 2023.

Vista Australia paid the NT government's costs of management and operation of the Project Site up to a maximum of AUD375,000 during the first year of the term, and assumed site management and management and operation costs in the following years. In the agreement, the NT government acknowledges its commitment to rehabilitate the site and the Lease Agreement provides that Vista Australia has no rehabilitation obligations for pre-existing environmental conditions until it submits and receives approval of a Mining Management Plan (MMP) for the resumption of mining operations. The most recent MMP (Vista Gold Australia 2016) addresses activities undertaken by Vista with respect to site management, infrastructure maintenance and environmental management (Section 20.3 – Permitting and Authorizations).

Recognizing the importance placed by the NT government upon local industry participation, Vista Australia has agreed to use, where appropriate, NT-sourced labor and services during the period of the Lease

Agreement in connection with the Mt Todd property, and further, in connection with any proposed mining activities prepare and execute a local Industry Participation Plan.

Pursuant to an agreement (the JAAC Agreement) with the Jawoyn Association Aboriginal Corporation (JAAC), Vista was required to issue Vista common shares with a value of Canadian dollars (CAD) 1.0 million as consideration for the JAAC entering into the JAAC Agreement and as rent for the use of the surface lands overlying the mineral leases during the period from the effective date of the agreement until a decision is reached to begin production. Vista pays the JAAC AUD5,000 per month in return for consulting with respect to aboriginal, cultural, and heritage issues.

Figure 4‑1: General Project Location Map

NOTE: Prepared by Vista Gold Corp.; updated on February 23, 2018

Figure 4‑2: Concessions

Figure 4‑3: General Arrangement

4.4Risks

Vista is in sole possession of the title and rights to perform work on the Project. Surface access is guaranteed through Vista’s agreement with the JAAC. Exploration or other similar activities require a MMP to be submitted to the Department of Regional Development, Primary Industry, Fisheries and Resources (DRDPIFR) with approvals typically occurring in thirty or less days. Vista has been in sole possession of the site for the last nearly 11 years and no MMPs have ever been withheld with regard to exploration or other activities.

5.0ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1Accessibility

The Project is located 56 km by road northwest of Katherine, and approximately 290 km southeast of Darwin in the NT of Australia. Access to the mine is via high quality, two-lane paved roads from the Stuart Highway, the main artery within the territory.

5.2Climate and Physiography

The Project area has a sub-tropical climate with a distinct wet season and dry season. The area receives most of its rainfall between the months of January and early March. During these months, the temperature usually ranges from 25° to 35°C, but temperatures can reach as high as 42°C. Winter temperatures in the dry season usually range from 14°C to 20°C, but can drop to as low as 10°C at night.

Mining and processing operations are planned year-round; however, pit dewatering will be required after large precipitation events.

5.3Local Resources and Infrastructure

Access to local resources and infrastructure is excellent. The Project is located sufficiently close to the city of Katherine to allow for an easy commute for workers. The area has both historic and current mining activity and therefore a portion of the skilled workforce will be sourced locally. In addition, Katherine offers the necessary support functions that are found in a medium-sized city with regard to supplies, accommodations, communications, etc.

The property has an existing high-pressure gas line and an electric power line that was used by previous operators. In addition, wells for potable water and a dam for process water are also located on or adjacent to the site. Finally, a fully functioning tailings dam is present on site.

The concessions are within 2 to 3 km of the Nitmiluk Aboriginal National Park on the east. This National Park contains a number of culturally and geologically significant attractions. The proximity to the National Park has not historically yielded any impediments to operating. It is not expected to yield any issues to renewed operation of the property in the future. The Project is wholly contained within the Aboriginal Freehold Land and will require no additional acquisition of surface rights.

5.4Topography, Elevation and Vegetation

The topography of the Project is relatively flat. The mineral leases encompass a variety of habitats forming part of the northern Savannah woodland region, which is characterized by eucalypt woodland with tropical grass understories. Surface elevations are on the order of 130 to 160 meters (m) above sea level in the area of the previous and planned site and waste dumps.

6.0HISTORY

The Project area has significant gold deposits. It is situated in a well-mineralized historical mining district that supported small gold and tin operations in the past.

The Shell Company of Australia (Billiton), who was the managing partner in an exploration program in joint venture with Zapopan NL (Zapopan), discovered the Mt Todd mineralization, or more specifically the Batman deposit, in May 1988. Zapopan acquired Billiton’s interest in 1992 by way of placement of shares to Pegasus Gold Australia Pty. Ltd. (Pegasus). Pegasus progressively increased their shareholding until they acquired full ownership of Zapopan in July 1995.

Feasibility studies (not NI 43-101 compliant) for Phase I, a heap leach operation which focused predominately on the oxide portion of the deposit, commenced during 1992 culminating in an engineering, procurement, construction management (EPCM) award to Minproc in November of that year. The Phase I project was predicated upon a 4 million tonnes per year (Mtpy) on an annualized basis heap leach pad, which came on stream in late 1993. The treatment rate was subsequently expanded to a rate of 6 Mtpy on an annualized basis in late 1994.

Historic production is shown in Table 6‑1.

Table 6‑1: Heap Leach – Historic Actual Production

Category | Historic Production Actual
Tonnes Leached (million) | 13.2
Head Grade (g-Au/t) | 0.96
Recovery (%) | 53.8
Gold Recovered (oz) | 220,755
Cost/t (AUD) | 8.33
Cost/oz (AUD) | 500

NOTE: All tonnages and grades are historic production numbers that pre-date Vista’s ownership. The QPs and issuer consider historic estimates to be relevant but not current.

Phase II involved expanding to 8 Mtpy and treatment through a flotation and carbon-in-leach (CIL) circuit. The feasibility study was conducted by a joint venture between Bateman Kinhill and Kilborne (BKK, 1996) and was completed in June 1995.

The Pegasus board approved the project on August 17, 1995, and awarded an EPCM contract to BKK in October 1995. Commissioning commenced in November 1996. Final capital cost to complete the project were AUD232 million (US$181 million).

Design capacity was never achieved due to inadequacies in the crushing circuit. An annualized throughput rate of just under 7 Mtpy was achieved by mid-1997; however, problems with the flotation circuit which resulted in reduced recoveries necessitated closure of this circuit. Subsequently, high reagent consumption as a result of cyanide soluble copper minerals further hindered efforts to reach design production. Operating costs were above those predicted in the feasibility study.

The spot price of gold deteriorated from above US$400 in early 1996 to below US$300 per ounce during 1997. According to the 1997 Pegasus Annual Report, the economics of the project were seriously affected by the slump. Underperformance of the project and higher operating costs led to the mine being closed and placed on care and maintenance on November 14, 1997.

In February 1999, General Gold Resources Pty. Ltd. (General Gold) agreed to form a joint venture with Multiplex Resources Pty Ltd (Multiplex Resources) and Pegasus to own, operate, and explore the mine. Initial equity participation in the joint venture was General Gold 2%, Multiplex Resources 93%, and Pegasus 5%. The joint venture appointed General Gold as mine operator, which contributed the operating plan in exchange for a 50% share of the net cash flow generated by the project, after allowing for acquisition costs and environmental sinking fund contributions. General Gold operated the mine from March 1999 to July 2000.

6.1History of Previous Exploration

The Batman gold prospect is part of a goldfield that was worked from early in the 20 th century. Gold and tin were discovered in the Mt Todd area in 1889. Most deposits were worked in the period from 1902 to 1914. A total of 7.80 tonnes of tin concentrate was obtained from cassiterite-bearing quartz-kaolin lodes at the Morris and Shamrock mines. The Jones Brothers reef was the most extensively mined gold-bearing quartz vein, with a recorded production of 28.45 kg Au. This reef consists of a steeply dipping ferruginous quartz lode within tightly folded greywackes.

The Yinberrie Wolfram field, discovered in 1913, is located 5 km west of Mt Todd. Tungsten, molybdenum and bismuth mineralization was discovered in greisenized aplite dykes and quartz veins in a small stock of the Cullen Batholith. Recorded production from numerous shallow shafts is 163 tonnes of tungsten, 130 kg of molybdenite and a small quantity of bismuth.

Exploration for uranium began in the 1950s. Small uranium prospects were discovered in sheared or greisenized portions of the Cullen Batholith in the vicinity of the Edith River. The area has been explored previously by Esso for uranium without any economic success.

Australian Ores and Minerals Limited (AOM) in joint venture with Wandaroo Mining Corporation and Esso Standard Oil took out a number of mining leases in the Mt Todd area during 1975. Initial exploration consisted of stream sediment sampling, rock chip sampling, and geological reconnaissance for a variety of commodities. A number of geochemical anomalies were found primarily in the vicinity of old workings.

Follow-up work concentrated on alluvial tin and, later, auriferous reefs. Backhoe trenching, costeaning, and ground follow-up were the favored mode of exploration. Two diamond drillholes were drilled at Quigleys. Despite determining that the gold potential of the reefs in the area was promising, AOM ceased work around Mt Todd. The Arafura Mining Corporation, CRA Exploration, and Marriaz Pty Ltd all explored the Mt Todd area at different times between 1975 and 1983. In late 1981, CRA Exploration conducted grid surveys, geological mapping and a 14-diamond drillhole program, with an aggregate meterage of 676.5 m, to test the gold content of Quigleys Reef over a strike length of 800 m. Following this program CRA Exploration did not proceed with further exploration.

During late 1986, Pacific Gold Mines NL (Pacific Gold Mines) undertook exploration in the area which resulted in small-scale open cut mining on the Quigleys and Golf reefs, and limited test mining at the Alpha, Bravo, Charlie and Delta pits. Ore was carted to a carbon-in-pulp (CIP) plant owned by Pacific Gold Mines at Moline. This continued until December 1987. Pacific Gold Mines ceased operations in the area in February 1988 having produced approximately 86,000 tonnes grading 4 g-Au/t (historic reported production, not NI 43-101

compliant). Subsequent negotiations between the Mt Todd Joint Venture partners (Billiton and Zapopan) and Pacific Gold Mines resulted in the acquisition of this ground and incorporation into the joint venture.

Table 6‑2 presents important historical events in a chronologic order.

Table 6‑2: Property History

1986 |
October 1986 – January 1987: | Conceptual Studies, Australia Gold PTY LTD (Billiton); Regional Screening (Higgins); Ground Acquisition, Zapopan N.L.
1987 |
February: June-July: October: | Joint Venture finalized between Zapopan and Billiton. Geological Reconnaissance, Regional BCL, stream sediment sampling. Follow-up BCL stream sediment sampling, rock chip sampling and geological mapping (Geonorth)
1988 |
Feb-March: March-April: May: May-June: July: July-Dec: | Data reassessment (Truelove) Gridding, BCL grid soil sampling, grid based rock chip sampling and geological mapping (Truelove) Percussion drilling Batman (Truelove) - (BP1-17, 1475m percussion) Follow-up BCL soil and rock chip sampling (Ruxton, Mackay) Percussion drilling Robin (Truelove, Mackay) – RP 1-14, (1584m percussion) Batman diamond, percussion and reverse circulation (RC) drilling (Kenny, Wegmann, Fuccenecco) - BP18-70, (6263m percussion); BD1-71, (8562m Diamond); BP71-100, (3065m R.C.)
1989 |
Feb-June: June: July-Dec: | Batman diamond and RC drilling: BD72-85 (5060m diamond); BP101-208, (8072m RC). Penguin, Regatta, Golf, Tollis Reef Exploration Drilling: PP1-8, PD1, RGP1-32 GP1-8, BP108, TP1-7 (202m diamond, 3090m RC); TR1-159 (501m RAB). Mining lease application (MLA's 1070, 1071) lodged. Resource estimates; mining-related studies; Batman EM-drilling: BD12, BD8690 (1375m diamond); RC pre-collars and H/W drilling, BP209-220 (1320m RC); Exploration EM and exploration drilling: Tollis, Quigleys, TP9, TD1, QP1-3, QD1-4 (1141 diamond, 278m RC); Negative Exploration Tailings Dam: E1-16 (318m RC); DR1-144 (701. RAB) (Kenny, Wegmann, Fuccenecco, Gibbs).
1990 |
Jan-March: | Pre-feasibility (PFS) related studies; Batman Inclined Infill RC drilling: BP222-239 (2370m RC); Tollis RC drilling, TP10-25 (1080m RC). (Kenny, Wegmann, Fuccenecco, Gibbs)
1993 - 1997 |
| Pegasus Gold Australia Pty Ltd reported investing more than $200 million in the development of the Mt Todd mine and operated it from 1993 to 1997, when the project closed as a result of technical difficulties and low gold prices. The deed administrators were appointed in 1997 and sold the mine in March 1999 to a joint venture comprised of Multiplex Resources Pty Ltd and General Gold Resources Ltd.
1999 - 2000 |
March - June | Operated by a joint venture comprised of Multiplex Resources Pty Ltd and General Gold Resources Ltd. Operations ceased in July 2000, Pegasus Gold Australia Pty Ltd., through the Deed Administrators, regained possession of various parts of the mine assets in order to recoup the balance of purchase price owed to it. Most of the equipment was sold in June 2001 and removed from the mine. The tailings facility and raw water facilities still remain at the site.

2000 - 2006 |
| The Deed Administrators, Pegasus Gold Australia Pty Ltd, the government of the NT, and the Jawoyn Association Aboriginal Corporation held the property.
2006 |
March | Vista Gold Corp. acquired mineral lease rights from the Deed Administrators.

6.2Historic Drilling

The following discussion centers on the historic drillhole databases that were provided to Tetra Tech for use in this Technical Report. Based on the reports by companies, individuals and other consultants, it is The QPs’ opinion that the drillhole databases used as the bases of this report contain all of the available data. Tetra Tech is unaware of any drillhole data that have been excluded from this report.

6.2.1Batman Deposit

There are 730 historic drillholes in the Batman deposit assay database. Figure 6‑1 shows the drillhole locations for the Batman deposit. These drillholes include 225-diamond drill core (DDH), 435 reverse circulation holes (RVC), and 70 open rotary holes (OP). Nearly all of the DDH and RVC holes were inclined 60° to the west. Samples were collected in one-meter intervals. DDH holes included both HQ and NQ core diameters. Core recoveries were reported to be very high with a mean of 98%. The central area of the deposit was extensively core-drilled. Outside of the central area, most of the drillholes were RVC and OP holes. All drillholes collars were surveyed by the mine surveyor. Down-hole surveys were conducted on most drillholes using an Eastman single shot instrument. All drillholes were logged on site.

A series of vertical RVC infill holes were drilled on a 25 m x 25 m grid in the core of the deposit to depths between 50 m and 85 m below the surface. Zapopan elected to exclude these drillholes from modeling the Batman deposit because the assays from these drillholes seemed to be downwardly biased and more erratic compared to assays from inclined RVC holes. Of the possible reasons cited as to why vertical RVC holes might report lower grades and have a more erratic character, the 1992 Mining & Resource Technology Pty Ltd (Khosrowshahi et al. 1992 – MRT) report states that "the orientation of vertical holes sub-parallel to mineralization caused preferential sampling of barren host rocks...”. This statement was, at least in part, borne out by the later sampling work done on the blast holes as it was credited with part of the reproducibility problems that were encountered when the Batman deposit was being mined.

6.2.2Drillhole Density and Orientation

Pegasus was aware of the potential problem of drillhole density within the Batman deposit. The feasibility study prepared by BKK (BKK, 1996) indicates that the drilling density decreases with depth. In the central area oxide and transition zone spacing was generally 25 m by 25 m. The spacing was wider on the periphery of the mineralized envelope. The drilling density in the central area of the primary zone ranged from 50 m by 50 m, but decreased to 50 m by 100 m and greater at depth. At the time of that study, there were 593 drillholes in the assay database 531 of which RSG used in the construction of the MRT block model.

At the time of The Winters Company’s (TWC) site visit in 1997, the drillhole database numbered 730 drillholes. It is not known if any drillholes were excluded from the Pegasus exploration models. Most of the new drilling that had been added since the 1994 MRT model was relatively shallow. TWC reviewed PGA's 50 m drill

sections through the Batman deposit and saw that there was a marked decrease in drillhole spacing below 1,000 RL (the model has had constant 1,000 m added to it in order to prevent the reporting of elevations below 0 m and have been denoted as RL for relative elevation) and another sharp break below 900 RL. The drillhole spacing in the south of 1,000 N on the 954 RL bench plan approached 80 m x 80 m. Pegasus was able resolve this problem by using very long search ranges in its grade estimation. In the main ore zone, Pegasus used maximum search distances in the north and east directions of nearly 300 m.

Another potential problem related to drilling is the preferred orientation of the drillholes. Most of the drillholes in the assay database are inclined to the west to capture the vein set which strikes N10° to 20°E, dips east, and which dominates the mineralized envelope. This orientation is the obvious choice to most geologists since these veins are by far the most abundant. Ormsby (1996) discussed that while the majority of mineralization occurs in these veins, the distribution of gold mineralization higher than 0.4 g-Au/t is controlled by structures in other orientations, such as east-west joints and bedding. For this reason, Ormsby stated, "[t]he result is that few ore boundaries (in the geological model) actually occur in the most common vein orientation." If this is truly the case, the strongly preferential drilling orientation has not crosscut the best mineralization and in cases may be sub-parallel to it.

Vertically oriented RVC holes were not included in the drillhole database for the 1994 MRT model because their assay results appeared to be too low compared to other drillhole orientations. If vertical drillhole orientations were actually underestimating the gold content during exploration drilling, the vertical and often wet blast holes, which are used for ore control, pose a similar problem and will need to be addressed prior to commencing any new mining on the site.

6.2.3Quigleys

Table 6‑3 details the Quigleys exploration database as of the time of this report. Figure 6‑1 also shows the drillhole locations for the Quigleys deposit.

Table 6‑3: Summary of Quigleys Exploration Database

Category | Count | Min | Max | Average
Count | 644 | - | - | -
Depth | - | 13 | 368 | 92
Collar Easting | 644 | 187,067 | 190,023 | 189,484
Collar Northing | 644 | 8,437,020 | 8,439,305 | 8,438,149
Collar Elevation | 644 | 129 | 208 | 156
Survey Azimuth | 2,057 | 0 | 359 | 87.36
Survey Dip | 2,057 | -90 | -40 | -60
Assay Au | 54,073 | 0 | 36 | 0.241
Assay Interval | 54,131 | 0.1 | 69 | 1.04

Snowden (1990) completed a statistical study of the Quigleys drillhole database in order to bias test it. A comparison of historic and recent data by Snowden suggested that a bias might exist. Further study concluded that a bias is not apparent where all drilling is oriented in a similar direction (and not clustered). This suggests the inclusion of assay data from all phases of drilling is reasonable. The March 2008 report entitled “Mt Todd Gold Project, Gold Resource Update” contains additional information regarding the Snowden findings.

Figure 6‑1: Drillhole Location Map – Batman and Quigleys Deposits

6.3Historic Sampling Method and Approach

NQ core intervals were sawed lengthwise into half core. HQ core was quartered. RVC samples were riffle split on site and a 3- to 4-kg sample was sent to an assay lab. The 1992 MRT resource report commented that many of the RVC holes were drilled wet and that Billiton and Zapopan were aware of possible contamination problems. Oddly, in some comparison tests, DDH had averaged assays five percent to six percent higher than RVC holes; for that reason, MRT elected to exclude RVC holes from the drillhole database for grade estimation of the central area of the Batman deposit.

Since the property is currently not operating, Tetra Tech did not witness any drilling and sampling personally. We have taken the following discussion from reports by the various operators and more importantly, from reports by independent consultants that were retained throughout the history of the property to audit and verify the sampling and assaying procedures. It is the opinion of the QP for this section that the reports by the various companies and consultants have fairly represented the sampling and assaying history at the site and that the procedures implemented by the operators, most notably General Gold, have resulted in an assay database that fairly represents the tenor of the mineralization at Batman.

6.4Historic Sample Preparation, Analysis and Security

The large number of campaigns and labs used in the Mt Todd drilling effort has resulted in a relatively complex sampling and assaying history. The database developed prior to August of 1992 was subjected to a review by Billiton, and has been subjected to extensive check assays throughout the project life. Furthermore, a number of consultants have reviewed the integrity of the database and have been content with the data for modeling purposes.

Drillhole samples were taken on one-meter intervals, though there are instances of two-meter intervals in the typically barren outlying drillholes. The procedure involved sawing the NQ core lengthwise in half. HQ core was quartered. RVC samples were riffle split on site and a 3- to 4-kg sample was sent to the laboratory for analyses. Pincock Allen and Holt (PAH) stated that they witnessed the sample preparation process at a number of steps and concurred with the methods in use (PAH, 1995).

Pegasus (and Zapopan, before) conducted a check assay program which is consistent with industry practice. Every 20th assay sample was subjected to assay by an independent lab. Standards were run periodically as well, using a non-coded sample number to prevent inadvertent bias in the labs.

6.4.1Sample Analysis

According to reports by Pegasus, various consultants, and others, the early exploration assays were largely done at various commercial labs in Pine Creek Geosyncline (PCG) and Darwin. Later assays were done at the Mt Todd mine site lab. At least three different sample preparation procedures were used at one time or another. All fire assays were conducted on 50-gram charges. Based on these reports, it appears that the assay labs did use their own internal assay blanks, standards, and blind duplicates.

Assay laboratories used for gold analysis of the Batman drill data were Classic Comlabs in Darwin, Australia, Assay Laboratories in Pine Creek and Alice Springs and Pegasus site Laboratory.

The exploration data consist of 91,225 samples with an average and median length of 1 m. The minimum sample length is 0.1 m and the maximum sample length is 5 m. 137 samples are less than 1 m and 65 samples are over 1 m in length.

All exploration drill data were used for the resource estimate. Four-meter down hole composite samples were calculated down hole for the resource estimate. The assay composited data were tabulated in the database field called “Comp”. The weighted average grades, the length, and the drillhole were recorded.

6.4.2Check Assays

Extensive check assaying was carried out on the exploration data. Approximately 5% of all RVC rejects were sent as duplicates and duplicate pulps were analyzed for 2.5% of all DDH intervals. Duplicate halves of 130 core intervals were analyzed as well. Overall, Mt Todd's check assay work is systematic and acceptable. The feasibility study showed that the precision of field duplicates of RVC samples is poor and that high errors exist in the database. The 1995 study stressed that because of the problems with the RVC assays, the RVC and OP assays should be kept in a separate database from the DDH assays (PAH, 1995). However, since that time, the majority of the identified assaying issues have been corrected by General Gold based on recommendations of consultants. It is the opinion of the QP responsible for this section that the assay database used in the creation of the current independent resource estimation exercise is acceptable and meets industry standards for accuracy and reliability.

6.4.3Security

The QP responsible for this section is unaware of any “special” or additional security measures that were in place and/or followed by the various exploration companies, other than the normal practices of retaining photographs, core splits, and/or pulps of the samples sent to a commercial assay laboratory.

6.5Historic Process Description

The Mt Todd deposit is a large, but low-grade gold deposit. The average grade of the gold mineralization is approximately 1 g-Au/t. The gold mineralization occurs in a hard, uniform greywacke host and is associated with sulfide and silica mineralization which has resulted from deposition along planes of weakness that had opened in the host rock. Gold is very fine grained (<30 microns) and occurs with both silica and sulfides. The host rock is very competent with a Bond Ball Mill Work Index (BWi) of 23 to 30.

Pegasus and earlier owners did extensive metallurgical testing from 1988 to 1995 to develop a process flowsheet for recovering gold from low-grade extremely hard rock. The treatment route, based on the metallurgical studies, was engineered to provide for the recovery of a sulfide flotation concentrate which was subsequently reground and leached in a concentrate leach circuit. Flotation tailings were leached in a separate CIL circuit.

The historic design process flowsheet for the Project is given in Figure 6‑2.

A brief description of the major unit operations is as follows:

Crushing: Four stages of crushing were employed to produce a product having a P80 of 2.6mm. The primary crusher was a gyratory followed by secondary cone crushers in closed circuit. Barmac vertical shaft impact crushers were used for tertiary crushing in closed circuit and quaternary crushing stages. The crushed product was stored under a covered fine ore stockpile.

Grinding: The crushed product was drawn from the fine ore stockpile into three parallel grinding circuits, each consisting of an overflow ball mill in closed circuit with cyclones to produce a grind with a P80 of 150 microns.

Flotation: Cyclone overflow was sent to the flotation circuit where a bulk concentrate was supposed to recover seven percent of the feed with 65% to 70% of the gold.

CIL of Tailing: The flotation tailing was leached in carbon-in-leach circuit. The leach residue was sent to the tailings pond. Approximately 60% of the gold in the flotation tailings was supposed to be recovered in the CIL circuit.

CIL of Flotation Concentrate: The flotation concentrate was reground in Tower mills to 15 microns and subjected to cyanide leaching to recover the bulk of the gold in this product (94.5% of the flotation concentrate). The leach residue was sent to the tailings pond.

Process Recycle: The process water was recycled to the milling circuit from the tailings pond. The overall gold recovery was projected to be 83.8% for the proposed circuit. However, during the initial phase of plant optimization, problems were encountered with high levels of cyanide in the recycled process water which, when returned to the mill, caused depression of pyrite and much lower recoveries to the flotation concentrate. As a result, the flotation plant was shut down and the ground ore was directly sent to the CIL circuit. The modified process flowsheet is given in Figure 6‑3.

Without the flotation circuit, the CIL plant recovered 72 to 75% of the gold.

The plant was shut down and placed on care and maintenance within one year of startup due to a collapse in gold price, under performance of the process plant and higher than projected operating costs.

6.6Technical Problems with Historical Process Flowsheet

There were several technical problems associated with the design flowsheet. These technical problems have been documented by plant engineers, TWC, and other investigators. They are briefly discussed in this section.

6.6.1Crushing

The four-stage crushing circuit was supposed to produce a product with P80 of 2.6mm. Also, historically the tonnage was projected to be 8 Mtpy on an annualized basis. The actual product achieved in the plant had a P80 of 3.2 to 3.5 mm and the circuit could handle a maximum of 7 Mtpy on an annualized basis. This resulted in an increased operating cost for gold production.

A four-stage crushing/ball mill circuit was selected over a SAG/ball mill/crusher circuit because crushers were available from the Phase I heap leach pad and could be used in the Phase II program. The use of this available equipment did reduce the overall capital cost.

Figure 6‑2: Plant Process Flowsheet for Project as Designed

Figure 6‑3: Modified Plant Process Flowsheet for Project

The following problems were encountered with the crushing circuit:

The mechanical availability of the Barmac vertical shaft impact crushers was extremely poor.

The Barmac crushers were not necessarily the best choice for the application. The three-stage crusher product could have been sent to the mills which would have had to have been larger size mills.

The crushing circuit generated extreme amounts of fines and created environmental problems. The dust also carried gold with it. The dust levels increased the wear on machinery parts and were a potential long-term health hazard.

The use of water spray to keep the dust down resulted in use of large amounts of fresh water. This was a strain on the availability of fresh water for the plant.

General Gold operated a whole-ore cyanide leach facility but no technical reports describing their process have been located by Vista to date.

6.6.2Flotation Circuit

The flotation circuit was supposed to recover 60 to 70% of the gold in a bulk sulfide concentrate which was 7% of the feed material. The flotation circuit recovered ± 1% of the weight of material and less than 50% of the gold values. This was due to the significant amount of cyanide in the recycle process water which depressed the sulfide minerals in the flotation process. If the process water had been detoxified, the problems would not have occurred. This was not done because of the cost associated with a cyanide detoxification plant.

Additional problems which were overlooked during the testwork and design of the plant included the following:

The presence of cyanide soluble copper was known but was not taken into consideration during the design of the process flowsheet; and

Removal of copper from the bulk sulfide in the form of a copper concentrate would have reduced the consumption of cyanide as well as the amount of weak acid dissociable (WAD) cyanide in the recycled process water. Pilot plant testing was undertaken in the plant to produce copper concentrate. Documented results do indicate ± 60% of copper recovery at a concentrate grade of +10% Cu. Approximately 45% of the gold reported to this concentrate. However, from our discussions with the engineering contractors and the Pegasus staff running the pilot plant, a copper concentrate assaying over 20% was achieved in some of the later tests.

6.6.3CIL of Flotation Concentrate and Tailings

A portion of the copper was depressed with cyanide with the recycled process water in the flotation process. Hence, the cyanide consumption was high even in the leaching of the flotation tailings. The availability of dissolved oxygen in leaching terms was very low thereby resulting in poor extraction of gold in the leach circuit. This resulted in an estimated reduction of 40% of gold recovery in the circuit.

7.0GEOLOGICAL SETTING AND MINERALIZATION

7.1Geological and Structural Setting

The Project is situated within the southeastern portion of the Early Proterozoic Pine Creek Geosyncline (Figure 7‑1). Meta-sediments, granitoids, basic intrusives, acid and intermediate volcanic rocks occur within this geological province.

Within the Mt Todd region, the oldest outcropping rocks are assigned to the Burrell Creek Formation. These rocks consist primarily of interbedded greywackes, siltstones, and shales of turbidite affinity, which are interspersed with minor volcanics. The sedimentary sequence incorporates slump structures, flute casts and graded beds, as well as occasional crossbeds. The Burrell Creek Formation is overlain by interbedded greywackes, mudstones, tuffs, minor conglomerates, mafic to intermediate volcanics and banded ironstone of the Tollis Formation. The Burrell Creek Formation and Tollis Formation comprise the Finniss River Group.

The Finniss River Group strata have been folded about northerly trending F1 fold axes. The folds are closed to open style and have moderately westerly dipping axial planes with some sections being overturned. A later north-south compression event resulted in east-west trending open style upright D2 folds.

The Finniss River Group has been regionally metamorphosed to lower green schist facies.

Late and Post Orogenic granitoid intrusion of the Cullen Batholith occurred from 1,789 Ma to 1,730 Ma, and brought about local contact metamorphism to hornblende hornfels facies.

Unconformably overlying the Burrell Creek Formation are sandstones, shales and tuffaceous sediments of the Phillips Creek sandstone, with acid and minor basic volcanics of the Plum Tree Creek Volcanics. Both these units form part of the Edith River Group, and occur to the south of the Project Area.

Relatively flat lying and undeformed sediments of the Lower Proterozoic Katherine River Group unconformably overlie the older rock units. The basal Kombolgie Formation forms a major escarpment, which dominates the topography to the east of the Project area.

7.2Local Geology

The geology of the Batman deposit consists of a sequence of hornfelsed interbedded greywackes, and shales with minor thin beds of felsic tuff. Bedding is striking consistently at 325°, dipping at 40° to 60° to the southwest. Minor lamprophyre dykes trending north-south pinch and swell, cross cutting the bedding.

Nineteen lithological units have been identified within the deposit and are listed in Table 7‑1 below from south to north (oldest to youngest).

Table 7‑1: Geologic Codes and Lithologic Units

Unit Code | Lithology | Description
1 | GW25 | Greywacke
2 | SH24 | Shale
3 | GW24A | Greywacke
4 | SHGW24A | shale/greywacke
5 | GW24 | Greywacke
6 | SHGW23 | shale/greywacke
7 | GWSH23 | greywacke/shale
8 | GW23 | Greywacke
9 | SH22 | Shale
10 | T21 | felsic tuff
11 | SH21 | Shale
12 | T20 | felsic tuff
13 | SH20 | Shale
14 | GWSH20 | greywacke/shale
15 | SH19 | Shale
16 | T18 | felsic tuff
17 | SH18 | Shale
18 | GW18 | Greywacke
Int | INT | lamprophyre dyke

Bedding parallel shears are present in some of the shale horizons (especially in units SHGW23, GWSH23 and SH22). These bedding shears are identified by quartz/ calcite sulfidic breccias. Pyrite, pyrrhotite, chalcopyrite, galena and sphalerite are the main primary sulfides associated with the bedding parallel shears.

East west trending faults and joint sets crosscut bedding. Only minor movement has been observed on these faults. Calcite veining is sometimes associated with these faults. These structures appear to be post mineralization.

Northerly trending quartz sulfide veins and joints striking at 0° to 20°, dipping to the east at 60° are the major location for mineralization in the Batman deposit. The veins are 1 millimeter (mm) to 100 mm in thickness with an average thickness of around 8 mm to 10 mm. The veins consist of dominantly quartz with sulfides on the margins. The veining occurs in sheets with up to 20 veins per horizontal m. These sheet veins are the main source of mineralization in the Batman deposit.

Figure 7‑1: General Geologic Map

7.3Mineralization

A variety of mineralization styles occur within the Mt Todd area. Of greatest known economic significance are auriferous quartz-sulfide vein systems. These vein systems include the Batman, Jones, Golf, Quigleys and Horseshoe prospects, which occur within a north-northeast trending corridor, and are hosted by the Burrell Creek Formation. Tin occurs in a north-northwest trending corridor. The tin mineralization comprises cassiterite, quartz, tourmaline, kaolin, and hematite bearing assemblages, which occur as bedding to parallel breccia zones and pipes. Polymetallic Au, W, Mo, and Cu mineralization occurs in quartz-greisen veins within the Yinberrie Leucogranite; a late stage highly fractionated phase of the Cullen Batholith. The Batman Deposit extends approximately 2,200 m along strike, 400 m across dip and drill tested to a depth of 800 m. Drilling indicates the Batman mineralization to be open along-strike and down-dip.

7.3.1Batman Deposit

7.3.1.1Local Mineralization Controls

The mineralization within the Batman deposit is directly related to the intensity of the north-south trending quartz sulfide veining. The lithological units impact on the orientation and intensity of mineralization.

Sulfide minerals associated with the gold mineralization are pyrite, pyrrhotite and lesser amounts of chalcopyrite, bismuthinite and arsenopyrite. Galena and sphalerite are also present, but appear to be post-gold mineralization, and are related to calcite veining in the bedding plains and the east-west trending faults and joints.

Two main styles of mineralization have been identified in the Batman deposit. These are the north-south trending vein mineralization and bedding parallel mineralization.

7.3.1.2North-South Trending Corridor

The north-south trending mineralization occurs in all rock units and is most dominant in the shales and greywackes designated SHGW23. Inspection of grade control and exploration data, drill logs, diamond core and the pit has shown that the north-south trending mineralization can be divided into three major zones based on veining and jointing intensity.

CORE COMPLEX

Mineralization is consistent and most, to all, joints have been filled with quartz and sulfides. Vein frequency per meter is high in this zone. This zone occurs in all rock types.

HANGING WALL ZONE

Mineralization is patchier than the core complex due to quartz veining not being as abundant as the core complex. The lithology controls the amount of mineralization within the hanging wall zone. The hanging wall zone doesn’t occur north of T21. South of reference line T21 to the greywacke shale unit designated GWSH23, the mineralization has a bedding trend. A large quartz/pyrrhotite vein defines the boundary of the hanging wall and core complex in places.

FOOTWALL ZONE

Like the hanging wall zone, the mineralization is patchier than the core complex and jointing is more prevalent than quartz veining. Footwall Zone mineralization style is controlled by the lithology and occurs in all lithological units.

Narrow bands of north-south trending mineralization also occur outside the three zones, but these bands are patchy.

BEDDING PARALLEL MINERALIZATION

Bedding parallel mineralization occurs in rock types SH22 to SH20 to the east of the Core complex. Veining is both bedding parallel and north south trending. The mineralization appears to have migrated from the south along narrow north-south trending zones and “balloon out” parallel to bedding around the felsic tuffs.

7.3.2Quigleys Deposit

The Quigleys deposit mineralization was interpreted by Pegasus and confirmed by Snowden (1990) to have a distinctive high-grade shallow dipping 30°-35° northwest shear zone extending for nearly 1 km in strike and 230 m vertical depth within a zone of more erratic lower grade mineralization. The area has been investigated by RVC and diamond drilling by Pegasus and previous explorers on 50 m lines with some infill to 25 m.

Drillhole intersections generally revealed an abrupt change from less than 0.4 g-Au/t to high grade (>1 g‑Au/t) mineralization at the hanging wall position of the logged shear, but also revealed a gradational change to lower grade mineralization with depth. Some adjacent drillholes were also noted with significant variation in the interpreted position of the shear zone, and some of the discrepancies appeared to have been resolved on the basis of selection of the highest gold grade. While the above method may result in a valid starting point for geological interpretation, the selection of such a narrow high grade zone is overly restrictive for interpretation of mineralization continuity and will require additional work prior to estimating any resources.

It was further thought that while the shear might be readily identified in diamond drillholes, interpretation in RVC drilling, and in particular later interpretation from previously omitted RVC holes, must invoke a degree of uncertainty in the interpretation. Snowden concluded that while the shear zone was identifiable on a broad scale, the local variation was difficult to map with confidence and therefore difficult to estimate with any degree of certainty at this time.

It is for these reasons that Vista has only drilled diamond drillholes. As reference above, the shears and other structural features are identifiable in drill core.

NOTE: Prepared by Vista Gold Corp.; updated on February 23, 2018

Figure 7‑2: Concessions

8.0DEPOSIT TYPES

According to Hein (2003), the Batman and Quigleys gold deposits of the Mt Todd Mine are formed by hydrothermal activity, concomitant with retrograde contact metamorphism and associated deformation, during cooling and crystallization of the Tennysons Leucogranite and early in D2 (Hein, submitted for publication). It is speculated that pluton cooling resulted in the development of effective tensile stresses that dilated and/or reactivated structures generated during pluton emplacement and/ or during D1 (Furlong et al., 1991, as cited in Hein, 2003), or which fractured the country rock carapace as is typical during cooling of shallowly emplaced plutons (Knapp and Norton, 1981, as cited in Hein, 2003). In particular, this model invokes sinistral reactivation of a northeasterly trending channelization basement strike–slip fault, causing brittle failure in the upper crust and/or dilation of existing north–northeasterly trending faults, fractures, and joints in competent rock units such as meta-greywackes and siltstones. The generation of dilatant structures above the basement structure (i.e., along a northeasterly trending corridor overlying the basement fault), coupled with a sudden reduction in pressure, and concomitant to brecciation by hydraulic implosion (Sibson, 1987; Je´brak, 1997; both as cited in Hein, 2003) may have facilitated channelization of predominantly metamorphic fluid in the intermediate contact metamorphic aureole (possibly suprahydrostatic-pressured) and into the upper crust (Furlong et al., 1991; Cox et al., 2001; both as cited in Hein, 2003). Rising fluids decompressed concurrent with mineral precipitation. Throttling of the conduit or fluid pathways probably resulted in over pressuring of the fluid (Sibson, 2001, as cited in Hein, 2003), this giving way to further fracturing, etc. Mineral precipitation accompanied a decrease in temperature although, ultimately, the hydrothermal system cooled as isotherms collapsed about the cooling pluton (Knapp and Norton, 1981).

Gold mineralization is constrained to a single mineralizing event that included:

Retrogressive contact metamorphism during cooling and crystallization of the Tennysons Leucogranite;

Fracturing of the country rock carapace;

Sinistral reactivation of a NE-trending basement strike-slip fault;

Brittle failure and fluid-assisted brecciation; and

Channelization of predominantly metamorphic fluid in the intermediate contact metamorphic aureole into dilatant structures.

The deposits are similar to other gold deposits of the porphyry copper gold (PCG) and are classified as orogenic gold deposits in the subdivision of thermal aureole gold style. The Batman deposit shares some characteristics with intrusion-related gold systems, especially in terms of the association of gold with bismuth and reduced ore mineralogies. This makes the deposit unique in the PCG.

The mineral deposit types being investigated and the geological model being applied are described in Section 9.0 – Exploration and Section 14.0 – Mineral Resource Estimates, respectively.

9.0EXPLORATION

Since acquiring the Mt Todd mining leases and exploration licenses, Vista has conducted an ongoing exploration program that includes prospecting, geologic mapping, rock and soil sampling, geophysical surveys and exploration drilling. Equipment and personnel were mobilized from the Mt Todd Mine site or from an exploration base camp established in the central part of the exploration licenses. The work was conducted by geologists and field technicians.

The exploration effort initially focused on follow up work on targets developed by Pegasus during their tenure on the property. These included the RKD target, Tablelands, and Silver Spray. During a review of Pegasus’ airborne geophysical survey data, five distinct magnetic highs were observed located within sedimentary rocks that should have a low magnetic signature. These features are remarkably similar to those at the Batman deposit, which, as a result of the included pyrrhotite, exhibits a strong magnetic high. The geophysical targets were prioritized following review of historic work in the area and site visits. To date, two of the geophysical targets (Golden Eye and Snowdrop) have been grilled and a third has been covered by soil sampling (Black Hill).

Table 9‑1 details soil geochemical samples collected on the exploration licenses (ELs) by year.

Table 9‑1: Exploration Sampling

Year | Soils | Samples Collected
2008 | 0 | 164
2009 | 1,333 | 45
2010 | 3,135 | 224
2011 | 1,925 | 79
2012 | 2,312 | 295
2013 | 572 | 51
2014 | 2,601 | 143
2015 | 841 | 53
2016 | 241 | 27
2017 | 1,098 | 78
Total Samples | 11,831 | 949

Table 9‑2: Exploration Prospects

Year | Drill Hole | Location | Zone | GDA94 Coords | Tasks Completed?
Prospect | Lease No | Easting | Northing | RL | Depth | Rehab Status
2010
| GE10-001 | Goldeneye | EL29886 | 53L | 200220 | 8455415 | 184 | 252 | Closed
| GE10-002 | Goldeneye | EL29886 | 53L | 200360 | 8455415 | 178 | 297 | Closed
| GE10-003 | Goldeneye | EL29886 | 53L | 200340 | 8455495 | 189 | 194 | Closed
| GE10-004 | Goldeneye | EL29886 | 53L | 200190 | 8455495 | 189 | 194 | Closed
| RKD10-001 | RKD | EL29882 | 53L | 197400 | 8450650 | 201 | 201 | Closed
| RKD10-002 | RKD | EL29882 | 53L | 197440 | 8450550 | 225 | 225 | Closed
| RKD10-003 | RKD | EL29882 | 53L | 197440 | 8450550 | 291 | 291 | Closed
| RKD10-004 | RKD | EL29882 | 53L | 197400 | 8450520 | 336 | 336 | Closed
| RKD10-005 | RKD | EL29882 | 53L | 197530 | 8450450 | 183 | 183 | Closed
| RKD10-006 | RKD | EL29882 | 53L | 197360 | 8450490 | 552 | 352 | Closed
2011
| SS11-001 | Silver Spray | EL29882 | 53L | 208572 | 8460026 | 217 | 369 | Closed
| SS11-002 | Silver Spray | EL29882 | 53L | 208607 | 8459933 | 211 | 438 | Closed
| LL11-001 | Limestone Quarry | EL28321 | 52L | 813950 | 8426350 | 95 | 60 | Closed
| LL11-002 | Limestone Quarry | EL28321 | 52L | 813950 | 8426300 | 95 | 60 | Closed
| LL11-003 | Limestone Quarry | EL28321 | 52L | 813950 | 8426250 | 95 | 60 | Closed
| LL11-004 | Limestone Quarry | EL28321 | 52L | 814050 | 8426350 | 95 | 64 | Closed
| LL11-005 | Limestone Quarry | EL28321 | 52L | 814050 | 8426300 | 95 | 61 | Closed
| LL11-006 | Limestone Quarry | EL28321 | 52L | 814050 | 8426250 | 95 | 60 | Closed
| GE11-001 | Goldeneye | EL29886 | 53L | 200300 | 8455555 | 177 | 195 | Closed
| GE11-002 | Goldeneye | EL29886 | 53L | 200240 | 8455455 | 182 | 351 | Closed
| GE11-003 | Goldeneye | EL29886 | 53L | 200350 | 8455455 | 182 | 241 | Closed
| GE11-004 | Goldeneye | EL29886 | 53L | 200400 | 8455500 | 186 | 267 | Closed
| GE11-005 | Goldeneye | EL29886 | 53L | 200400 | 8455555 | 186 | 240 | Closed
2012
| SD12-001 | Snowdrop | EL29882 | 53L | 195169 | 8457484 | 171 | 219 | Closed
2015
| SD15-001 | Snowdrop | EL29882 | 53L | 195164 | 8457302 | 170 | 250 | Closed
| SD15-002 | Snowdrop | EL29882 | 53L | 195142 | 8457248 | 170 | 250 | Closed
| SD15-003 | Snowdrop | EL29882 | 53L | 195305 | 8457599 | 170 | 250 | Closed
| WD15-001 | Wandie | EL29882 | 53L | 190947 | 8455709 | 169 | 46 | Closed
| WD15-002 | Wandie | EL29883 | 53L | 190920 | 8455696 | 168 | 100 | Closed
| WD15-003 | Wandie | EL29884 | 53L | 190890 | 8455679 | 167 | 135 | Closed
2016
| WD16-001 | Wandie | EL29882 | 53L | 190859 | 8455663 | 166 | 204 | Closed
|  |  |  |  |  |  |  | 6,445 |

9.1Golden Eye Target

At Golden Eye, an initial 100m x 100m soil program identified 2 anomalous samples, one of 70PPB and one of 50ppb, follow-up rock chip sampling, in an area with limited exposure, returned a 25.0 g-Au/t sample from a small outcrop of Laminated Fe rich sediments. Further sampling returned 23.0 g-Au/t and 7.7 g-Au/t assays in vein and breccias located 15 m and 50 m, respectively, north of the original sample. Due to the sparse outcrop, the orientation and thickness of the mineralized zone is not currently known. An infill soil sampling program over the area was completed on a 20 m grid. The survey returned a strong coherent gold anomaly approximately 400 m in diameter with coincident anomalous base metals and arsenic.

In 2010 Vista completed four drillholes on the target. All four drillholes intersected strong sulfide mineralization associated with laminated Fe rich Burrell Creek Formation with interesting concentrations of copper, lead zinc and anomalous gold mineralization, with the best intercept occurring in drillhole GE10-003 and consisting of 1.1 m of 7.69 g-Au/t including 0.3 m of 26.7 g-Au/t.

Five additional drillholes were completed during the 2011 field season. Drilling intersected several narrow weakly mineralized zones; however, none that can yet be correlated with any confidence between different drillholes or between the drillholes and the mineralization identified on the surface. The most encouraging mineralization was intersected by GE11-002, consisting of a sheared, chloritic and broken sulfide-rich unit from 54.2 m to 55 m which assayed 1.41 g-Au/t and a siliceous lode from 162.07m to 162.82 m which assayed 1.86 g-Au/t. The remaining drillholes all intersected widespread quartz sulfide veining containing pyrrhotite, chalcopyrite, and arsenopyrite and contained anomalous gold, copper, bismuth, and arsenic. Although thin and patchy, this mineralization is at least a clear indication that there is a mineralized system at Golden Eye which is yet to be defined with confidence.

A detailed ground magnetic survey was completed over the area in 2012 and an airborne UTS geophysical survey was conducted in 2013. One IP line was conducted in 2017 to determine if a more extensive program would be helpful, this defined a thin target zone. The survey results, combined with detailed mapping and the drillhole data is currently being reviewed and additional drilling may be recommended.

9.2RKD Target

Six drillholes totaling 1,587.4 m were completed on the target known as RKD during 2011. The drillholes intersected a NNW trending mineralized shear zone dipping steeply to the west. The best gold intercept was in drillhole RKD11-003 which contained 2.7 m of 2.3 g-Au/t. Drillhole RKD11-005 intersected 3 m of 3.4% copper and 50 ppm silver a chalcocite-rich part of the shear zone. All of the drillholes intersected anomalous gold with values up to 0.4 and 0.5 g-Au/t. Extensive surface mapping and rock-chip sampling indicates that RKD is likely to be thin and is strike constrained.

9.3Silver Spray Target

Two drillholes totaling 806.8 m were completed at Silver Spray. The drillholes intersected strong chloritic alteration throughout both drillholes. Both drillholes intersected several 20-m zones of strong quartz veining with a thin (30 cm) zone of galena, pyrrhotite and arsenopyrite. These zones contained anomalous lead, zinc, and arsenic but only sporadic anomalous gold (up to 0.18 g-Au/t).

9.4Snowdrop Target

In 2011, 100m x 100m soil geochemical lines were completed across the Snowdrop magnetic anomaly. These soils were later closed in on a 20-m spacing. The results confirmed and refined the a gold-copper-arsenic-bismuth anomaly with 146 samples of 481 samples containing 100 ppm or greater copper and 60 samples containing greater than 5 ppb gold (high value 97 ppb). The onset of the wet season has suspended work on the target until next spring. A drill plan will be included in the updated mine management plan to permit drilling in 2012.

In 2012, the detailed 20 m by 20 m infill soil sampling program was continued. A total of 3,376 soils have been collected in the target area. Results show a coherent gold anomaly that is 200-m wide and at least 700‑m long. It is oriented NE-SW and flanks a strong magnetic high. There is a strong correlation with As, Bi and Fe with zoned Cu and Zn on the margins. Rock chip sampling in the area has identified the highest grades within gossanous rocks associated with quartz float. Rock chip samples range up to 6 ppm.

In late November, 2012, a single diamond drillhole was completed on the target before the onset of the wet season. SD12-01 was drilled at an angle across the target zone to a depth of 219.1m. The drillhole intersected zones of intensely silicified greywackes and shales with minor sheeted quartz veins. The alteration and veining is notably similar to that observed at the Batman deposit in the vicinity of the core zone. The greywacke units are coarser grained than at Batman, but the frequency of lithological changes and alteration types are all very similar. Sulfides are present within the quartz veining and as disseminated blebs within intensely silicified siltstones. Common sulfide minerals include pyrite, pyrrhotite, chalcopyrite, and arsenopyrite with traces of galena, sphalerite and bornite. Veining has a steep dip to the east, similar to Batman, but appears richer in base metals. Disseminated sulfides are also more abundant, while the vein density is not as intense as Batman. Although the drillhole did not intersect significant ore grade mineralization, assay results were encouraging and additional drilling is warranted. The highest grade intercept was 0.90 g-Au/t with six intervals returning greater than 0.4 g-Au/t. In total, 80 intervals out of 272 samples contained detectable gold with two intervals greater than 30 m containing detectable gold. Two geochemical signatures are apparent in the assay data; one with gold associated with anomalous base metals and one with an association with As, Bi, Co, and Te.

To date, this early-stage exploration program has not produced an announceable discovery on the ELs. While the work is promising and will be ongoing, there are no quantifiable resources or reserves. Once an announceable discovery is made, Vista will detail that discovery according to all applicable disclosure regulations.

10.0DRILLING

10.1Drilling

The drilling discussed in this section is limited to the Batman deposit since the filing of the NI 43-101 Technical Report – Mt Todd Gold Project, 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia, Amended & Restated; July 7, 2014. These drillhole data were used to complete the resource estimate in August 2017 as presented in this Technical Report. All of the drillhole data presented in the aforementioned 2014 Technical Report are still valid and have been incorporated as presented therein.

Between the fourth quarter of 2012 and the end of the first quarter of 2017, the Vista exploration program at the Batman deposit consisted of 22 diamond core drillholes containing 12,530 m that targeted both infill definitional drilling and step-out drilling. Table 10‑1 contains information for the 22 drillholes completed.

Figure 10‑1 is a plan map that details the locations of all exploration drillholes drilled at the Batman deposit up to and including VB17-003.

Table 10‑1: Batman Deposit Drillholes Added for Resource Update

Drillhole ID | Northing m (MGA94 z53) | Easting m (MGA94 z53) | Elevation (masl) | Bearing (°) | Dip (°) | Total Depth (m) | Drillhole Type
VB12-015 | 8434901.6 | 187446.7 | 144.4 | 268 | -55 | 745.85 | Diamond
VB12-016 | 8434703.6 | 187262.7 | 147.3 | 267 | -61 | 713.5 | Diamond
VB12-017 | 8435349.1 | 187391.2 | 150.8 | 277 | -61 | 833.28 | Diamond
VB12-018 | 8434849.2 | 187429.9 | 144.7 | 270 | -56 | 177 | Diamond
VB12-019 | 8434846.9 | 187429.4 | 144.8 | 269 | -61 | 731.8 | Diamond
VB12-020 | 8435852.4 | 187359.6 | 167.3 | 272 | -67 | 611.9 | Diamond
VB12-021 | 8435954.0 | 187378.8 | 149.9 | 271 | -65 | 602.9 | Diamond
VB12-022 | 8434453.4 | 187179.3 | 153.3 | 269 | -57 | 647.9 | Diamond
VB12-023 | 8435801.3 | 187371.0 | 161.3 | 265 | -60 | 650.88 | Diamond
VB12-024 | 8434482.1 | 187094.7 | 149.8 | 266 | -58 | 460.14 | Diamond
VB12-025 | 8435656.2 | 187344.7 | 158.6 | 261 | -60 | 650.63 | Diamond
VB12-026 | 8434393.4 | 187066.8 | 144.8 | 270 | -59 | 378.9 | Diamond
VB12-027 | 8435717.0 | 187259.7 | 169.8 | 291 | -54 | 434.75 | Diamond
VB15-001 | 187431 | 8434480 | 147 | 268.3 | -75.812 | 455.5 | Diamond
VB15-001W1 | 187431 | 8434480 | 147 | 268.3 | -75.812 | 831.8 | Diamond
VB15-001W2 | 187431 | 8434480 | 147 | 268.3 | -75.812 | 746 | Diamond
VB15-002 | 187277 | 8434703 | 147.268 | 266.07 | -76.19 | 446.3 | Diamond
VB15-002W1 | 187277 | 8434703 | 147.268 | 266.07 | -76.19 | 705 | Diamond
VB16-002* | 187195 | 8434849 | 134.84 | 328.6 | -64 | 485.7 | Diamond
VB17-001* | 187094 | 8435292 | 161.5 | 184.6 | -55 | 166.6 | Diamond
VB17-002* | 187194 | 8434848 | 134.84 | 330.6 | -64 | 485 | Diamond
VB17-003* | 187091 | 8435290 | 161.5 | 188.2 | -55 | 568.9 | Diamond

NOTE: Metallurgical drillholes not used in the resource estimation.

Figure 10‑1: Drillhole Location Map Batman Deposit to VB17-003

10.2Sampling

The sampling method and approach for drillholes completed between 2012 and 2017 was the same as has been used by Vista at the Batman Deposit for all of the Vista diamond drilling. The drill core, upon removal from the core barrel, is placed into plastic core boxes. The poly core boxes are transported to the sample preparation building where the core is marked, geologically logged, geotechnically logged, photographed, and sawn into halves. One-half is placed into sample bags as one meter sample lengths, and the other half retained for future reference. The only exception to this is when a portion of the remaining core has been flagged for use in the ongoing metallurgical testwork.

The bagged samples have sample tags placed both inside and on the outside of the sample bags. The individual samples are grouped into “lots” for submission to Northern Analytical Laboratories for preparation and analytical testing. All of this work was done under the supervision of a Vista geologist.

Neither Vista nor Tetra Tech are aware of any drilling, sampling, or assaying issues that would materially impact the accuracy or the results presented in this Technical Report.

The QP has observed the sampling, statistically tested the approach, confirmed quality control procedures employed, and quality assurance actions taken for the Project, and is of the opinion that the data accurately represent the nature and extent of the deposit.

11.0SAMPLE PREPARATION, ANALYSES AND SECURITY

The following section describes the sample preparation, analyses and security undertaken by Vista through the March 2018 resource update.

All of the sample preparation, sample analysis and sample security information presented below is the same and unchanged from Section 11 of the NI 43-101 Technical Report – Mt Todd Gold Project, 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia, Amended & Restated; July 7, 2014, with the exception that the same methodologies, practices, analytical preparation and security has been applied to the drillholes referenced in Section 10 of this Technical Report.

11.1Sample Preparation

The diamond drilling program was conducted under the supervision of the geologic staff composed of a chief geologist, several experienced geologists, and a core handling/cutting crew. The core handling crew was labor recruited locally.

Facilities for the core processing included an enclosed logging shed and a covered cutting and storage area that was fenced in. Both of these facilities were considered to be limited access areas and kept secured when work was not in progress.

The diamond drill core was boxed and stacked at the rig by the drill crews. Core was then picked up daily by members of the core cutting crew and transported directly into the logging shed.

Processing of the core included photographing, geotechnical and geologic logging, and marking the core for sampling. The nominal sample interval was one meter. When this process was completed, the core was moved into the core cutting/storage area where it was laid out for sampling. The core was laid out using the following procedures:

One meter depth intervals were marked out on the core by a member of the geologic staff;

Core orientation (bottom of core) was marked with a solid line when at least three orientation marks aligned and used for structural measurements. When orientation marks were insufficient an estimation orientation was indicated by a dashed line;

Geologic logging was then done by a member of the geologic staff. Assay intervals were selected at that time and a cut line marked on the core. The standard sample interval was one-m, with a minimum of 0.4 m and a maximum of 1.4 m;

Blind sample numbers were then assigned based on pre-labeled sample bags. Sample intervals were then indicated in the core tray at the appropriate locations;

Each core tray was photographed and restacked on pallets pending sample cutting and stored on site indefinitely; and

9,635 assays were added for the October 2012 resource update, an additional 7,601 assay intervals were added for the March 2013 resource, and 729 assay intervals were added for the 2017 model update.

The core was then cut using diamond saws with each interval placed in sample bags. At this time, the standards and blanks were also placed in plastic bags for inclusion in the shipment. A reference standard or a blank was inserted at a minimum ratio of 1 in 10 and at suspected high grade intervals additional blanks

sample were added. Standard reference material was sourced from Ore Research & Exploration Pty Ltd and provided in 60 g sealed packets. When a sequence of five samples was completed, they were placed in a shipping bag and closed with a zip tie. All of these samples were kept in the secure area until crated for shipping.

Samples were placed in crates for shipping with 100 samples per crate (20 shipping bags). The crates were stacked outside the core shed until picked up for transport.

11.2Sample Analyses

The following laboratories have been used for lab preparation, analyses, and check assays (Table 11‑1).

Table 11‑1: Assay and Preparation Laboratories

Laboratory | Address | Purpose | Abbreviation | Certifications
ALS | Minerals | 31 Denninup Way Malaga, WA 6090 | Main assay analyses | ALS | ISO:9001:2008 and ISO 17025 Certified
ALS | Minerals | 13 Price St Alice Springs, NT 0870 | Sample Preparation | ALS Alice Springs | ISO 9001:2008 and ISO 17025 Certified
Genalysis Laboratory Services (Intertek Group) | 15 Davison St Maddington, WA 6109 | Check Analyses | Genalysis | Unable to verify
North Australian Laboratories Pty Ltd | MLN 792 Eleanor Rd Pine Creek, NT 0847 | Alternative assay analyses | NAL | ISO 10725 Certified
NT Environmental Laboratories (Intertek Group) | 3407 Export Dr Berrimah, NT 0828 | Check Analyses | NTEL | ISO 17025

Prior to the 2011 drilling campaign, the majority of samples were transported first to ALS in Alice Springs (NT) for sample preparation. After preparation, samples were then forwarded on to ALS in Malaga (WA) for assay analyses. One in every 20 pulp or reject was sent from ALS in Alice Springs to Northern Australian Laboratories (NAL), Vista was notified by email which samples were sent to NAL. For the 2011-2012 drilling campaign samples for assay were sent to NAL lab in Pine Creek, NT. Check assays on one in every 20 pulps or rejects were completed by NT Environmental Laboratories.

Following completion of assay results, all pulps and reject material was shipped back to the Project site and stored.

Vista is completely independent of any analytical testing entity presented in this Technical Report, other than they have engaged said entities as a customer.

11.3Sample Security

NAL is the primary laboratory for the current drilling program. The NAL laboratory is located in the town of Pine Creek, approximately 100 km distant by road. Samples were picked up and transported by NAL employees.

Sample shipments were scheduled for approximately once a week. The crates were picked up on site by NAL for direct road transport to the assay lab. A sample transmittal form was prepared and included with each shipment and a copy was filed in the geologist office on site.

When the shipment left site, sample transmittals were prepared and e-mailed to NAL. When the shipment arrived at the preparation facility the samples were lined out and a confirmation of sample receipt was e-mailed back to Vista.

The QP is satisfied with the adequacy of sample preparation, security and analytical procedures employed by Vista given the fact that Vista has completed more than 50,000 m of core drilling in the Batman deposit, to verify the approximately 98,000 m of historic drilling. Statistical analysis of the various drilling populations and quality assurance/quality control (QA/QC) samples has not either identified or highlighted any reasons to not accept the data as representative of the tenor and grade of the mineralization estimated at the Batman deposit.

12.0DATA VERIFICATION

12.1Drill Core and Geologic Logs

Multiple site visits were performed by Mr. Rex Bryan, QP for the resource estimation portion of this Technical Report. During that visit, Tetra Tech found a comprehensive drillhole database comprised of drill core, photographs of the drill core, assay certificates and results, and geologic logs. All data were readily available for inspection and verification. In addition, most of the subsequent companies or their consultants that have examined the project have completed checks of the data and assay results. The author reviewed drill core, drill core logs and assay certificates and found a minimal number of errors (i.e., mislabeled intervals, number transpositions), which were corrected in development of the resource estimation. It is the opinion of the QP responsible for this section that the databases and associated data were of a high quality in nature and valid for use in mineral resource and reserve estimation.

The QP responsible for this section found no significant discrepancies with the existing drillhole geologic logs and is satisfied that the geologic logging, as provided for the development of the three-dimensional geologic models, fairly represents both the geologic and mineralogic conditions of each of the deposits that comprise the Project.

12.2Topography

The topographic map of the project area was delivered electronically in an AutoCAD® compatible format and represents the topography in half-meter accuracy. The native coordinate system of the topography is GDA94 / Map Grid of Australia (MGA) zone 53, and for this resource update and as the Project goes forward GDA94 / MGA zone 53 will be the used coordinate system. The surveyed drillhole collar coordinates, once translated to GDA94 / MGA zone 53 agree well with the topographic map; it is the opinion of the QP for this section that the current topographic map is accurate and accurately represents the topography of the project area. In addition, it is suitable for the development of the geologic models, mineral resource estimates, and mineral reserve estimates.

12.3Verification of Analytical Data

As part of the 2007 exploration program, an exercise to both verify the historic assay results and ensure that future analytical work meets current NI 43-101 standards for reporting of mineral resources was completed. This program consisted of two components; re-assaying of a portion of the historic drillholes, and assaying of the new core drillholes.

A multi-phase program evaluated the accuracy of gold assays generated by NAL on Mt Todd core samples. The test involved three phases including, 1) cross checking assay standards used in the program between NAL and ALS-Chemex, 2) preparing and assaying 30, one-m intervals of remaining half-core and detailed analysis of crushing and analytical performance between the two labs, and 3) screen sieve assay analysis of 45 coarse reject samples plus the 45 comparable remaining half core samples.

Analysis of the results from the two labs confirmed that finer material tends to be higher grade and that this fine material had been preferentially lost through the coarse-weave sample bags during storage and handling of the coarse reject samples. Vista now uses commercial polyester sample bags and loss of fines is no longer an issue. The test also showed good reproducibility between labs in all tests at grade ranges typical of the

deposit. Greater variance, which is not unexpected, showed up in the few samples assaying in the 5-20 g-Au/t range.

Figure 12‑1, Figure 12‑2, and Figure 12‑3 detail the results of the analytical check program that was completed on the 2007 exploration drillholes. The program was designed to check both internal laboratory accuracy and inter-laboratory accuracy. NAL was the primary laboratory for completion of the sample analyses. ALS-Chemex in Sydney, Australia performed the inter-laboratory analyses. As can be seen from the plots, the correlation coefficient was 0.997 for the re-splits of original assays, 0.992 for pulp repeats, and 0.986 for inter-laboratory analyses, respectively.

Figure 12‑1: NAL Resplit Analyses

Figure 12‑2: NAL Pulp Repeats

Figure 12‑3: Original Pulp Cross Lab Checks

12.3.1Latest Drilling Data Verification

For the March 2018 resource estimate, a detailed data verification procedure was undertaken by Tetra Tech which focused on the latest two drilling campaigns (VB12-015 through VB17-003 inclusive). This verification was accomplished by: reviewing the assay database received from Vista, comparing results with laboratory certificates received directly from the laboratory and reviewing results of the field QA/QC samples.

For the 13 drillholes from the 2012 exploration program, there were 7,601 intervals assayed. For the nine drillholes from the combined 2015-2017 exploration programs, there were 1,770 intervals. In addition to Au and other precious metals, most intervals had multi-element and environmental test results as well. Similar to previous work, the assay interval averaged one meter with a minimum interval of 0.4 m and a maximum interval of 1.4 m. No errors were noted in the assay data received other than selenium results for one drillhole that were erroneously entered. This was corrected by Vista. A spot-check of approximately 14% of the received database with laboratory certificates requested and received from NAL showed a 100% correct correlation of reported values.

Field QA/QC samples (those submitted with the drillhole samples to the laboratory) were also analyzed. Five standards (standard reference materials [SRMs]) were used by Vista with ranges of Au between 0.334 and 5.49 ppm of variable mineral/rock composition. Results of the SRMs were plotted as the relative difference to the average SRM certified Au concentration and are shown in Figure 12‑4. Of the 385 results, no drift was noted over time and all but four were within 10% of the certified value. Of the four that fell outside that range the highest offset was 13.8%. One value was clearly a mislabeled sample and when plotted with the assumed correct standard fell within the 10% range. Figure 12‑4 demonstrates the variance is greatest at lower Au concentrations and this is normally seen with most Au analytical data.

Field blanks were also reviewed and found to be acceptable. Of 388 blank results, six blanks had Au concentrations greater than detection limit of 0.01. The maximum value was 0.11 ppm. Again, no drift was noted in the data over time.

Because the current drilling campaign uses core, a regular program of field duplicates is not instituted at this time, but approximately 30% of samples have at least one replicate assay performed and an additional 3% of these have a second replicate assay. Replicates are taken from pulp when the primary sample is taken and run in the same analytical “batch.” Variability is highest at concentrations near detection limit, but overall trends are very good for the drillholes. Figure 12‑5 shows the first replicate value against the primary value by drillhole. Equally good correlation is seen for the second replicates against the original and against the first replicate value.

The author is of the opinion that the current field QA/QC program and results meet industry standards and that the assay database adequately reflects values reported from the laboratory and is suitable for use in mineral resource and mineral reserve estimation.

Figure 12‑4: Scatterplot of Relative Au Value to Certified Standard Reference Material Value

Figure 12‑5: Scatterplots (Log Scale) of Replicates by Drillhole

13.0MINERAL PROCESSING AND METALLURGICAL TESTING

This section reports on the work done to develop the understanding of the metallurgical characteristics of the remaining ore in the Batman deposit. This understanding contributes to the design of a technically effective and economically efficient gold recovery operation.

13.1Summary

Key conclusions drawn from the metallurgy studies to date are:

Mt Todd (and in particular the Batman deposit) ore is among the hardest and most competent ore types processed for mineral recovery. The most energy efficient comminution circuit has been determined to be the sequence of primary crushing, closed circuit secondary crushing, and closed circuit HPGR tertiary crushing and ore sorting, followed by two stages of grinding.

The ore is free-milling, is not preg-robbing, and is amenable to gold extraction by conventional cyanidation processes.

The ore has moderate cyanide consumption, determined to be 0.4 kg of sodium cyanide per tonne of ore. This is largely due to the presence of sulfides and cyanide consuming copper.

The ore requires a P80 grind of 60 µm and 30-hour leach residence time to achieve a nominal 86.7% gold recovery (86.2% net of solution loss) from ore with a pit head grade of 0.84 g‑Au/t.

13.2Historic Metallurgical Test Programs

The Mt Todd deposit is a large low-grade gold deposit. The average grade of the gold mineralization is approximately 1 g-Au/t. The gold mineralization occurs in a hard, uniform greywacke host and is associated with sulfide and silica mineralization which has resulted from deposition along planes of weakness that had opened in the host rock. Gold is fine grained (<30 µm) and occurs with both silica and sulfides. The host rock is very competent with a Bond Ball Mill Work Index (BWi) of 23 to 30 kWh/t.

A substantial body of knowledge has been accumulated for the metallurgy of the Mt Todd ore, some from the historical operation of the mine, but more importantly, detailed information has been developed from recent sampling of the remaining ore body. Observations are as follows:

1988-1997 metallurgical studies by previous owners (Pegasus) led to the design and construction of a treatment plant comprised of crushing, milling to a P80 of 150 µm, sulfide flotation, concentrate regrind and cyanidation, and, separate CIL cyanidation of flotation tailings. Operational efficiencies were lower than planned due to ore hardness, presence of cyanide-soluble copper minerals, and inefficient flotation performance resulting from the presence of free cyanide in the process water (from recycled tailings decant water). These process difficulties together with the collapse of the gold price led to the cessation of operations in November of 1997.

In 2006, Vista acquired the Project with the belief that each of these challenges could be overcome through the use of current technology, adequate metallurgical testing and higher gold prices. Vista’s consultant, Resource Development Inc. (RDi), completed a study using historical metallurgical data and test results from transition ore samples. RDi proposed a flowsheet consisting of crushing and grinding followed by rougher flotation to produce a sulfide concentrate containing 85% of the gold. Rougher tailings, substantially barren of gold

and sulfides, would be discarded to a benign tailings dam. Rougher concentrate would be reground to enable upgrading in a cleaner flotation circuit to produce a saleable copper concentrate containing 50% of the gold. Cleaner tailings would be cyanide leached in a CIL circuit for gold recovery. The cleaner tailings would be subjected to cyanide destruction and stored in a separate sulfide tailings dam.

The design incorporated energy efficient HPGR technology in the comminution circuit to handle the hard ore. These processing advantages combined with a higher gold price significantly improved the viability of the proposed operation. It then became necessary to confirm if the remaining ore had the same metallurgical characteristics as the historically processed ore.

In 2007/2008 two exploration drilling programs were completed focusing on the deeper ore beneath the existing Batman pit. The following composites/samples were prepared for RDi’s testwork conducted on the samples of the deeper Batman ore from the 2007/2008 drilling program:

Composite 1 – 1,200 kg composite sample made up from 2007 drill core. The composite consisted of samples from five drillholes selected to be representative of a cross section of the deposit. The head assay was 1.3 g-Au/t, 0.92% S and 447 ppm Cu. The sequential copper analysis indicated 80.4% of the copper in the sample was primary copper. The dominant sulfide in the sample was pyrrhotite.

Composite 2 – 140 kg composite sample made up from 2008 drill core. The head assay was 0.89 g-Au/t and 450 ppm Cu. The sequential copper analysis indicated 80.3% of the copper in the sample was primary copper. The dominant sulfide in the sample was pyrrhotite.

Drillhole 41 sample was sourced from the oxide and transitional zones (depth of 0–65 m). The head assay was 1.78 g-Au/t, 1.42% S, 448 ppm Cu.

The new cores were more representative of the remaining resource and samples were selected for confirmatory metallurgical test work. It was confirmed that the ore was extremely hard but it was not possible to repeat the flotation results previously achieved. The tests indicated that gold recovery into the rougher flotation concentrate was ± 80% at a grind P80 of 74µm but copper could not be upgraded to saleable concentrate grade of ± 20% Cu. The best results were ± 6% Cu using the same test procedure as employed for earlier core testing (2006).

Investigations revealed that the historical core tested in 2006 was transition zone material containing copper minerals predominantly as secondary copper which is known to be a major consumer of cyanide. The major sulfide mineral was pyrite. However, the 2007 and 2008 drill core had primary copper as predominant copper species and pyrrhotite as the major sulfide mineral. Pyrrhotite is known to float readily as compared to pyrite and is significantly more difficult to depress in the flotation process. It was difficult to selectively float copper minerals and produce a copper concentrate without the dilutive effect of pyrrhotite and other gangue minerals. Consequently flotation was dropped from the flow sheet and replaced with whole ore leach.

In 2010/2011 a confirmatory drilling campaign and metallurgical test program was conducted on the remaining Batman resource. The objective was to validate the findings of the 2007/2008 programs and to expand the level of understanding of variability of metallurgical performance within the Batman ore body. Samples used for the 2011 metallurgical testwork program were sourced from eight drillholes drilled

2010/2011. The drillholes were orientated to intersect the main Batman ore body beneath the existing pit and are representative of the ore within the PFS pit shell.

All samples from drillholes labeled VB11 were drilled in 2011, logged, packaged then shipped directly to the laboratory for processing. Drillholes labeled MHT were drilled and logged during 2010 and were stored in cold storage before being transported to the laboratory in 2011.

The test program was designed by Vista, supervised by Ausenco Limited (Ausenco), and executed by ALS Ammtec in Perth, Western Australia. There were a total of ninety-nine composited gold ore drill core intervals originating from the Project area. The metallurgical testwork included head analyses, crushing tests (HPGR and conventional crush), comminution testing, mineralogical analyses, leaching tests, cyanide detoxification and thickening and rheology testing. The test results confirmed that gold recovery by whole ore leach was the appropriate approach to process design.

Vista Gold had additional testwork undertaken in 2016 at RDi on the 2011 drilling samples. The test results indicate that the recovery was independent of the ore types but was somewhat dependent on the content of quartz in the ore. Also testing of the HPGR product indicated that the plus 5/8-in material had the potential to be treated by ore sorting to reject non-sulfide material. Since this was undertaken in small-scale tests, it provided incentive to undertake large scale tests to improve the process flowsheet and economics of gold production.

13.32017 Metallurgical Testwork

During January and February 2017 Vista completed drilling and logging of approximately 1,700 m of PQ (3.75 in diameter) core to obtain four 5-tonne bulk samples of ore representing different parts of the deposit. These composites were selected to represent both near-term and longer-term mining and are spatially located to provide variability both horizontally and vertically.

The primary objective of this phase of the test program was to perform sufficient metallurgical testwork to confirm the preferred process flowsheet developed during the last two years and associated reagent consumptions.

13.3.1HPGR Testing at Thyssen-Krupp Industries (TKI)

The four composite samples were sent to TKI (formerly Polysius) in Germany for the HPGR crushing component of the test program. The material was crushed in a one meter diameter HPGR unit. The material was subjected to a single pass through the HPGR and then screened on 16mm (5/8 inch) and each composite had the coarse fraction weighed and placed into a drum. The fine fraction was weighed and placed into several drums. The coarse fractions were sent to Tomra Sorting Solutions/Outotec for ore sorting.

The test protocol is given in Error! Reference source not found.. The weights of the plus and minus 16mm fractions for each composite are given in Table 13‑1.

Table 13‑1: Material Balance for HPGR Tests

Composite No. | Sample Weight, Kgs | HPGR Products %
+16 mm | -16 mm
1. | 4399.9 | 17.5 | 82.5
2. | 4977.7 | 17.8 | 82.2
3. | 4370.7 | 16.6 | 83.4
4. | 4317.3 | 18.7 | 81.3

13.3.2Tomra/Outotec Ore Sorting Testwork

Each plus 16mm sample was weighed at Tomra facility. Each composite was split into three parts. Each split sample was subjected to a two-step automated sorting test designed to separate the gold-bearing sulfide minerals and quartz veining from non-gold bearing waste material. The first step (XRT) sorts the material by measuring differences in density to target the gold-bearing sulfide material. Three different sensitivities (1%, 2% and 5%) were tested. The X-ray Transmission (XRT) material was then washed to remove the fines which could interfere with the laser ore sorting. The second step (laser) separates the gold bearing quartz veining material.

The test results, summarized in Table 13‑2, indicate the following:

Open-circuit HPGR produced approximately 18% of the feed as a plus 16mm fraction.

The ore sorting rejected approximately 10% of the run-of-mine feed as below cut-off grade material. Approximately 1.3% of the gold was rejected with the waste fraction.

Removal of waste resulted in approximately 8% improvement in estimated mill feed grade (average life-of-mine grade of 0.91 g/t Au compared to 0.84 g/t Au reserve grade).

Table 13‑2: Tomra Sorting Test Results

| Test | Product (XRT + Laser + Wash) | Final Reject | Head Grade of +16mm To Sorting
Wt | Au (g/mt) | Ag (g/mt) | Sulfur (%) | CN Soluble Cu (ppm) | Wt | Au (g/mt) | Ag (g/mt) | Sulfur (%) | CN Soluble Cu (ppm) | Wt | Au (g/mt) | Ag (g/mt)
Composite # 1 |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
XRT Sensitivity at 5% | 1.1 | 190.2 | 0.817 | 0.7 | 1.09 | 45.6 | 125.5 | 0.103 | 0.2 | 0.24 | 10.0 | 315.7 | 0.533 | 0.6 | 0.89 | 30.65
| Distribution | 60.2% | 92.3% | 83.7% | 87.3% | 87.0% | 39.8% | 7.7% | 16.3% | 12.7% | 13.0% | 100.0% | 100.0% | 100.0% | 100.0% | 100.0%
XRT Sensitivity at 2% | 2.1 | 101.5 | 0.541 | 1.0 | 1.12 | 38.9 | 118 | 0.110 | 0.4 | 0.23 | 11.2 | 219.53 | 0.309 | 0.9 | 0.83 | 23.21
| Distribution | 46.2% | 80.9% | 68.8% | 80.8% | 74.1% | 53.8% | 19.1% | 31.2% | 19.2% | 25.9% | 100.0% | 100.0% | 100.0% | 100.0% | 100.0%
XRT Sensitivity at 1% | 3.1 | 71.4 | 0.758 | 2.2 | 0.94 | 42.4 | 124.5 | 0.086 | 0.2 | 0.27 | 12.5 | 195.9 | 0.331 | 1.7 | 0.80 | 22.94
| Distribution | 36.4% | 83.5% | 86.0% | 66.7% | 65.4% | 63.6% | 16.5% | 14.0% | 33.3% | 34.6% | 100.0% | 100.0% | 100.0% | 100.0% | 100.0%
Composite # 2 |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
XRT Sensitivity at 5% | 4.1 | 193.2 | 0.365 | 2.1 | 0.73 | 27.8 | 117.5 | 0.075 | 0.2 | 0.23 | 6.5 | 310.7 | 0.255 | 1.5 | 0.55 | 19.25
| Distribution | 62.2% | 88.9% | 94.6% | 83.9% | 87.2% | 37.8% | 11.1% | 5.4% | 16.1% | 12.8% | 100.0% | 100.0% | 100.0% | 100.0% | 100.0%
XRT Sensitivity at 2% | 5.1 | 138.4 | 0.449 | 11.8 | 1.03 | 40.7 | 114.5 | 0.106 | 0.2 | 0.18 | 8.1 | 252.86 | 0.294 | 10.1 | 0.90 | 25.26
| Distribution | 54.7% | 83.6% | 98.6% | 87.3% | 85.5% | 45.3% | 16.4% | 1.4% | 12.7% | 14.5% | 100.0% | 100.0% | 100.0% | 100.0% | 100.0%
XRT Sensitivity at 1% | 6.1 | 132.9 | 0.566 | 35.7 | 0.86 | 32.4 | 151.5 | 0.185 | 0.2 | 0.22 | 10.7 | 284.4 | 0.363 | 23.6 | 0.45 | 20.23
| Distribution | 46.7% | 72.9% | 99.4% | 77.4% | 71.9% | 53.3% | 27.1% | 0.6% | 22.6% | 28.1% | 100.0% | 100.0% | 100.0% | 100.0% | 100.0%
Composite # 3 |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
XRT Sensitivity at 5% | 7.1 | 110.3 | 0.255 | 1.0 | 0.51 | 64.1 | 94 | 0.072 | 0.4 | 0.12 | 23.2 | 204.3 | 0.171 | 0.9 | 0.41 | 43.12
| Distribution | 54.0% | 80.6% | 75.0% | 83.4% | 75.2% | 46.0% | 19.4% | 25.0% | 16.6% | 24.8% | 100.0% | 100.0% | 100.0% | 100.0% | 100.0%
XRT Sensitivity at 2% | 8.1 | 106.4 | 0.570 | 3.6 | 0.64 | 94.7 | 139.5 | 0.233 | 0.4 | 0.13 | 36.7 | 245.87 | 0.379 | 2.7 | 0.62 | 59.23
| Distribution | 43.3% | 65.1% | 87.2% | 78.9% | 64.8% | 56.7% | 34.9% | 12.8% | 21.1% | 35.2% | 100.0% | 100.0% | 100.0% | 100.0% | 100.0%
XRT Sensitivity at 1% | 9.1 | 86.2 | 0.282 | 13.5 | 0.58 | 96.7 | 153.5 | 0.055 | 0.2 | 0.11 | 33.6 | 239.7 | 0.136 | 10.6 | 0.69 | 54.55
| Distribution | 36.0% | 74.2% | 97.4% | 74.8% | 60.5% | 64.0% | 25.8% | 2.6% | 25.2% | 39.5% | 100.0% | 100.0% | 100.0% | 100.0% | 100.0%

Composite # 4 |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
XRT Sensitivity at 5% | 10.1 | 148.0 | 0.901 | 1.4 | 0.99 | 43.0 | 98 | 0.192 | 0.4 | 0.23 | 18.9 | 246.0 | 0.619 | 1.3 | 0.88 | 32.67
| Distribution | 60.2% | 87.6% | 83.8% | 86.7% | 77.0% | 39.8% | 12.4% | 16.2% | 13.3% | 23.0% | 100.0% | 100.0% | 100.0% | 100.0% | 100.0%
XRT Sensitivity at 2% | 11.1 | 127.2 | 0.933 | 2.3 | 1.32 | 46.4 | 136 | 0.127 | 0.4 | 0.21 | 13.8 | 263.17 | 0.516 | 2.1 | 1.18 | 28.88
| Distribution | 48.3% | 87.3% | 84.3% | 85.5% | 75.2% | 51.7% | 12.7% | 15.7% | 14.5% | 24.8% | 100.0% | 100.0% | 100.0% | 100.0% | 100.0%
XRT Sensitivity at 1% | 12.1 | 112.9 | 1.005 | 17.0 | 1.67 | 44.3 | 161.5 | 0.113 | 0.4 | 0.26 | 9.9 | 274.4 | 0.480 | 15.3 | 1.70 | 23.61
| Distribution | 41.1% | 86.1% | 96.7% | 81.8% | 75.3% | 58.9% | 13.9% | 3.3% | 18.2% | 24.7% | 100.0% | 100.0% | 100.0% | 100.0% | 100.0%

13.3.3Preparation of Composites for Metallurgical Testwork

The HPGR product (minus 16mm) and the ore sorting products were weighed for each composite. This was followed by blending of the minus 16mm product and splitting using a cone and quarter process to obtain quarter portions of the material for each composite. The ore sorting product was proportioned into the split samples to prepare the composite samples, representing the product of crushing with sortation as feed to leach processing.

The composite samples were stage crushed to nominal 6 mesh. However, the required samples were split out at 3/4 inch material for abrasion testing. The minus 6 mesh material was thoroughly blended and split into 1 kg and 10 kg charges, and approximately half the material was stored in drums.

13.3.4Mineralogical Study

The four prepared composite samples were submitted for mineralogical study with emphasis on gold, silver, and speciation of pyrrhotite. Each sample was prepared as a standard polished thin section for study by transmitted/reflected light microscopy.

The highlights of the study indicate the following:

The mineralogy of the four composites was very similar.

Quartz was the primary phase in all samples and account for over 60% of the volume.

Quartz occurs as very fine mosaic grains (5 to 10 µm) or as angular to rounded grains in sizes from 5 to 125 µm. Some very course fragments of quartz up to several millimeters were also present in all samples.

The coarse quartz was commonly associated with coarse grain sulfides.

Other silicate minerals identified in the samples were mica, biotite, muscovite, chlorite and plagioclase feldspar.

Sulfide minerals represented 2% to 3% in each composite. Pyrite was common in all samples and occurred as euhedral cubes and anhydral grains (3 to 300 µm).

Pyrite concentration was highest in Composites 1 and 2. It was intermixed with marcasite and arsenopyrite.

Arsenopyrite was most prominent in Composite 3 with a grain size of up to 100 µm.

Other sulfide minerals present included chalcopyrite, sphalerite and galena.

Pyrrhotite was identified in all four composites. It was determined to have monoclinic structure.

Most of the gold grains identified were associated with pyrite and ranged in sizes from 3 to 28 µm.

No discrete silver minerals were identified in any of the composite samples.

13.3.5Head Analyses

The composite samples were submitted for head analysis. The test results are summarized in Table 13‑3, Table 13‑4, and Table 13‑5. The results indicate the following:

The samples assayed from 0.348 g/t Au to 0.760 g/t Au.

The total sulfur content ranged from 0.43% to 1.26%.

The copper valves ranged from 241 ppm to 467 ppm.

The samples contained significantly lower gold values than projected from the drilling data as shown in Table 13‑5.

Table 13‑3: Head Analyses of Composite Samples

Element | Composite
1 | 2 | 3 | 4
Au, g/t | 0.679 | 0.350 | 0.350 | 0.699
Assay 1 |  |  |  |
Assay 2 | 0.672 | 0.346 | - | 0.713
Average | 0.675 | 0.348 | 0.350 | 0.706
Ag, g/t | 1.6 | 3.7 | 1.2 | 0.8
STotal, % | 1.26 | 0.67 | 0.43 | 0.76

Table 13‑4: Whole Rock Analyses of Composite Samples

Element Percent | Composite
1 | 2 | 3
Al | 7.33 | 7.65 | 7.44 | 6.97
Ca | 0.33 | 0.32 | 0.17 | 0.37
Fe | 5.48 | 5.02 | 5.44 | 4.97
K | 3.59 | 3.63 | 3.03 | 3.06
Mg | 1.14 | 1.23 | 1.26 | 1.16
Na | 0.29 | 0.36 | 0.50 | 0.36
Ti | 0.19 | 0.21 | 0.20 | 0.22
ppm
As | 50 | 103 | 403 | 113
Ba | 579 | 622 | 574 | 548
Bi | <10 | <10 | <10 | <10
Cd | 8 | 9 | 7 | 7
Co | 21 | 22 | 22 | 18
Cr | 83 | 97 | 111 | 88
Cu | 467 | 285 | 241 | 384
Mn | 352 | 372 | 360 | 368
Mo | <175 | <1 | <1 | <1

Ni | 213 | 72 | 78 | 74
Pb | 17 | 81 | 302 | 222
Sr | 83 | 23 | 17 | 20
V | <10 | 86 | 100 | 92
W | 575 | 11 | <10 | <10
Zn |  | 240 | 392 | 421

Table 13‑5: Assayed vs. Projected Head Analyses

| g/t Au
Assayed | Projected
Composite 1 | 0.675 | 1.54
Composite 2 | 0.348 | 0.99
Composite 3 | 0.350 | 0.74
Composite 4 | 0.706 | 0.56

13.3.6Abrasion Indices

The samples were submitted for bond Abrasion index determination. The test results are summarized in Table 13‑6. The test results indicate that the material is low to moderately abrasive.

Table 13‑6: Abrasion Indices for the Various Composite Samples

Sample | Ai, g
Composite CC ¾ X ½ in | 0.1603
Composite 1 – 16mm | 0.2278
Composite 2 – 16mm | 0.1616
Composite 3 – 16mm | 0.2006
Composite 4 – 16mm | 0.2250

13.3.7Bond Ball Mill Work Indices

Bond ball mill work indices (BWi) were determined at a grind size of P80 of 100 mesh for the various products, namely HPGR, ore-sorting, composite samples and waste material. The results are summarized in Table 13‑7 and Table 13‑8.

The test results indicate the following:

The BWi for the plus 16mm sorted product was higher than the composite samples prepared from the crushed products. Hence, it is reasonable to conclude that the uncrushed material in the HPGR is harder than the crushed product.

The rejected plus 16mm material has a BWi harder than the composite sample and harder than the plus 16mm sorted product.

The BWi for the products ranged from 23.1 to 24.28. A BWi of 24.5 was selected for the design of the primary ball mill circuit.

Table 13‑7: Bond Ball Mill Work Indices for Composite Samples

Composite | BWi (kwh/mt)
1 | 23.10
2 | 24.41
3 | 23.79
4 | 24.48

Table 13‑8: Bond Ball Mill Work Indices for Ore Sorting Products and Wastes

No. | Composite | Sample | BWi (kwh/mt) | Average BWi
1 | 1 | 1.1 XRT Product | 23.0 |
2 | 1 | 2.1 XRT Product | 25.15 | 25.00
3 | 1 | 3.1 XRT Product | 25.98 |
4 | 2 | 4.1 XRT Product | 26.55 |
5 | 2 | 5.1 XRT Product | 26.91 | 26.63
6 | 2 | 6.1 XRT Product | 26.44 |
7 | 3 | 7.1 XRT Product | 24.54 |
8 | 3 | 8.1 XRT Product | 24.63 | 24.87
9 | 3 | 9.1 XRT Product | 25.44 |
10 | 4 | 10.1 XRT Product | 25.37 |
11 | 4 | 11.1 XRT Product | 25.89 | 25.62
12 | 4 | 12.1 XRT Product | 25.61 |
13 | 2 | 4.2 Laser Waste | 26.34 |
14 | 4 | 10.2 Laser Waste | 23.89 |
15 | Composite Sample (before HPGR | 25.01 |

13.3.8Leach Tests

Several series of leach tests were performed to evaluate the effect of grind size, leach pulp density, cyanide concentration and two-stage grind on the gold extraction and reagent consumption.

The test procedure consisted of grinding the ore to the desired particle size in a single stage or two stages as would be done in the plant and the ground pulp was transferred to a bottle. The pulp density was adjusted to the desired level and then the pH was adjusted to 11 with hydrated lime. The slurry was pre-aerated for 4 hours with 50ppm lead nitrate. Sodium cyanide was then added to a calculated level of cyanide concentration. The pH and cyanide concentration were determined at 6 and 24 hours and a sample of solution was taken and assayed for gold and silver. Activated carbon was added at 24 hours at a level of 20g/L. After 30 hours, the solution was measured to determine pH, free cyanide, and gold and silver content. The carbon was screened and dried. The slurry was filtered, washed and dried. The products were prepared and assayed for gold and silver.

The test results are summarized Table 13‑9 to Table 13‑13. The test results indicate the following:

The gold extraction is size dependent. The finer the grind size, the higher the gold extraction.

The gold extraction for average grade composites 1 and 4 were 82.8% to 87.6% at a P80 of 46 µm in a single-stage grind. However, for two-stage grind to P80 of 53 µm, the gold extraction improved from 86.4% to 89.7%.

The NaCN consumption in the two-stage grind tests was also lower by ± 20% as compared to single-stage grind.

The preliminary optimization study indicated that the leach circuit could potentially operate at higher pulp density (± 50% solids) and lower cyanide concentration (750 ppm initial concentration) without impacting gold extraction.

Table 13‑9: Gold Extraction vs. Grind Size for the Four Composites

Test No. | Composite | P80, mesh | Extraction % Au (30 hrs.) | Residue g/t Au | Cal. Head g/t Au | Consumption Kg/t
NaCN | Lime
1 | 1 | 200 | 84.4 | 0.12 | 0.75 | 0.515 | 3.782
2 | 1 | 200 | 84.9 | 0.10 | 0.68 | 0.512 | 3.000
3 | 1 | 230 | 85.1 | 0.10 | 0.65 | 0.471 | 3.351
4 | 1 | 230 | 85.4 | 0.10 | 0.66 | 0.514 | 2.987
5 | 1 | 325 | 85.1 | 0.10 | 0.66 | 0.516 | 3.578
6 | 1 | 325 | 87.6 | 0.10 | 0.77 | 0.515 | 3.446
13 | 2 | 200 | 77.0 | 0.10 | 0.42 | 0.336 | 3.743
14 | 2 | 200 | 76.4 | 0.10 | 0.44 | 0.393 | 3.460
15 | 2 | 230 | 77.3 | 0.10 | 0.45 | 0.393 | 3.533
16 | 2 | 230 | 75.1 | 0.11 | 0.44 | 0.394 | 3.493
17 | 2 | 325 | 68.3 | 0.16 | 0.50 | 0.453 | 3.631
18 | 2 | 325 | 75.5 | 0.12 | 0.48 | 0.453 | 3.678
19 | 3 | 200 | 65.2 | 0.10 | 0.30 | 0.456 | 4.554
20 | 3 | 200 | 64.0 | 0.10 | 0.27 | 0.397 | 4.545
21 | 3 | 230 | 66.5 | 0.10 | 0.29 | 0.454 | 4.555

Test No. | Composite | P80, mesh | Extraction % Au (30 hrs.) | Residue g/t Au | Cal. Head g/t Au | Consumption Kg/t
NaCN | Lime
22 | 3 | 230 | 69.8 | 0.09 | 0.29 | 0.396 | 4.678
23 | 3 | 325 | 69.8 | 0.08 | 0.27 | 0.454 | 4.700
24 | 3 | 325 | 70.0 | 0.08 | 0.27 | 0.454 | 4.632
7 | 4 | 200 | 80.0 | 0.13 | 0.65 | 0.551 | 3.237
8 | 4 | 200 | 79.7 | 0.14 | 0.71 | 0.516 | 2.992
9 | 4 | 230 | 81.8 | 0.14 | 0.75 | 0.576 | 2.980
10 | 4 | 230 | 82.9 | 0.12 | 0.72 | 0.513 | 3.008
11 | 4 | 325 | 82.8 | 0.12 | 0.72 | 0.575 | 3.458
12 | 4 | 325 | 84.1 | 0.10 | 0.66 | 0.576 | 2.939

NOTE: Lime Consumption was assumed to be the same as lime addition to the test.

Table 13‑10: Gold Extraction at P80 of 270 mesh (53µm) with Two-stage Grind

for the Four Composites

Test No. | Composite | Extraction % Au (30 hrs.) | Residue g/t Au | Cal. Head g/t Au | Consumption Kg/t
NaCN | Lime
25 | 1 | 86.6 | 0.09 | 0.67 | 0.393 | 4.972
26 | 1 | 86.2 | 0.09 | 0.67 | 0.336 | 4.866
27 | 2 | 85.8 | 0.06 | 0.44 | 0.398 | 4.446
28 | 2 | 85.2 | 0.07 | 0.44 | 0.458 | 4.529
29 | 3 | 80.1 | 0.06 | 0.31 | 0.514 | 4.773
30 | 3 | 80.5 | 0.06 | 0.32 | 0.513 | 4.930
31 | 4 | 86.1 | 0.10 | 0.69 | 0.392 | 4.521
3 2 | 4 | 86.4 | 0.09 | 0.68 | 0.397 | 4.501

NOTE: Lime Consumption was assumed to be the same as lime addition to the test.

Table 13‑11: Effect of Pulp Density and NaCN Concentration on Gold Extraction for Composite No. 1

at P80 of 270 mesh (53µm) with Two-stage Grinding

Test No. | Composite | NaCN g/t | Pulp Density % Solids | Extraction % Au (30 hrs.) | Residue g/t Au | Cal. Head g/t Au | Consumption Kg/t
NaCN | Lime
62 | 1 | 1.0 | 40 | 87.8 | 0.08 | 0.65 | 0.399 | 3.010
63 | 1 | 1.0 | 40 | 88.8 | 0.08 | 0.67 | 0.399 | 3.003
64 | 1 | 1.0 | 45 | 89.1 | 0.07 | 0.66 | 0.273 | 3.008
65 | 1 | 1.0 | 45 | 88.7 | 0.07 | 0.64 | 0.271 | 3.011
66 | 1 | 0.75 | 45 | 87.5 | 0.08 | 0.63 | 0.270 | 3.028
67 | 1 | 0.75 | 45 | 88.4 | 0.07 | 0.62 | 0.221 | 3.024

Test No. | Composite | NaCN g/t | Pulp Density % Solids | Extraction % Au (30 hrs.) | Residue g/t Au | Cal. Head g/t Au | Consumption Kg/t
NaCN | Lime
68 | 1 | 0.5 | 45 | 88.8 | 0.07 | 0.64 | 0.210 | 3.007
69 | 1 | 0.5 | 45 | 88.4 | 0.08 | 0.65 | 0.212 | 3.007
70 | 1 | 1.0 | 50 | 89.5 | 0.07 | 0.66 | 0.305 | 3.021
71 | 1 | 1.0 | 50 | 89.7 | 0.07 | 0.63 | 0.344 | 3.015

NOTE: Lime Consumption was assumed to be the same as lime addition to the test.

Table 13‑12: Effect of Pulp Density and NaCN Concentration on Gold Extraction for Composite No. 3

at P80 of 270 mesh (53µm) with Two-stage Grinding

Test No. | Composite | NaCN g/t | Pulp Density % Solids | Extraction % Au (30 hrs.) | Residue g/t Au | Cal. Head g/t Au | Consumption Kg/t
NaCN | Lime
82 | 3 | 1.0 | 40 | 84.7 | 0.04 | 0.25 | 0.460 | 3.011
83 | 3 | 1.0 | 40 | 84.9 | 0.04 | 0.25 | 0.272 | 3.011
84 | 3 | 1.0 | 45 | 84.7 | 0.04 | 0.25 | 0.271 | 3.010
85 | 3 | 1.0 | 45 | 84.8 | 0.04 | 0.25 | 0.372 | 3.010
86 | 3 | 0.75 | 45 | 83.2 | 0.04 | 0.24 | 0.269 | 3.017
87 | 3 | 0.75 | 45 | 86.3 | 0.03 | 0.25 | 0.322 | 3.010
88 | 3 | 0.50 | 45 | 83.8 | 0.04 | 0.25 | 0.211 | 3.011
89 | 3 | 0.50 | 45 | 84.4 | 0.04 | 0.24 | 0.211 | 3.016
90 | 3 | 1.0 | 50 | 85.0 | 0.04 | 0.25 | 0.347 | 3.011
91 | 1 | 1.0 | 50 | 84.9 | 0.04 | 0.25 | 0.346 | 3.011

NOTE: Lime Consumption was assumed to be the same as lime addition to the test.

Table 13‑13: Effect of Pulp Density and NaCN Concentration on Gold Extraction for Composite No. 4

at P80 of 270 mesh (53µm) with Two-stage Grinding

Test No. | Composite | NaCN g/t | Pulp Density % Solids | Extraction % Au (30 hrs.) | Residue g/t Au | Cal. Head g/t Au | Consumption Kg/t
NaCN | Lime
72 | 4 | 1.0 | 40 | 86.7 | 0.08 | 0.62 | 0.337 | 3.014
73 | 4 | 1.0 | 40 | 86.8 | 0.08 | 0.62 | 0.275 | 3.012
74 | 4 | 1.0 | 45 | 85.9 | 0.09 | 0.61 | 0.315 | 3.024
75 | 4 | 1.0 | 45 | 86.8 | 0.08 | 0.62 | 0.270 | 3.017
76 | 4 | 0.75 | 45 | 86.4 | 0.08 | 0.60 | 0.222 | 3.013
77 | 4 | 0.75 | 45 | 86.0 | 0.09 | 0.62 | 0.270 | 3.018
78 | 4 | 0.50 | 45 | 86.5 | 0.09 | 0.64 | 0.210 | 3.015
79 | 4 | 0.50 | 45 | 86.1 | 0.09 | 0.62 | 0.210 | 3.022
80 | 4 | 1.0 | 50 | 88.4 | 0.09 | 0.63 | 0.264 | 3.014
81 | 4 | 1.0 | 50 | 86.0 | 0.09 | 0.64 | 0.263 | 3.023

NOTE: Lime Consumption was assumed to be the same as lime addition to the test.

13.3.9Cyanide Destruction

The cyanide leach residue for composites No. 1 and No. 4 were subjected to cyanide destruction tests using the air/SO2 method. Approximately 1.5 liters of leach residue at 50% solids was agitated with sodium meta-bi-sulfite (SMBS) three times the stoichiometric amount of free cyanide and copper sulfate. Samples were taken every hour and free cyanide determined. Though no free cyanide was detected after one hour, the test was run for four hours.

The cyanide specification before and after destruction for the two tests are given in Table 13‑14. The test results indicate the following:

The air-SO2 process successfully reduced CNWAD to levels of <10 ppm.

There is sufficient dissolved copper in solution for precipitation of copper iron cyanide compounds in the earlier years of operation. Hence, addition of copper sulfate may not be needed.

One hour of detox residence time is sufficient for the process.

Table 13‑14: Cyanide Destruction Test Results

Forms of Cyanide ppm | Composite
1 | 4
Before | After | Before | After
Free | 600 | 6.3 | 590 | 4.0
Total | 587 | 3.6 | 615 | 2.2
WAD | 590 | 5.0 | 560 | 2.6

13.3.10Thickening Tests

Thickening tests on leach residue having a grind size of P80 of 53 µm generated in two-stages of grinding were performed for the four composites. The test results, given in Table 13‑15, indicate the following:

Approximately 8 g/t of high molecular weight low anionic acrylamide/sodium acrylate flocculant will be required for the settling of the slurry.

Unit area required to settle the slurry to 45% solids ranges from 0.044 to 0.182 m 2/mt/day.

The unit area requirements increase significantly if the desired underflow solids is 50%.

Table 13‑15: Unit Area Requirements for Thickener for Composite Samples

Composite | P80, µm | pH | Flocculent | Feed % Solids | Unit Area Required m 2/mt/day
40% | 45% | 50% | 55%
1 | 53 | 11 | 8 g/t DAF-10 | 25 | 0.031 | 0.044 | 0.164 | 2.41
2 | 53 | 11 | 8 g/t DAF-10 | 25 | 0.050 | 0.069 | 0.150 | 2.448
3 | 53 | 11 | 8 g/t DAF-10 | 25 | 0.042 | 0.081 | 0.191 | 2.436
4 | 53 | 11 | 8 g/t DAF-10 | 25 | 0.083 | 0.182 | 0.650 | 2.425

13.4Process Flowsheet

The process flowsheet given in Figure 13-2 and Figure 13-3 has significant advantages over the process flowsheet provided in the NI 43-101 Technical Report – Mt Todd Gold Project, 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia, Amended & Restated; July 7, 2014. The samples tested in the 2017 program are believed to be representative of the deposit as stated in Section 13.3. These composites were selected to represent both near term and longer term mining and spastically located to provide variability both horizontally and vertically. The recovery was estimated from the leach test results for the various composites given in Table 13‑9 to Table 13‑13 based on the proportion of each composite over the life of the mine. The only deleterious element in the deposit will be oxide and secondary copper which increases the cyanide consumption but will not impact gold extraction. Though the amount of oxide and secondary copper which decrease with depth, the cyanide consumption was not corrected for it. Therefore, the cyanide consumption is conservatively estimated. These results have not been compiled in the final metallurgical report which will be issued by RDi later this year. The significant changes tested in the recent testwork and incorporated in the process flowsheet include the following:

Ore sorting of the coarse HPGR product which rejects ± 10% of the feed as waste product.

Two-stage grinding which reduces the quantity of material that needs to be ground in the second stage.

These modifications have resulted in producing much finer product to the leach circuit. This has resulted in enhancing gold extraction by ± 4%.

Figure 13‑1: Conceptual Process Flowsheet for Mt Todd Ore (2/2)

14.0MINERAL RESOURCE ESTIMATES

14.1Introduction

The following sections summarize the thought process, procedures, and results of Tetra Tech’s independent estimate of the contained gold resources of the:

1)Batman Deposit,

2)The Quigleys Deposit, and

3)The Heap Leach Pad.

Only these three deposits currently have resource estimates classified in accordance with CIM Standards. Each of the mineral resources for the Batman and Quigleys deposits have been reported within a shell generated using Whittle TM’s 4-D Lerchs-Grossman algorithm. Mineral resources within such a shell are not mineral reserves and do not demonstrate economic viability.

It is the opinion of the QP for this section that the reported mineral resource classifications comply with current CIM definitions for each mineral class.

Geostatistics resource estimation and 3-D visualization was done with various mining software. The primary software used were MicroModel®, MicroMine®, Vulcan®, GemCom® and Whittle TM. Additional statistical analysis was done with Statistica® and Excel®.

Figure 14‑1 shows the relative locations of the three resource estimations for the Project. The Batman deposit is located approximately 500 meters west of the original plant site, the Quigleys deposit and the Heap Leach Pad are north and south of the existing tailings area respectively. Table 14‑1 summarizes the resources of each.

Cautionary statements regarding mineral resource estimates:

Mineral resources are not mineral reserves and do not have demonstrated economic viability. There is no certainty that all or any part of the mineral resources will be converted into mineral reserves. Inferred resources are that part of a mineral resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply but not verify geological and grade or quality continuity. It is reasonably expected that the majority of inferred mineral resources could be upgraded to indicated mineral resources with continued exploration

Figure 14‑1: Drillhole Location Map Batman & Quigleys Deposits and Heap Leach Pad

Table 14‑1: Summary of the Batman, Heap Leach Pad and Quiqleys Deposits

| Batman Deposit (August 2017) | Heap Leach Pad (May 2013) | Quigleys Deposit (August 2017)
| Tonnes (000s) | Grade (g/t) | Contained Ounces (000s) | Tonnes (000s) | Grade (g/t) | Contained Ounces (000s) | Tonnes (000s) | Grade (g/t) | Contained Ounces (000s)
Measured (M) | 77,725 | 0.88 | 2,191 | - | - | - | 457 | 1.27 | 19
Indicated (I) | 200, 112 | 0.80 | 5,169 | 13,354 | 0.54 | 232 | 5743 | 1.12 | 207
Measured & Indicated | 277,837 | 0.82 | 7,360 | 13,354 | 0.54 | 232 | 6,200 | 1.13 | 225
inferred (F) | 61,323 | 0.72 | 1,421 | - | - | - | 1,600 | 0.84 | 43

NOTES:

(1)Measured & indicated resources include proven and probable reserves.

(2)Batman and Quigleys resources are quoted at a 0.40g-Au/t cut-off grade. Heap Leach resources are the average grade of the heap, no cut-off applied.

(3)Batman: Resources constrained within a US$1,300/oz gold Whittle TM pit shell. Pit parameters: Mining Cost US$1.50/tonne, Milling Cost US$7.80/tonne processed, G&A Cost US$0.46/tonne processed, 50K TPD Ore, 355 Days/Yr., TPY Ore 17,750,000 TPY, G&A/Year 8,201 K US4, Au Recovery, Sulfide 85%, Transition 80%, Oxide 80%, 0.2g-Au/t minimum for resource shell. Selling Cost: US$/oz recovered US412.00.

(4)Quigleys: Resources constrained within a US$1200/oz gold Whittle TM pit shell. Pit parameters: Mining cost US$2.07/tonne, Milling Cost US$9.623/tonne processed, Sale Cost US$/oz US$15.18, Royalty 1% NPR (Jawoyn), Gold Recovery All Types, 70%.

(5)Differences in the table due to rounding are not considered material

(6)Rex Bryan of Tetra Tech is the qualified person responsible for the Statement of Mineral Resources for the Batman, Heap Leach Pad and Quiqleys deposits.

(7)Thomas Dyer of Mine Development Associates is the qualified person responsible for developing the resource Whittle TM pit shell for the Batman Deposit.

(8)The effective date of the Batman and Quigleys resource estimate is August 2017, the effective date of the Heap Leach resource is May 2013.

(9)Mineral resources that are not mineral reserves have no demonstrated economic viability and do not meet all relevant modifying factors.

14.2Geologic Modeling of the Batman Deposit (2017)

Gold mineralization in the Batman deposit at the Project occurs in sheeted veins within silicified greywackes/shales/siltstones. The Batman deposit strikes north-northeast and dips steeply to the east. Higher grade zones of the deposit plunge to the south. The core zone is approximately 200-250 meters wide and 1.5 km long, with several hanging wall structures providing additional width to the orebody. Mineralization is open at depth as well as along strike, although the intensity of mineralization weakens to the north and south along strike.

The Batman deposit contains 94% of the gold resources classified as measured and indicated within the Project. Only the Batman resources have been further converted to classified reserves of Proven and possible ore.

Over several drilling campaigns, the shape of the mineralized shear zone has been adjusted and resized to accommodate this new data. Deeper step-out drilling by Vista indicated that the lower footwall of the core complex was previously not drill tested. The additional drilling confirmed the previously indicated higher grade plunge of the core complex. The new data was used to re‑define the granite contact that constrains the lower footwall of the core complex. The granite contact is a mineral exclusionary zone and has been modeled as a triangulated surface, which can be seen in Figure 14‑2.

In addition to resizing the core complex wireframe solid, three structures paralleling the core complex to the east were also resized and constructed into wireframe solids and used for this resource estimate. The interpreted parallel structures represent an echoing of the main mineralization controls of the core complex nearer the surface and to the east. Wireframe solids for the parallel structures were interpreted on sections using Au mineralization, veining percentage, visual sulfide percentages, structural orientations and multi-element data. Deep drilling conducted in 2011 and through 2012 confirmed the existence of these structures and indicates a possible increasing definition and grade at depth.

The Batman Deposit resource was updated to reflect the increase in available data provided by drilling conducted in 2015 through 2017. A redefinition of the geometry of a granite contact reducing primarily inferred resources at depth. A Whittle TM pit further constrained the reported 2017 resources.

Figure 14‑2 is a schematic of domain designations and crucial parameters used in the resource model. The figure lists the resource classification codes, the rock codes, density assignments. Also schematically shown are the constraining surfaces for current topography, levels of oxidation, granite basement and the US$1,300/oz gold pit shell constrain reported resources.

Figure 14‑2: Schematic of Codes and Surface Designations (Looking North)

Figure 14‑3 shows a sectional view of the drillhole data as of June 2013 unchanged in 2017 due to the limited amount of new drilling. The predominate direction of Batman Deposit drilling is dipping at approximately 45‑degrees to the west. The figure shows the original 2013 and new 2017 granite contact surfaces. Yellow highlight shows the 2017 truncation of the core zone.

Figure 14‑3: Sectional View of Drillhole Data 8,434,803 mN (Looking North)

14.2.1Batman Deposit Density Data

Drillhole data through 2012 for a total of 16,373 samples were tested for bulk density (diamond core). These bulk densities were carried out on a 10 to 15 centimeters (cm) piece of core from a 1-m sample. Based on this work, the bulk densities applied to the resource model are presented in Table 14‑2.

Table 14‑2: Summary of Batman Bulk Density Data by Oxidation State

Oxidation | No. of Samples | Min | Max | Mean | Variance | CV
Oxide | 2,341 | 1.77 | 3.28 | 2.47 | 0.04 | 0.08
Transitional | 1,316 | 2.07 | 3.55 | 2.67 | 0.01 | 0.04
Primary | 12,716 | 1.58 | 3.90 | 2.77 | 0.006 | 0.03

Since then, an additional 3,370 samples have confirmed these results for Primary material bulk density.

14.2.2Grade Capping

Review of the log probability plot of the composited gold grades shows that there is a distinct break in the distribution at 50 g-Au/t. All gold composites were capped at this value. Inspection of the cumulative frequency plot of data from the core domain codes 600, 700, 800 and 1000 suggest that the 1m assay values when composited to 4 m limits the higher gold grades to a maximum value of 10.9 g-Au/t.

14.3Batman Block Model Parameters

Table 14‑3 details the physical limits of the Batman deposit block model utilized in the estimation of mineral resources.

Table 14‑3: Block Model* Physical Parameters – Batman Deposit

Direction (dir) | Minimum (m) MGA94 z53 | Maximum (m) MGA94 z53 | Block Size | #Blocks
y-dir | 8,433,801 mE | 8,436,213 mE | 12 m | 201
x-dir | 185,999 mN | 187,931 mN | 12 m | 161
z-dir | -994 m | 224 m | 6 m | 203

* Model changed from previous Tetra Tech estimates to reflect the 2011 drillhole locations and depths.

14.3.1Geostatistics of the Batman Deposit

Geology of the Batman Deposit consists of a sequence of hornfelsed interbedded greywackes, and shales with minor thin beds of felsic tuffs. Minor lamprophyre dykes trending north-south crosscut the bedding. The mineralized lithologic package consists of a tabular deposit striking at 325 o with a dip of 40 o to 60 o to the southeast. The majority of drilling slants at a dip of approximately 65 o with an azimuth of 270 o.

Bedding parallel shears are present in some of the shale horizons (especially in lithologic units SHGW23, GWSH23, and SH22). These bedding shears are identified by quartz/calcite sulfidic breccias. Pyrite, pyrrhotite, chalcopyrite, galena, and sphalerite are the main primary sulfides associated with the bedding parallel shears.

NE-SW trending faults and joint sets crosscut bedding. Only minor movement has been observed on these faults. Calcite veining is sometimes associated with these faults. These structures appear to be post mineralization.

Northerly trending quartz sulfide veins and joints striking at 0 o to 20 o, dipping to the east at 60 o, are the major location for mineralization in the Batman Deposit. The veins are 1 to 100 mm in thickness with an average thickness of around 8 to 10 mm. The veins consist of dominantly quartz with sulfides on the margins. The veining occurs in sheets with up to 20 veins per horizontal meter. These sheet veins are the main source of mineralization in the Batman Deposit.

The mineralization within the Batman Deposit is directly related to the intensity of the north-south trending quartz sulfide veining. The lithological units impact on the orientation and intensity of mineralization. Sulfide minerals associated with the gold mineralization are pyrite, pyrrhotite and lesser amounts of chalcopyrite, bismuthinite and arsenopyrite. Galena and sphalerite are also present but appear to be post gold mineralization and are related to calcite veining bedding and the east-west trending faults and joints.

Multiple directional variograms explored the best continuity of mineralization given the combination of control by bedding and sulfide veining. Figure 14‑4 is an example of two log variograms in the core complex.

Figure 14‑4: 2012 Example Log Variograms of Gold within the Core Complex

Table 14‑4 shows the resource classification criteria and variogram for the Batman resource model.

Table 14‑4: Batman Resource Classification Criteria and Variogram

Category | Search Range & Kriging Variance | No. of Sectors/ Max Points per DH | Search Anisotropy | Min Points | Composite Codes | Block Codes | CORE
Indicated | Core Complex: 150 m & KV < 0.45 Pass 1 | 4/2 | (1.0:0.7:0.4) [110:80:0] | 2 | 1000 | 1000 | CORE COMPLEX
Measured | Core Complex: 60 m & KV < 0.30) Pass 2 (overwrite Pass 1) | 4/3 | (1.0:0.7:0.4) [110:80:0] | 4 | 1000 | 1000
inferred | Core Complex KV >= 0.34 Classification Step | 4/2 | (1.0:0.7:0.4) [110:80:0] | 2 | 1000 | 1000
inferred | Outside Core Complex: 150 m & KV <= 0.45 Pass 3 | 4/3 | (1.0:0.7:0.4) [110:80:0] | 3 | 500/3500 | 500/ 3500 | OUTSIDE CORE COMPLEX
inferred | Outside Core Complex: 50 m & KV > = 0.45 Pass 4 (overwrite Pass 3) | 4/3 | (1.0:0.7:0.4) [110:80:0] | 8 | 500/3500 | 500/ 3500
inferred | Primary Satellite Deposit: 150 m & KV >= 0.45 Pass 5 | 4/3 | (1.0:0.7:0.4) [110:80:0] | 3 | 600 | 600
Indicated | Primary Satellite Deposit: 50 m & KV < 0.45 Pass 6 --(overwrite Pass 5) | 4/3 | (1.0:0.7:0.4) [110:80:0] | 8 | 600 | 600
inferred | Secondary Satellite Deposit: 150 m & KV >= 0.45 Pass 7 | 4/3 | (1.0:0.7:0.4) [110:80:0] | 3 | 700 | 700
Indicated | Secondary Satellite Deposit: 50 m & KV < 0.45 Pass 8 (overwrite Pass 7) | 4/3 | (1.0:0.7:0.4) [110:80:0] | 8 | 700 | 700
inferred | Tertiary Satellite Deposit: 150 m & KV >= 0.45 Pass 9 | 4/3 | (1.0:0.7:0.4) [110:80:0] | 3 | 800 | 800
Indicated | Tertiary Satellite Deposit: 50 m & KV < 0.45 Pass 10 (overwrite Pass 9) | 4/3 | (1.0:0.7:0.4) [110:80:0] | 8 | 800 | 800
VARIOGRAM FOR ALL CATEGORIES
Type: Spherical Primary Axis: 150m Nugget: 0.6 First Rotation (Azimuth: 110) Secondary Axis: 105m Sill 1: 0.3 Range 1: 40m Second Rotation (Dip: 80) Tertiary Axis: 60m Sill 2: 0.2 Range 2: 500m Third Rotation (Tilt: 0)

INDEX
Zone Codes | Zone Names | Notes
3500 | Footwall | Ranges In meters (m) KV = kriging variance, Passes refer to multiple re-estimations of blocks with greater constraints (minimum points, search ranges, etc.) imposed. Core and Satellites have more consistent gold grades, while the Footwall and Hanging Wall have patchy gold grades, Search Ranges (a:b:c) Proportion of Maximum Range for: a. Primary Axis Length: b. Secondary Axis Length: c. Tertiary Axis Length Orientation of Ellipse [1:2:3] 1. Azimuth of Primary Axis : 2. Dip of Primary Axis: 3. Rotation (Tilt) around Primary Axis
1000 | Core Complex
800 | Tertiary Satellite (between 600 and 700)
700 | Secondary Satellite (in HW farthest from Core)
600 | Primary Satellite (in HW nearest to Core)
500 | Hanging Wall Area

Figure 14‑5 through Figure 14‑10 are a series of sections and plan views of the Batman deposit block model from 2013. An insert showing the 2017 block model within the core and outlier zones has been added. Note that the granite layer that truncates the lowest portion of the deposit was re-modelled from a steeply plunging surface in 2013 to more horizontal surface in 2017.

Figure 14‑5: 2013 Study – Blocks Kriged Au – Cross-section 8,434,900 mN Looking North, Batman Deposit

Figure 14‑6: 2013 Study – Classified Blocks Measured, Indicated, and Inferred – Cross-section 8,434,900 mN

Looking North, Batman Deposit

Figure 14‑7: 2013 Study -- Blocks Kriged Au – Level Plan -100m msl Batman Deposit

Figure 14‑8: 2013 Study – Classified Blocks Measured, Indicated, and Inferred – Level Plan -100m msl

Batman Deposit

Figure 14‑9: 2013 Study -- Blocks Kriged Au – Long Section of the Core Complex Looking West

Figure 14‑10: 2013 Study – Classified Blocks Measured, Indicated, and Inferred –

Long Section of the Core Complex Looking West

Table 14‑5 lists the current 2017 Batman measured and indicated resource estimates at cutoff grades ranging from 0.3 g-Au/t to 2.0 g-Au/t. Table 14‑6 lists the current 2017 Batman inferred resource estimates at cutoff grades ranging from 0.3 g-Au/t to 2.0 g-Au/t.

Figure 14‑11 graphically shows the grade-tonnage for Measured plus Indicated classified Batman deposit resources.

Table 14‑5: Batman Deposit Measured and Indicated Gold Resource Estimate

Cutoff Grade g-Au/t | Tonnes (x1000) | Average Grade g-Au/t | Total Au Ounces (x1000)
MEASURED | 2.00 | 2,474 | 2.40 | 191.1
1.75 | 4,616 | 2.15 | 319.4
1.50 | 8,186 | 1.92 | 505.1
1.25 | 13,205 | 1.71 | 725.8
1.00 | 21,512 | 1.48 | 1,024.0
0.90 | 26,481 | 1.38 | 1,175.2
0.80 | 33,167 | 1.27 | 1,357.2
0.70 | 41,594 | 1.17 | 1,560.0
0.60 | 52,492 | 1.06 | 1,787.3
0.50 | 64,597 | 0.96 | 2,001.1
0.40 | 77,725 | 0.88 | 2,191.1
0.3 | 90,719 | 0.80 | 2,337.8
2.00 | 5,413 | 2.55 | 443
INDICATED | 1.75 | 9,124 | 2.27 | 666
1.50 | 15,194 | 2.01 | 982
1.25 | 25,183 | 1.75 | 1,420
1.00 | 43,059 | 1.49 | 2,057
0.90 | 54,104 | 1.38 | 2,394
0.80 | 68,845 | 1.26 | 2,796
0.70 | 88,256 | 1.15 | 3,262
0.60 | 115,528 | 1.03 | 3,830
0.50 | 153,278 | 0.91 | 4,494
0.40 | 200,112 | 0.80 | 5,169
0.30 | 253,187 | 0.71 | 5,765
2.00 | 7,887 | 2.50 | 634
MEASURED + INDICATED | 1.75 | 13,740 | 2.23 | 985
1.50 | 23,380 | 1.98 | 1,487
1.25 | 38,387 | 1.74 | 2,145
1.00 | 64,571 | 1.48 | 3,081
0.90 | 80,585 | 1.38 | 3,569
0.80 | 102,012 | 1.27 | 4,153
0.70 | 129,850 | 1.16 | 4,822
0.60 | 168,021 | 1.04 | 5,617
0.50 | 217,875 | 0.93 | 6,495
0.40 | 277,837 | 0.824 | 7,360
0.30 | 343,906 | 0.73 | 8,104
0.30 | 343,906 | 0.73 | 8,104

NOTE:

(1)The measured and indicated resource estimates presented in this table include the proven and probable reserves presented in Section 15 of this Technical Report.

(2)Mineral resources that are not mineral reserves have no demonstrated economic viability and do not meet all relevant modifying factors.

Mineral resources that are not mineral reserves have no demonstrated economic viability and do not meet all relevant modifying factors.

Figure 14‑11: Grade Tonnage Curve of Measured and Indicated Resource for the Batman Deposit

Table 14‑6: Batman Deposit Inferred Gold Resource Estimate

Cutoff Grade g-Au/t | Tonnes (x1000) | Average Grade g-Au/t | Total Au Ounces (x1000)
INFERRED | 2.00 | 1,664 | 2.94 | 157
1.75 | 2,196 | 2.68 | 189
1.50 | 2,975 | 2.40 | 230
1.25 | 4,532 | 2.05 | 298
1.00 | 7,914 | 1.65 | 419
0.90 | 10,170 | 1.49 | 487
0.80 | 14,327 | 1.30 | 601
0.70 | 19,576 | 1.15 | 726
0.60 | 27,798 | 1.00 | 897
0.50 | 40,964 | 0.86 | 1,128
0.40 | 61,323 | 0.72 | 1,421
0.30 | 94,532 | 0.59 | 1,790

NOTE:

(1)Resources constrained within a US$1,300/oz gold Whittle TM Pit Shell. Pit parameters: Mining Cost US$1.50/tonnes, Milling Cost US$7.80/tonnes processed, G&A Cost US$0.46/tonnes processed, 50K TPD Ore, 355 Days/Yr., TPY Ore 17,750,000 TPY, G&A/Year 8,201 K US$, Au Recovery, Sulfide 85%, Transition 80%, Oxide 80%, 0.2g-Au/t minimum for resource shell. Tonnage, grades and totals may not total due to rounding. The reported resources at a cutoff of 0.4 g/t is highlighted.

(2)Mineral resources that are not mineral reserves have no demonstrated economic viability and do not meet all relevant modifying factors.

14.4Batman Estimation Quality

Several methods were used to validate the block model to determine the adequacy of the Batman deposit resource. Confirmatory drilling was used to ascertain the general good quality of the model within the core zone. In addition, overlaid cumulative frequency plots of blocks, composites, and assays were used. The three overlaid plots showed the expected decrease in the variability of the gold distributions going from assays to assay composites and then to kriged blocks. In addition:

Jackknife studies were employed to determine the optimum kriging search parameters and the overall quality of the estimation as required by classification. Figure 14‑12 shows the Jackknife results for the measured class.

Numerous swath plots were analyzed in the direction of rows and columns were used to verify that composite and block gold grades are spatially in sync. Several examples of these swath plots are shown in Figure 14‑13.

The use of visual inspection of the kriged blocks models in section and plan and the inspection of gold histograms of assays, composites and blocks.

Figure 14‑12: Jackknife Correlation Plot for Measured Blocks

Figure 14‑13: Jackknife Correlation Plot for Inferred Blocks

14.5Modeling of the Quigleys Deposit

The Quigleys Deposit is located approximately 3.5 km northeast of the Batman Deposit. The deposit is not as deep as the Batman deposit; it reaches a maximum depth of approximately 200 m. The deposit has been sampled with 57,600 m of drilling by 631 drillholes, with the majority reaching a depth of 100m at a 60 degree dip; oriented 83 degrees azimuth. Assays were taken at a nominal one meter interval. Geologic interpretation in section produced wireframes modeling thin ore zones dipping west. Material inside the wire frames was given a code of 1. Outside the mineralization zones, the material was given a code of 9999.

Bulk density data were supplied by Pegasus for two ore types and waste within the oxide, transition and primary zones, based on a total of 39 samples collected from RC drilling. The two densities supplied were for stockwork and shear, with the density of the shear material substantially higher, particularly in the transition and primary zones. These samples were over one-m to two-m intervals and thus selected the narrow high grade portion of the shear zone as originally interpreted by Pegasus. The final mineralization envelope was much broader than this, and the bulk density was therefore estimated by assuming the final envelope contained 15% shear and 85% stockwork and weighting the density values accordingly. Table 14‑7 shows the specific gravity data assigned to the Quigleys area according to oxidation state.

Table 14‑7: Quigleys Deposit Specific Gravity Data

Oxide within modeled shear (t/cm) | 2.60
Oxide Waste (t/cm) | 2.62
Transition within modeled shear (t/cm) | 2.65
Transition Waste (t/cm) | 2.58
Primary within modeled shear (t/cm) | 2.70
Primary Waste (t/cm) | 2.61

14.5.1Quigleys Exploration Database

Table 14‑8 summarizes the Quigleys exploration database.

Table 14‑8: Summary of Quigleys Exploration Database

Drillhole Statistics
| Northing (m) AMG84 z53 | Easting (m) AMG84 z53 | Elevation (m) | Azimuth | Dip | Depth (m)
Minimum | 8,430,1876 | 188,445.7 | 129.7 | 0 | 45 | 0
Maximum | 8,432,290 | 189,746.5 | 209.0 | 354.0 | 90 | 330.5
Average | 8,431,129.5 | 189,230.8 | 155.9 | 83.4 | 62.5 | 91.3
Range | 2,104.0 | 1,300.8 | 79.3 | 354.0 | 45.0 | 330.5
Cumulative Drillhole Statistics |
Total Count | 631 |
Total Length (m) | 57,821
Assay Length (m) | 1 (approx.)
Drillhole Grade Statistics | Number | Average | Std. Dev. | Min. | Max. | Missing
Au (g/t) | 52,152 | 0.2445 | 0.8764 | 0 | 36.00 | 82
Cu (%) | 40,437 | 0.0105 | 0.0305 | 0 | 2.98 | 11,897

14.5.2Quigleys Block Model Parameters

Quigleys’ block model parameters are shown in Table 14‑9. The model consisted of 37,082 blocks within the modeled mineralized zones (blocks within the modeled grade zones are coded as 1). Each of the blocks is 250 m 3 (5x25x2m) with a defined density of 2.77 g/cm (692.5 tonnes).

Table 14‑9: Block Model Physical Parameters – Quigleys Deposit

Direction | Minimum (m) AMG84 z53 | Maximum (m) AMG84 z53 | Block Size | # Blocks
x-dir | 188,250 mE | 189,900 mE | 5m | 330
y-dir | 8,430,337.5 mN | 8,432,487.5mN | 25m | 86
z-dir | -200 m | 208m | 2m | 204

Figure 14‑14 shows the rock codes used for the Quigleys estimation.

Figure 14‑14: 3-D Visualization of the Quigleys Deposit Mineralized Zone Positions with Wireframe Codes

The cap value of 12.0 g-Au/t has been chosen based on review of natural log transformed histograms, cumulative frequency and probability plots. Review of the log probability plot of the composited gold grades shows that there is a distinct break in the distribution at 12 g-Au/t. All gold composites were capped at this value.

Two surfaces were generated based on historic downhole logging of drill holes. The first surface represents the boundary between weathered mineral type (oxide) and transition mineral type (mixed), and the second surface represents the boundary between transition mineral type and fresh mineral type (sulfide).

A number of absolute, log and indicator directional variograms were calculated and modeled. From these, Figure 14‑15 shows the log (Au) variogram for along strike, down dip and down hole coded as AStrk, DDip, Dhole respectively. These variograms have a nugget of 0.77, with an ultimate sill of 2.74. The ranges are 90 meters Along Strike (AStrk) and 30 m Down Dip (DDip). Table 14‑10 shows the search parameters selected for each domain.

Figure 14‑15: Quiqleys Median Indicator Variogram

Table 14‑10: Search Parameters for each Domain

Code | Azimuth | Dip | Axis1 m | Axis2 m | Axis3 m
0 | 280 | 35 | 90 | 90 | 30
1000 | 266 | 26 | 90 | 90 | 30
2000 | 273 | 35 | 90 | 90 | 30
3000 | 266 | 26 | 90 | 90 | 30
4000 | 273 | 35 | 90 | 90 | 30
5000 | 275 | 30 | 90 | 90 | 30
6100 | 280 | 35 | 90 | 90 | 30
6200 | 280 | 55 | 90 | 90 | 30
6300 | 280 | 70 | 90 | 90 | 30
7000 | 300 | 25 | 90 | 90 | 30

Table 14‑11 lists the resource classification criteria. The classification was accomplished by a combination of search distance, kriging variance, number of points used in the estimate, and number of sectors used. The block model was estimated using ordinary kriging. The estimation searched for four composites in a sector, allowing a maximum of three composites per drillhole. Inside the ore zone (blocks coded as “1”); composites were selected only if they also were coded as “1”. Separate kriging passes were done at increasing search distances. The first pass and second pass restricted points to be within 30 m and 90 m as defined by the search ellipsoid axis to produce provisional resources classes of measured and indicated. Review of the kriging error plotted as a log-probability graph indicated that the gold estimates were particularly poor when kriging variances were greater than 1.0 and 1.55 for the measured and indicated classes respectively. Hence the provisional Measured, Indicated, inferred (MIF) codes were then adjusted to a more restricted class when a blocks kriging error exceeded this value.

Table 14‑11: Search Parameters and Sample Restrictions

Domain | Class | Drill Holes | Max Sample Per Drill Hole | Search Major | Search Semi-major | Search Minor | Kriging Error
1000 to 7000 | Measured | >= 3 | 4 | 30 | 30 | 10 | <=1.00
1000 to 7000 | Indicated | >=2 | 4 | 90 | 90 | 30 | <=1.55
1000 to 7000 | inferred | >=1 | 4 | 90 | 90 | 30 | NA
0 | inferred | >=2 | 2 | 30 | 30 | 10 | NA

For the outside zone, a two-stage kriging for MIF class 3 was done inside and outside of modelled wireframes with a maximum search ellipse range of 90 m and 30 m respectively.

Each domain was assigned a unique search orientation; however kriging parameters were the same for all domains. Blocks with a given domain code were estimated only by composites of the same code.

Several methods used to validate the block model were used to determine the adequacy of the Quigleys resource. Cumulative frequency plots of blocks, composites, and assays were overlaid. The three overlaid plots showed the expected decrease in the variability of the gold distributions going from assay to assay composites and then to kriged blocks. Additional verification of the block model was completed by the use of jackknife studies (model validation) where known assays were estimated using surrounding samples, visual

inspection of the kriged blocks models in section and plan and the inspection of gold histograms of assays, composites and blocks.

Table 14‑12 lists the parameters used to generate a Whittle TM pit shell for reporting the measured plus indicated resource in Table 14‑13 and the inferred resource in Table 14‑14.

Mineral resources that are not mineral reserves have no demonstrated economic viability and do not meet all relevant modifying factors.

Figure 14‑16 shows the grade tonnage relations for measured plus indicated classified resource.

Table 14‑12: Whittle TM Pit Shell Parameters

Item | Input
Gold Price | US$1,200 per troy ounce
Gold Recovery | 82% Sulfide 78% Transition 78% Oxide
Payable Gold | 99.90%
Overall Mining Cost | US$1.90 per tonne
Processing Cost | US$9.779 per tonne processed
Tailings | US$0.985 per tonne processed
Water Treatment | US$0.09 per tonne processed
Royalty | 1% NPR (JAAC)
Sell Cost | US$3.19

Table 14‑13: Quigleys Deposit Measured and Indicated Gold Resource Estimate within M&I Whittle TM Shell (August 2017)

Cutoff Grade g-Au/t | Tonnes (x1000) | Average Grade g-Au/t | Total Au Ounces (x1000)
Measured + Indicated | 2.0 | 700 | 3.04 | 68
1.5 | 1,200 | 2.51 | 97
1.0 | 2,300 | 1.86 | 139
0.4 | 6,200 | 1.13 | 225
0.0 | 6,400 | 1.11 | 228

NOTE:

(1)Resources constrained within a US$1,200/oz gold Whittle TM Pit Shell. Pit parameters: Mining cost US$2.07/tonnes, Milling Cost US$9.623/tonnes processed, Sale Cost US$/oz US$15.18, Royalty 1% NPR (JAAC), Gold Recovery All Types, 70%.

(2)Tonnage, grades and totals may not total due to rounding. The reported resources at a cutoff of 0.4 g/t is highlighted.

(3)There are no mineral reserves at the Quigleys deposit at this time.

(4)For measured and indicated defined at the chosen cutoff grade, reference Table 14‑1.

Mineral resources that are not mineral reserves have no demonstrated economic viability and do not meet all relevant modifying factors.

Figure 14‑16: Measured and Indicated Grade Tonnage Curve for the Quigleys Deposit

Table 14‑14: Quigleys Deposit Inferred Gold Resource Estimate within M&I Whittle TM Shell (August 2017)

Cutoff Grade g-Au/t | Tonnes (000s) | Average Grade g-Au/t | Total Au Ounces (000s)
inferred | 2.0 | 91 | 2.74 | 8
1.5 | 204 | 2.19 | 14
1.0 | 338 | 1.79 | 19
0.4 | 1,600 | 0.84 | 43
0.0 | 20,400 | 0.18 | 120

NOTE:

(1)Resources constrained within a US$1,200/oz gold Whittle TM Pit Shell. Pit parameters: Mining cost US$2.07/tonnes, Milling Cost US$9.623/tonnes processed, Sale Cost US$/oz US$15.18, Royalty 1% NPR (JAAC), Gold Recovery All Types, 70%.

(2)Tonnage, grades and totals may not total due to rounding. The reported resources at a cutoff of 0.4 g/t is highlighted.

(3)There are no mineral reserves at the Quigleys deposit at this time.

14.6Existing Heap Leach Gold Resource

In addition to the in-situ gold resource for the Batman Deposit, a historical heap leach pad (HLP) adjacent to the current Mt Todd pit was analyzed for gold. The HLP is a remnant of the Pegasus operation, pre-2006. The HLP’s geometry was analyzed using historical maps to determine the pile bottom and current surveys of the present day surface. This work produced two surfaces which were used to calculate the volume of the pile. The concentration of gold was analyzed with 24 vertical drillholes separated by an approximately 100 meters. Drilling depth was terminated 5-meters before the final depth of the heap to keep from piercing the bottom liner. The 363 assays from 1-m composites were analyzed for gold and copper grade. Density of the pile was estimated from 11 drillholes using 1,162 dual density sidewall gamma probe technology. Note that the probe uses a gamma source and a scintillation detector to estimate density via the Compton Effect.

A nearest neighbor (polygon) method was employed to estimate grades within the heap leach pad since there is no apparent spatial correlation between samples. The existing heap leach pad is estimated to contain 230,000 ounces of gold within 13.4 Mt of indicated mineral resource at an average grade of 0.54 g-Au/t. It is the opinion of Tetra Tech that the heap leach resource can be classified as an indicated mineral resource as the surveyed volume, the tonnage derived from density measurements, and grade assays from drillhole sampling reconciles with Pegasus’ original reported values.

Table 14‑15 lists the indicated mineral resources for the existing heap leach pad. Note the grade-tonnage plot in Figure 14‑17. The majority of tonnage has an average grade of 0.5 g-Au/t as indicated by the flat portion of the tonnage curve. Note too that no cutoff grade was applied to the heap leach pad resource as all material will be processed as part of the site rehabilitation process. Copper was also estimated, but the copper results are not presented here.

Table 14‑15: Existing Heap Leach Indicated Gold Resource Estimate (May 2013)

Cutoff Grade g-Au/t | Tonnes (000s) | Average Grade g-Au/t | Total Au Ounces (000s)
INDICATED | 0 | 13,400 | 0.541 | 230

NOTE:

(1)No cutoff grade is technically applied due to all heap leach material will be re-processed. Resources are reported at 0.4 g/t cutoff gold grade to be consistent with the reported Batman and Quigleys resource is Resource is defined by the geometry of the existing heap leach pad.

(2)Resource & reserve estimates for the heap leach materials are the same because 100% of the heap leach material is processed at the conclusion of mining the Batman Pit.

Mineral resources that are not mineral reserves have no demonstrated economic viability and do not meet all relevant modifying factors.

Figure 14‑17: Inferred Resource Grade Tonnage Curve for the Quigleys Deposit

14.7Relevant Factors Affecting Resource Estimates

There are currently no known environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other relevant factors which could affect the mineral resource estimate.

15.0MINERAL RESERVES

The Base Case is identified as a 50,000 tpd operation, as presented in this section. Another option, defined as the Alternate Case (33,000 tpd) is presented in Section 24.7 – Alternate Case.

The measured and indicated resource estimates as of August 2017 were used to estimate reserves.

Reserve definition is done by first identifying ultimate pit limits using economic parameters and pit optimization techniques. The resulting optimized pit shells were then used for guidance in pit design to allow access for equipment and personnel. Several phases of mining were defined to enhance the economics of the project, and MDA used the phased pit designs to define the production schedule to be used for cash-flow analysis for the preliminary feasibility study.

The following section details the definition of reserves used for the production scheduling. Later sections detail the production schedule and the mining costs used in the Tetra Tech cash-flow model.

15.1Pit Optimization

Pit optimization was done using Geovia’s Whittle TM software (version 4.7) to define pit limits with input for economic and slope parameters. The optimization used parameters provided by Vista and their consultants based on current and previous studies.

Optimization used only measured and indicated material for processing. All inferred material was considered as waste.

Varying gold prices were used to evaluate the sensitivity of the deposit to the price of gold, as well as to develop a strategy for optimizing project cash flow. To achieve cash-flow optimization, mining phases or push backs were developed using the guidance of Whittle TM pit shells at lower gold prices.

15.1.1Economic Parameters

Initially, several iterations of pit optimizations were reviewed for the final determination of pit limits based on the Base Case parameters.

Initial mining cost parameters were based on the economic parameters provided in Table 15‑1. The final mining costs from this study have turned out to be lower than those in Table 15‑1, thus making the pit optimization conservative with respect to costs and resulting reserves.

Table 15‑1: Initial Economic Parameters

Parameter | Base Case
Gold Recovery | 85% Sulfide 80% Transition 80% Oxide
Payable Gold | 99.9%
Overall Mining Cost | US$1.90 per tonne
Processing Cost | US$7.80 per tonne processed
Tailings | US$0.90 per tonne processed
General & Administrative | $0.46 per tonne processed

Water Treatment | US$0.10 per tonne processed
Royalty | 1% gross proceeds

The mining costs used were varied by bench. An incremental cost of US$0.010/tonne was added for each 6-meter bench below the 145 meter elevation. This represents the incremental increase in cost of haulage for both waste and ore for each bench that is to be mined. Reference mining costs of US$1.64/tonne for the Base case. The incremental cost was determined based on truck operating costs, truck cycle time to haul and return through a six-meter gain in elevation, and truck capacity. The reference mining cost was determined using first principles from previous studies. The total mining cost (reference plus incremental) is US$1.90 for the Base Case.

Processing, tailings construction, tailings reclamation, and water treatment costs were provided by Vista and their consultants. Calculated cutoff grades based on the economic parameters are 0.38 and 0.33 g-Au/t for the Alternate and Base cases respectively. At Vista’s request, MDA used a minimum cutoff grade of 0.40 g-Au/t for the Base Case. This was done to maintain higher grades with respect to material to be processed.

A base gold price of US$1,250 per ounce was determined by Vista for use in scenario analysis. However, various gold prices from US$300 to US$2,000 per ounce, in increments of US$20 per ounce, were used to determine different optimized pit shells.

Final recoveries for the sulfide ore is slightly higher (85.6% vs. 85%) and slightly lower (79.5% vs. 80%) for the oxide and transition ores. Final process operating costs are higher; however, the imposition of an elevated cutoff grade and selection of pit limits using a lower gold price pit shell create additional conservatism that offset the processing cost and recovery adjustments.

15.1.2Slope Parameters

The slope parameters were based on studies provided by Golder Associates and Ken Rippere as detailed in a Golder memo dated September 13, 2011 (“Mt Todd Gold Project: Batman Pit Slope Design Guidance in Support of the Definitive Feasibility Study”). Minor modifications were made based on comments from Call & Nicholas, Inc. (2016). The Golder parameters suggested that the catch benches would not be maintainable on the east side of the pit and that these catch benches should not be placed in the design. For safety, the roads on the east side were widened to allow a berm to be maintained along the road to contain any rock that would slough off of the wall.

The primary change suggested by Call & Nicholas (2016) is to either place catch benches in the high wall on the east side, or bolt and mesh the high wall. In both cases the ramp along the wall would be reduced to a normal width.

For this study, catch benches were inserted in preliminary pit phases. However, the ultimate pit used a flat slope with bolting and mesh. This helps to reduce the overall slope and the resulting stripping required.

For pit optimization, slopes were divided into five sectors in fresh rock and three additional sectors for weathered rock. The sectors are based on the wall orientation and are shown in Figure 15‑1. Each sector was modeled into a zone resulting in eight zones. Slopes on the eastern side of the pit were reduced to account for ramps in the high wall. The recommended and adjusted inner-ramp angles are shown in Table 15‑2.

Table 15‑2: Slope Angles for Pit Optimization

Zone | Sector | Slope Angle (°) | Adjusted Angle (°)
1 | Northeast | 36 | 33
2 | East | 40 | 36
3 | South | 55 | 50
4 | Southwest | 55 | 55
5 | Northwest | 51 | 51
6 | Northeast & East Weathered | 33 | 33
7 | South & Southwest Weathered | 45 | 45
8 | Northwest - Weathered | 45 | 45

Figure 15‑1: Mt Todd Geotechnical Sectors

15.1.3Pit-Optimization Results

Whittle TM pit optimizations were run using the economic and slope parameters described in previous sections. Pit optimizations were completed using prices of US$300 to US$2,000 per ounce gold with increments of US$20 per ounce. One additional pit shell was created using a US$1,250 per ounce gold price. These pits were used to assess the deposit’s sensitivity to gold prices for both scenarios. Results for US$100 per ounce increments, from US$300 to US$2,000 per ounce of gold, are shown in Table 15‑3, with a highlighted price of US$1,000/oz-Au as the pit shell used to guide the ultimate pit design and US$1,250/oz-Au used as the base price for Whittle TM analysis. The pit optimizations only used measured and indicated resources. Inferred materials are considered waste.

Graphs of the tonnes and contained ounces from the Whittle TM results are shown in Figure 15‑3.

Table 15‑3: Whittle TM Pit Optimization Results – Base Case using 0.40 g-Au/t Cutoff

Pit | Gold Price (US$) | Material Processed | Waste Tonnes | Total Tonnes | Strip Ratio
K Tonnes
1 | $ 300 | 3,282 | 1.77 | 186 | 2,797 | 6,078 | 0.85
6 | $ 400 | 8,578 | 1.54 | 425 | 7,507 | 16,085 | 0.88
11 | $ 500 | 15,988 | 1.34 | 686 | 15,740 | 31,728 | 0.98
16 | $ 600 | 37,253 | 1.12 | 1,340 | 53,757 | 91,010 | 1.44
21 | $ 700 | 89,301 | 0.99 | 2,855 | 171,617 | 260,918 | 1.92
26 | $ 800 | 121,187 | 0.92 | 3,585 | 222,919 | 344,106 | 1.84
31 | $ 900 | 159,485 | 0.87 | 4,442 | 316,889 | 476,374 | 1.99
36 | $ 1,000 | 185,915 | 0.85 | 5,093 | 429,208 | 615,123 | 2.31
41 | $ 1,100 | 212,340 | 0.84 | 5,741 | 566,907 | 779,247 | 2.67
46 | $ 1,200 | 230,587 | 0.83 | 6,184 | 675,714 | 906,302 | 2.93
49 | $ 1,250 | 234,858 | 0.83 | 6,278 | 700,519 | 935,376 | 2.98
51 | $ 1,300 | 240,195 | 0.83 | 6,416 | 742,833 | 983,029 | 3.09
56 | $ 1,400 | 243,306 | 0.83 | 6,498 | 771,190 | 1,014,497 | 3.17
61 | $ 1,500 | 249,389 | 0.83 | 6,658 | 829,933 | 1,079,321 | 3.33
66 | $ 1,600 | 254,050 | 0.83 | 6,779 | 880,583 | 1,134,633 | 3.47
70 | $ 1,700 | 254,348 | 0.83 | 6,785 | 883,222 | 1,137,571 | 3.47
74 | $ 1,800 | 259,140 | 0.83 | 6,908 | 943,012 | 1,202,152 | 3.64
78 | $ 1,900 | 259,964 | 0.83 | 6,927 | 952,872 | 1,212,836 | 3.67
81 | $ 2,000 | 260,099 | 0.83 | 6,929 | 953,985 | 1,214,083 | 3.67

Pit 36 was used for design purposes and Pit 49 illustrates the potential floating cone using a US$1,250/oz-Au price.

Figure 15‑2: Measured and Indicated Graph of Whittle TM Results – Base Case using 0.40 g-Au/t Cutoff

15.1.4Ultimate Pit Limit Selection

The ultimate pit limit was determined based on various iterations analyzed by MDA, Vista, and Tetra Tech. A lower gold price pit shell was used as a guide for the ultimate pit design. This decision was made to reduce the project footprint while still capturing the most valuable material in the pit optimizations.

For consistency, pit shells that closely replicate the tonnage from the previous NI 43-101 Technical Report (Tetra Tech, 2014) were selected for guidance in pit design. MDA selected US$1,000 and US$1,020 gold pit shells for the Base Case.

15.2Pit Designs

Detailed pit designs were completed, including an ultimate pit and three internal pits for the Base Case. The ultimate pits were designed to allow mining economic resources identified by Whittle TM pit optimization, while providing safe access for people and equipment. Internal pits or phases within the ultimate pits were designed to enhance the project by providing higher-value material to the processing plant earlier in the mine life.

15.2.1Bench Height

Pit designs used six-meter benches for mining. This corresponds to the resource model block heights, and the QP for this section believes this to be reasonable with respect to dilution and equipment anticipated to be used in mining. In areas where the material is consistently ore or waste, so that dilution is not an issue, benches may be mined in 12-meter heights.

15.2.2Pit Design Slopes

The slope parameters were based on geotechnical studies by Golder Associates and Ken Rippere (Golder, 2011). These were reviewed by Ross Barkley of Call & Nicholas (Barkley, 2016), and the slope parameters were modified based on his recommendations. The largest difference between the previous and 2017 slopes is in the use of catch benches on the eastern walls. The previous parameters specified the use of a flat wall on the east without any catch benches. To keep rocks from rolling down on trucks, the ramps were designed to be 28 m wide (total width of 50 m), so that a berm could be placed and rock would collect at the base of the slope behind the berm.

Call & Nicholas (2016) commented that “for interim phases, assuming a 47- to 50-degree bedding dip the interramp angles should be 37- to 39-degrees in order to maintain a 9+ meter wide catch bench”. For the final walls in the northeast area, the recommendations continue to state “… the walls will be smooth and excavated to the bedding dip” and “To mitigate the rock fall risk in the final wall, it is recommended that mesh be installed over the interramp slopes between the ramps.”

The recommended slopes are developed around five different sectors in fresh rock and three sectors in weathered rock as shown in Table 15‑4. The design parameters used are shown in Table 15‑4 for the ultimate pit and Table 15‑5 shows the sector 1 and 2 slope parameters for interim pit designs. The parameters are applied based on height between catch benches in meters (BH), safety berm widths in meters (berms), bench face angles in degrees (BFA) and inner-ramp angles also in degrees (IRA).

Table 15‑4: Pit Design Slope Parameters

| Due North | Sector 1 | Sector 2 | Sectors 3 & 4 | Sector 5 | Sector 6 | Sector 7 | Sector 8
BH (m) | 24 | 24 | 24 | 24 | 24 | 30 | 30 | 30
BFA (°) | 61 | 47 | 49 | 73 | 68 | 35 | 60 | 60
Berm (m) | 9.5 | - | - | 9.5 | 9.5 | 12.0 | 12.0 | 12.0
Net IRA (°) | 46.5 | 47.0 | 49.0 | 54.9 | 51.3 | 28.7 | 45.7 | 45.7

In the northern direction the slope azimuth must be 205 degrees or better.

Table 15‑5: Interim Pit Slope Parameters (Sectors 1 & 2)

| Northeast | East
BH (m) | 24 | 24
BFA (*) | 48 | 49
Berm (m) | 9.5 | 9.5
Net IRA (*) | 37.6 | 38.3
Zone | 1 | 2

For design purposes, weathered material is considered to be the top 30 meters from the surface.

15.2.3Haulage Roads

Ramps were designed to have a maximum centerline gradient of 10%. In areas where the ramps may curve along the outside of the pit, the inside gradient may be up to 11% or 12% for short distances. Designs use switchbacks to maintain the ramp system on the east side of the pit. This is done to better match the dip of the deposit and also allows better traffic connectivity between pit phases. In areas where switchbacks are employed, a maximum centerline gradient of 8% is used.

Ramp width was determined as a function of the largest haul truck width to be used. Mine plans use 226-tonne capacity trucks with operating widths of 8.30 meters. For haul roads inside of the pit, a single safety berm on the inside of the roadway will be required to be at least half the height of the largest vehicle tire that uses the road. MDA has designed safety berms with a 1.5 horizontal to 1 vertical slope using run-of-mine material, and a height of 1.97 meters, which provides half of the truck tire height plus 10% for the haul trucks. The 10% addition is used to ensure that the berm height exceeds half of the truck tire height in all cases. The resulting base width of safety berms is 5.9 meters.

Haul-roads inside of the pit, where only one safety berm is required, are designed to be 32 meters wide for two-way traffic. Subtracting berm widths, this provides 3.14 times the width of haul trucks for running width.

In lower portions of the pit, where haulage requirements allow use of one-way traffic, haul roads are designed to have a width of 20 meters. This provides 1.7 times the width of haul trucks for running width.

Haul roads outside of pit designs have been designed to be 42 meters wide to account for an additional safety berm.

15.2.4Ultimate Pit

The final ultimate pit design uses switchbacks to maintain the ramp system on the east side of the pit. This allows for better traffic flow between pit phases and allows the west side of the pit to best follow the dip of the deposit. In all, there are four switchbacks in the ultimate pit design and the lower portions of the pit have spirals to achieve the ultimate pit design.

The ultimate pit design, along with the ultimate dump and stockpile designs, and planned infrastructure, are shown in Figure 15‑3.

Figure 15‑3: Mt Todd Ultimate Pit Design – Base Case (February 12, 2018)

15.2.5Pit Phasing

Phase 1 essentially continues the western wall down from that done by prior operators, and wraps the ramp around the pit clockwise from the south. Phase 2 expands the pit to the east, north, and south, maintaining a portion of the phase 1 west wall. The phase 2 ramp is placed on the east wall and has a total of five switchbacks located in the north and south ends of the pit. Phase 3 will be mined to the final wall on the western side of the pit. Phase 4 expands the pit to the north, east, and south and mines under the phase 3 pit to the ultimate pit limits. The first three phases, prior to the ultimate pit, are presented in Figure 15‑4 to Figure 15‑6.

Figure 15‑4: Base Case Phase 1 Design (February 12, 2018)

Figure 15‑5: Base Case Phase 2 Design (February 12, 2018)

Figure 15‑6: Base Case Phase 3 Design (February 12, 2018)

15.3Cutoff Grade

The breakeven and internal cutoff grades calculated using the economic parameters are shown in Table 15‑6. The internal cutoff grade assumes that mining is constrained to an economic pit and does not include the mining cost.

To enhance projects economics, Vista has decided to use an elevated cutoff grade for reserves and scheduling. Reserves are reported using 0.40 g-Au/t cutoff grades for the Base Case.

Table 15‑6: US$1,250 Gold Price Cutoff Grades (g-Au/t)

| Base Case
Sulfide | Transition | Oxide
Breakeven | 0.33 | 0.35 | 0.35
Internal | 0.28 | 0.29 | 0.29
Cutoff Grade Used | 0.40 | 0.40 | 0.40

For purposes of production scheduling, low-grade, medium-grade, and high-grade material was designated. The low-grade material used the reserve cutoff grade (0.40 g-Au/t). Medium-grade and high-grade cutoffs used were 0.55 and 0.85 g-Au/t, respectively.

15.4Dilution

The resource block model was estimated with block sizes of 12m by 12m by 6m, and this model was used to define the ultimate pit limit, and to estimate proven and probable reserves. The QP responsible for this section considers the 12m by 12m by 6m block size to be reasonable for open pit mining of the deposit and believes that this represents an appropriate amount of dilution for statement of reserves.

15.5Reserves and Resources

Mineral reserves for the project were developed by applying relevant economic criteria (modifying factors) in order to define the economically extractable portions of the estimated resources. MDA developed the reserves to be in accordance with NI 43-101, which is based on the CIM Standards. CIM Standards define modifying factors as:

Modifying factors are considerations used to convert mineral resources to mineral reserves. These include, but are not restricted to, mining, processing, metallurgical, infrastructure, economic, marketing, legal, environmental, social and governmental factors.

CIM Standards define mineral reserves as:

Mineral reserves are sub-divided in order of increasing confidence into probable mineral reserves and proven mineral reserves. A probable mineral reserve has a lower level of confidence than a proven mineral reserve.

A mineral reserve is the economically mineable part of a measured and/or indicated mineral resource. It includes diluting materials and allowances for losses, which may occur when the material is mined or extracted and is defined by studies at pre-feasibility or feasibility level as appropriate that include

application of modifying factors. Such studies demonstrate that, at the time of reporting, extraction could reasonably be justified.

The reference point at which mineral reserves are defined, usually the point where the ore is delivered to the processing plant, must be stated. It is important that, in all situations where the reference point is different, such as for a saleable product, a clarifying statement is included to ensure that the reader is fully informed as to what is being reported.

The public disclosure of a mineral reserve must be demonstrated by a pre-feasibility study or feasibility study.

Mineral reserves are those parts of mineral resources which, after the application of all mining factors, result in an estimated tonnage and grade which, in the opinion of the qualified person(s) making the estimates, is the basis of an economically viable project after taking account of all relevant modifying factors. Mineral reserves are inclusive of diluting material that will be mined in conjunction with the mineral reserves and delivered to the treatment plant or equivalent facility. The term ‘mineral reserve’ need not necessarily signify that extraction facilities are in place or operative or that all governmental approvals have been received. It does signify that there are reasonable expectations of such approvals.

‘Reference point’ refers to the mining or process point at which the qualified person prepares a mineral reserve. For example, most metal deposits disclose mineral reserves with a “mill feed” reference point. In these cases, reserves are reported as mined ore delivered to the plant and do not include reductions attributed to anticipated plant losses. The qualified person must clearly state the ‘reference point’ used in the mineral reserve estimate.

Probable Mineral Reserve

A probable mineral reserve is the economically mineable part of an indicated, and in some circumstances, a measured mineral resource. The confidence in the modifying factors applying to a probable mineral reserve is lower than that applying to a proven mineral reserve.

The qualified person(s) may elect, to convert measured mineral resources to probable mineral reserves if the confidence in the modifying factors is lower than that applied to a proven mineral reserve. Probable mineral reserve estimates must be demonstrated to be economic, at the time of reporting, by at least a Pre-Feasibility Study.

Proven Mineral Reserve

A proven mineral reserve is the economically mineable part of a measured mineral resource. A proven mineral reserve implies a high degree of confidence in the modifying factors.

Application of the proven mineral reserve category implies that the qualified person has the highest degree of confidence in the estimate with the consequent expectation in the minds of the readers of the report. The term should be restricted to that part of the deposit where production planning is taking place and for which any variation in the estimate would not significantly affect the potential economic viability of the deposit. Proven mineral reserve estimates must be demonstrated to be economic, at the time of reporting, by at least a pre-feasibility study. Within the CIM Definition standards the term proved mineral reserve is an equivalent term to a proven mineral reserve.

Proven and probable reserves are stated based on the Base Case pit designs. Table 15‑7 reports the proven and probable reserves, by pit phase, along with waste material for the pit designs discussed in previous sections.

RDi is responsible for reporting of the heap-leach pad reserves (Section 15.7 – Heap Leach Reserve). This is based on the tonnage and grade of heap-leach material that was loaded onto a heap-leach pad buy a historical operator, but due to pad operations issues, the gold was not recovered properly. The tonnes and grades are well known based on record keeping of the historical operator as discussed in Section 14.0 – Mineral Resource Estimate. The heap-leach reserves are shown with the Batman reserves in Table 15‑8.

The Base Case reserves are shown to be economically viable based on cash flow analysis provided by Tetra Tech. The QP responsible for this section has reviewed the cash flow and believes that they are reasonable for the statement of proven and probable reserves.

Table 15‑7: Base Case Proven and Probable Reserves by Pit Phase

| Proven |  |  | Probable |  |  | Total P&P |  |  | Waste | Total | Strip
| K Tonnes | g Au/t | K Ozs Au | K Tonnes | g Au/t | K Ozs Au | K Tonnes | g-Au/t | K Ozs Au | K Tonnes | K Tonnes | Ratio
Ph_1 | 13,551 | 1.09 | 473 | 6,245 | 1.10 | 221 | 19,796 | 1.09 | 694 | 19,312 | 39,109 | 0.98
Ph_2 | 18,980 | 0.80 | 490 | 19,008 | 0.88 | 538 | 37,988 | 0.84 | 1,028 | 59,167 | 97,155 | 1.56
Ph_3 | 20,356 | 0.87 | 571 | 27,471 | 0.85 | 747 | 47,827 | 0.86 | 1,318 | 122,372 | 170,198 | 2.56
Ph_4 | 19,785 | 0.82 | 523 | 82,291 | 0.78 | 2,052 | 102,076 | 0.78 | 2,576 | 322,138 | 424,215 | 3.16
Total | 72,672 | 0.88 | 2,057 | 135,015 | 0.82 | 3,559 | 207,687 | 0.84 | 5,616 | 522,990 | 730,677 | 2.52

NOTES:

(1)Proven and probable mineral reserves are reported using a cutoff grade of 0.40 g-Au/t.

(2)The reserves point of reference is the point where material is fed into the mill.

Table 15‑8: Total Batman Project Reserves (Base Case plus Heap Leach)

| Batman Deposit (January 2018) | Heap Leach Pad (May 2013) | Total P&P Reserves (January 2018)
| Tonnes (000s) | Grade (g/t) | Contained Ounces (000s) | Tonnes (000s) | Grade (g/t) | Contained Ounces (000s) | Tonnes (000s) | Grade (g/t) | Contained Ounces (000s)
Proven | 72,672 | 0.88 | 2,057 | - | - | - | 72,672 | 0.88 | 2,057
Probable | 135,015 | 0.82 | 3,559 | 13,354 | 0.54 | 232 | 148,369 | 0.79 | 3,791
Proven & Probable | 207,687 | 0.84 | 5,616 | 13,354 | 0.54 | 232 | 221,041 | 0.82 | 5,848

NOTES:

(1)Thomas L. Dyer, P.E., is the QP responsible for reporting the Batman deposit proven and probable reserves.

(2)Batman deposit reserves are reported using a 0.40 g-Au/t cutoff grade.

(3)Deepak Malhotra is the QP responsible for reporting the heap-leach pad reserves.

(4)Because all of the heap-leach pad reserves are to be fed through the mill, these reserves are reported without a cutoff grade applied.

(5)The reserves point of reference is the point where material is fed into the mill.

15.6In-Pit inferred Resources

Inferred resources were considered as waste and not used in the economic analysis. Note that CIM Standards define inferred resources as:

An inferred mineral resource is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply but not verify geological and grade or quality continuity.

An inferred mineral resource has a lower level of confidence than that applying to an indicated mineral resource and must not be converted to a mineral reserve. It is reasonably expected that the majority of inferred mineral resources could be upgraded to indicated mineral resources with continued exploration.

An inferred mineral resource is based on limited information and sampling gathered through appropriate sampling techniques from locations such as outcrops, trenches, pits, workings and drill holes. Inferred mineral resources must not be included in the economic analysis, production schedules, or estimated mine life in publicly disclosed pre- feasibility or feasibility studies, or in the Life of Mine plans and cash flow models of developed mines. Inferred mineral resources can only be used in economic studies as provided under NI 43-101.

There may be circumstances, where appropriate sampling, testing, and other measurements are sufficient to demonstrate data integrity, geological and grade/quality continuity of a measured or indicated mineral resource, however, quality assurance and quality control, or other information may not meet all industry norms for the disclosure of an indicated or measured mineral resource. Under these circumstances, it may be reasonable for the qualified person to report an inferred mineral resource if the qualified person has taken steps to verify the information meets the requirements of an inferred mineral resource.

Table 15‑9 shows the inferred resources inside of the pit designs for each phase. For statement of reserves and economic analysis, inferred resources in the pit are reported as waste.

Table 15‑9: In-Pit Inferred Resources Inside Base Case Pits

| In-Pit Inferred Resource
Tonnes | g-Au/t | Ozs Au
Ph_1 | 992 | 0.58 | 18
Ph_2 | 3,868 | 0.60 | 75
Ph_3 | 3,048 | 0.57 | 56
Ph_4 | 6,585 | 0.58 | 123
Total | 14,494 | 0.58 | 272

NOTE:

(1)Base Case inferred resources are reported using a cutoff grade of 0.40 g-Au/t

(2)Mineral resources that are not mineral reserves have no demonstrated economic viability and do not meet all relevant modifying factors.

15.7Heap Leach Reserve Estimate

Heap leach reserves are provided in Table 15‑8. In addition to the ore mined from the Batman open pit, the mine plan contemplates processing the 13.4 Mt of ore from the existing heap leach pad through the mill at the end of the mine life.

The bottle roll and column leach test work undertaken at the ALS Metallurgy Laboratory in Australia has been reviewed (ALS, 2013). The testwork indicated the following:

Cyanidation leach tests on “as is” material on the heap will extract ± 30% of the gold.

CIP cyanidation tests at a grind size of P80 of 90 microns will extract on average 72% of gold (range: 64.14% to 80.37%) in 24 hours of leach time. The average lime and cyanide consumptions were 1.75 kg/t and 0.78 kg/t, respectively.

The limited testwork indicates that it is economically feasible to process and recover gold from the heap leach material. Hence, the 13.4 Mt of heap leach ore meets the criteria necessary to be called “reserves” for the Mt Todd Gold Project and should be included in the reserve tabulation based on the following:

 The heap leach material is already mined;

Is readily recoverable using the planned flowsheet; and

Can be economically processed in the plant which will be built to process fresh ore.

These reserves should be considered as probable, since limited drilling and assaying was undertaken to estimate the gold content of the heap leach residues.

16.0MINING METHODS

The Base Case is identified as a 50,000 tpd operation, as presented in this section. Another option, defined as the Alternate Case (33,000 tpd) is presented in Section 24.7 – Alternate Case.

16.1Methods

The Project has been planned as an open-pit truck and shovel operation. The truck and shovel method provides reasonable cost benefits and selectivity for this type of deposit. Only open-pit mining methods are considered for mining at Mt Todd.

The mining method and approach are based on the geotechnical and hydrological data and studies described in Sections 15.1.2 – Slope Parameters, Section 15.2.2 – Pit Design Slopes, and Section 24.5 – Regional Groundwater Model and Mine Dewatering.

16.2Site Landforms and Impoundments

For reference, a description of the site landforms and impoundments, as well as their naming conventions and abbreviations is included as Table 16‑1.

Table 16‑1: Description of Landforms and Impoundments

Landform/Impoundment | Abbreviated Name
Tailings Storage Facility 1 | TSF 1
Tailings Storage Facility 2 | TSF 2
Raw Water Dam | RWD
Low Grade Ore Stockpile | LGOS
Low Grade Ore Stockpile Retention Pond | LGRP
Heap Leach Pad | HLP
Batman Pit | RP3
Process Plant Retention Pond | PRP
Waste Rock Dump | WRD
Waste Rock Dump Retention Pond | RP1
Process Water Pond | PWP
Water Treatment Plant | WTP
Process Plant | PP

16.3Waste Material Definition

Some of the waste material at Mt Todd contains sulfide minerals, which can result in acid generation. Tetra Tech provided MDA with classification criteria for waste material so that the resulting production schedule can include the segregation of waste types for proper handling. Waste was classified into three classes based on total sulfur content as follows:

Non-PAGTotal Sulfur <= 0.25%

UncertainTotal Sulfur > 0.25% and <= 0.40%

PAG Total Sulfur > 0.40%

Material classified as uncertain or potentially acid generating (PAG) material was scheduled so that it could be placed inside of the ultimate waste dump. Non-potentially acid generating material (Non-PAG) was scheduled to be used to encapsulate the uncertain and PAG material, and for reclamation cover and construction material for TSF 1 and TSF 2. Due to the scheduled timing of mining of Non-PAG material, some of this material will be stockpiled on the outer portions of the waste dump and rehandled as required.

16.4Mine-Waste Facilities

Total contained waste tonnage is 523 million tonnes for the Base Case.

For both cases, Non-PAG mine waste will be used for construction and final reclamation cover for the mine. The construction material will be used at the tailings storage facilities (“TSF 1” and “TSF 2”). Other Non-PAG material will be used for reclamation purposes covering tailings and other facilities at the end of the mine life. Sorter tailings will be generated from the process plant sorter and hauled to a temporary stockpile near the sorter. This material is considered Non-PAG and will be rehandled at the end of the mine life as part of the reclamation material.

Tetra Tech provided the amount of material that would be required to be mined for construction and reclamation. These totals are shown in Table 16‑2 for the Base Case.

Table 16‑2: Base Case Construction and Reclamation Requirements

| Units | Yr -1 | Yr 1 | Yr 2 | Yr 3 | Yr 4 | Yr 5 | Yr 6 | Yr 7 | Yr 8 | Yr 9 | Yr 10 | Yr 11 | Yr 12 | Yr 13 | Yr 14 | Total
CONSTRUCTION MATERIAL REQUIREMENTS
Total TSF 1 | K m^3 | 675 | 625 | - | - | 210 | - | 219 | 222 | - | - | 235 | 235 | 367 | - | - | 2,787
| K Tonnes | 1,350 | 1,250 | - | - | 420 | - | 438 | 443 | - | - | 470 | 470 | 734 | - | - | 5,575
Total TSF 2 | K m^3 | - | - | 3,381 | 2,902 | 1,878 | 1,348 | 1,919 | 1,387 | 3,296 | - | 2,935 | 2,904 | 218 | - | - | 22,169
| K Tonnes | - | - | 6,763 | 5,804 | 3,757 | 2,697 | 3,837 | 2,774 | 6,592 | - | 5,869 | 5,808 | 437 | - | - | 44,337
Total - All TSFs | K m^3 | 675 | 625 | 3,381 | 2,902 | 2,088 | 1,348 | 2,137 | 1,609 | 3,296 | - | 3,169 | 3,139 | 586 | - | - | 24,956
| K Tonnes | 1,350 | 1,250 | 6,763 | 5,804 | 4,176 | 2,697 | 4,275 | 3,218 | 6,592 | - | 6,339 | 6,278 | 1,171 | - | - | 49,912
RECLAMATION MATERIAL REQUIREMENTS
Sorter Reject to TSF 1 |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  | 10,771 | 10,771
Sorter Reject to TSF 2 |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  | 9,997 | 9,997
Total Sorter Reject Re-handle |  | - | - | - | - | - | - | - | - | - | - | - | - | - | - | 20,769 | 20,769
TSF 1_Closure |  | - | - | - | - | - | - | - | - | - | - | - | - | - | 135 | 3,750 | 3,885
TSF 2_Closure |  | - | - | - | - | - | - | - | - | - | - | - | - | - | 664 | 3,506 | 4,170
Total Non-PAG |  | - | - | - | - | - | - | - | - | - | - | - | - | - | 799 | 48,793 | 49,593

The mine waste facility has been designed to permanently contain the remaining waste material associated with reserves in the pits for both cases. This facility is an extension of the existing waste dump at site with the ultimate dump fully encapsulating the current dump. The ultimate design incorporates an angle of repose slope of 1.5 vertical to 1.0 horizontal, with catch benches of 8.0 meters every 30 meters in height. During the construction of the ultimate dump, PAG and uncertain waste materials will be dumped in the interior of each lift of the waste dump. Non-PAG material will be dumped to the outer edge of each lift. It is anticipated that at least a 10 meter rind of Non-PAG material will surround uncertain and PAG type waste material.

For closure of the waste rock facility reference Section 20.0 – Environmental Studies, Permitting, and Social or Community Impact.

A 40% swell factor and an average specific gravity of 2.67 (bank) have been assumed for volume calculations. The Base Case dump design has a total capacity to contain 485 million tonnes, but it is over designed and only about 440 million tonnes of the capacity will be used. Prior to filling into the final footprint of the waste dump, the design should be optimized to minimize the footprint and minimize reclamation requirements.

16.5Mine-Production Schedule

Proven and probable reserves and the associated waste material were used to schedule mine production. Inferred resources inside of the pit were considered as waste. The final production schedule uses trucks and shovels necessary to produce the ore required to be fed into the process plant and maintain stripping requirements for each case.

Production scheduling was done using MineSched (version 9.1). This was summarized in Excel spreadsheets where additional waste re-handling was added to the schedule. Table 16‑3 shows the mine production schedule, including re-handle from stockpiles, waste material re-handle, and sorter stockpile material movement requirements for the Base Case. For the purpose of production scheduling, low-grade, medium-grade, and high-grade ore was designated. The medium-grade and high-grade cutoffs used 0.55 and 0.85 g-Au/t, respectively.

Ore from the mine is to be sent from the pit directly to the crusher, or to a mill ore stockpile. During pre-stripping, high-grade, medium-grade, and low-grade ore is stockpiled in the stockpile area northeast of the waste dump facility. High-grade and medium-grade ore is processed in the mill when mill capacity becomes available in Year 1.

For the purpose of scheduling, three ore stockpiles are assumed:

High-grade ore stockpile for high-grade ore (> 0.85 g-Aug/t;

Medium-grade stockpile for medium-grade ore (0.55 to 0.85 g-Au/t); and

Low-grade stockpile for low-grade ore (0.40 to 0.55 g-Au/t).

The high-grade and medium-grade stockpiles are to be built within the low-grade stockpiling areas but will be exhausted during the first year of processing when mill capacity becomes available. During the life of mine, the low-grade stockpile is to be used as needed to feed the mill to full capacity. For this reason, the stockpile grows and shrinks through the life of mine. The maximum stockpile balance through the life of mine is estimated to be 30.5 million tonnes for the Base Case. Ultimate stockpile designs have been created north-east of the processing facility.

Rehandling of stockpiled material will be done using a loader and trucks to haul ore to the crusher. Table 16‑4 shows the yearly ore stockpile balances.

Ore sent to the mill is shown in Table 16‑5 and is a combination of ore shipped directly from the mine, and ore that is reclaimed from stockpiles. Ore sent to the mill is summarized based on the level of oxidation. The recovered ounces shown are based on the recoveries used for pit optimizations.

Table 16‑3: Annual Mine Production Schedule – Base Case

| Units | Yr -1 | Yr 1 | Yr 2 | Yr 3 | Yr 4 | Yr 5 | Yr 6 | Yr 7 | Yr 8 | Yr 9 | Yr 10 | Yr 11 | Yr 12 | Yr 13 | Yr 14 | Total
CONSTRUCTION MATERIAL REQUIREMENTS
Total TSF 1 | K m^3 | 675 | 625 | - | - | 210 | - | 219 | 222 | - | - | 235 | 235 | 367 | - | - | 2,787
| K Tonnes | 1,350 | 1,250 | - | - | 420 | - | 438 | 443 | - | - | 470 | 470 | 734 | - | - | 5,575
Total TSF 2 | K m^3 | - | - | 3,381 | 2,902 | 1,878 | 1,348 | 1,919 | 1,387 | 3,296 | - | 2,935 | 2,904 | 218 | - | - | 22,169
| K Tonnes | - | - | 6,763 | 5,804 | 3,757 | 2,697 | 3,837 | 2,774 | 6,592 | - | 5,869 | 5,808 | 437 | - | - | 44,337
Total - All TSFs | K m^3 | 675 | 625 | 3,381 | 2,902 | 2,088 | 1,348 | 2,137 | 1,609 | 3,296 | - | 3,169 | 3,139 | 586 | - | - | 24,956
| K Tonnes | 1,350 | 1,250 | 6,763 | 5,804 | 4,176 | 2,697 | 4,275 | 3,218 | 6,592 | - | 6,339 | 6,278 | 1,171 | - | - | 49,912
RECLAMATION MATERIAL REQUIREMENTS
Sorter Reject to TSF 1 |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  | 10,771 | 10,771
Sorter Reject to TSF 2 |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  | 9,997 | 9,997
Total Sorter Reject Re-handle |  | - | - | - | - | - | - | - | - | - | - | - | - | - | - | 20,769 | 20,769
TSF 1_Closure |  | - | - | - | - | - | - | - | - | - | - | - | - | - | 135 | 3,750 | 3,885
TSF 2_Closure |  | - | - | - | - | - | - | - | - | - | - | - | - | - | 664 | 3,506 | 4,170
Total Non-Pag |  | - | - | - | - | - | - | - | - | - | - | - | - | - | 799 | 48,793 | 49,593

Table 16‑4: Annual Stockpile Balance – Base Case

|  |  | Pre-Prod | Yr 1 | Yr 2 | Yr 3 | Yr 4 | Yr 5 | Yr 6 | Yr 7 | Yr 8 | Yr 9 | Yr 10 | Yr 11 | Yr 12 | Yr 13
Hg_StkPl | Added | K Tonnes | 4,644 | 25 | 297 | 42 | 3,909 | 125 | - | - | - | 419 | 3,428 | - | - | -
| g-Au/t | 1.29 | 1.02 | 1.14 | 1.01 | 1.22 | 1.11 | - | - | - | 1.06 | 1.09 | - | - | -
| K Ozs Au | 193 | 1 | 11 | 1 | 153 | 4 | - | - | - | 14 | 120 | - | - | -
Removed | K Tonnes | - | 4,669 | - | 339 | - | 4,034 | - | - | - | - | - | 3,847 | - | -
| g-Au/t | - | 1.29 | - | 1.12 | - | 1.21 | - | - | - | - | - | 1.08 | - | -
| K Ozs Au | - | 194 | - | 12 | - | 157 | - | - | - | - | - | 134 | - | -
Balance | K Tonnes | 4,644 | - | 297 | - | 3,909 | - | - | - | - | 419 | 3,847 | - | - | -
| g-Au/t | 1.29 | - | 1.14 | - | 1.22 | - | - | - | - | 1.06 | 1.08 | - | - | -
| K Ozs Au | 193 | - | 11 | - | 153 | - | - | - | - | 14 | 134 | - | - | -
Mg_StkPl | Added | K Tonnes | 3,373 | 125 | 737 | 233 | 5,823 | 618 | - | - | - | 1,164 | 9,732 | - | - | -
| g-Au/t | 0.69 | 0.66 | 0.67 | 0.66 | 0.68 | 0.65 | - | - | - | 0.67 | 0.67 | - | - | -
| K Ozs Au | 74 | 3 | 16 | 5 | 127 | 13 | - | - | - | 25 | 210 | - | - | -
Removed | K Tonnes | - | 2,595 | 902 | 970 | - | 4,713 | 1,728 | - | - | - | - | 10,897 | - | -
| g-Au/t | - | 0.72 | 0.60 | 0.67 | - | 0.71 | 0.60 | - | - | - | - | 0.67 | - | -
| K Ozs Au | - | 60 | 17 | 21 | - | 107 | 33 | - | - | - | - | 236 | - | -
Balance | K Tonnes | 3,373 | 902 | 737 | - | 5,823 | 1,728 | - | - | - | 1,164 | 10,897 | - | - | -
| g-Au/t | 0.69 | 0.60 | 0.67 | - | 0.68 | 0.60 | - | - | - | 0.67 | 0.67 | - | - | -
| K Ozs Au | 74 | 17 | 16 | - | 127 | 33 | - | - | - | 25 | 236 | - | - | -
Lg_StkPl | Added | K Tonnes | 2,420 | 2,539 | 6,762 | 6,305 | 8,425 | 1,097 | - | - | - | 3,300 | 7,033 | - | - | -
| g-Au/t | 0.47 | 0.47 | 0.47 | 0.47 | 0.47 | 0.47 | - | - | - | 0.46 | 0.48 | - | - | -
| K Ozs Au | 37 | 39 | 103 | 95 | 128 | 17 | - | - | - | 49 | 108 | - | - | -
Removed | K Tonnes | - | - | 966 | 2,957 | 1,759 | - | 9,595 | 7,369 | 2,785 | - | - | 160 | 12,291 | -
| g-Au/t | - | - | 0.52 | 0.52 | 0.52 | - | 0.49 | 0.44 | 0.43 | - | - | 0.52 | 0.46 | -
| K Ozs Au | - | - | 16 | 50 | 30 | - | 151 | 105 | 38 | - | - | 3 | 183 | -
Balance | K Tonnes | 2,420 | 4,959 | 10,755 | 14,104 | 20,770 | 21,867 | 12,273 | 4,903 | 2,119 | 5,418 | 12,452 | 12,291 | - | -
| g-Au/t | 0.47 | 0.47 | 0.47 | 0.46 | 0.46 | 0.46 | 0.43 | 0.43 | 0.43 | 0.45 | 0.46 | 0.46 | - | -
| K Ozs Au | 37 | 75 | 162 | 207 | 306 | 322 | 172 | 67 | 29 | 78 | 186 | 183 | - | -
All StkPl | Balance | K Tonnes | 10,437 | 5,861 | 11,790 | 14,104 | 30,502 | 23,596 | 12,273 | 4,903 | 2,119 | 7,002 | 27,195 | 12,291 | - | -
| g-Au/t | 0.91 | 0.49 | 0.50 | 0.46 | 0.60 | 0.47 | 0.43 | 0.43 | 0.43 | 0.52 | 0.64 | 0.46 | - | -
| K Ozs Au | 304 | 93 | 189 | 207 | 586 | 356 | 172 | 67 | 29 | 117 | 555 | 183 | - | -

Table 16‑5: Annual Ore Delivery to the Mill Crusher – Base Case

|  | Pre-Prod | Yr 1 | Yr 2 | Yr 3 | Yr 4 | Yr 5 | Yr 6 | Yr 7 | Yr 8 | Yr 9 | Yr 10 | Yr 11 | Yr 12 | Yr 13 | Total
Sulfide Ore | K Tonnes | - | 17,694 | 17,739 | 17,136 | 17,663 | 17,077 | 16,912 | 17,285 | 17,542 | 17,750 | 17,750 | 17,671 | 12,133 | - | 204,354
g-Au/t | - | 1.20 | 0.83 | 0.96 | 1.00 | 0.96 | 0.56 | 0.56 | 0.61 | 0.78 | 1.24 | 0.84 | 0.46 | - | 0.84
K Ozs Au | - | 681 | 473 | 528 | 569 | 526 | 303 | 310 | 346 | 448 | 709 | 477 | 181 | - | 5,551
Recovery | 0% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 0% | 85%
K Ozs Au Rec | - | 581 | 404 | 451 | 486 | 449 | 259 | 265 | 296 | 382 | 605 | 407 | 155 | - | 4,739
Mixed Ore | K Tonnes | - | 20 | 7 | 564 | 79 | 434 | 518 | 309 | 123 | - | - | 127 | 93 | - | 2,272
g-Au/t | - | 0.75 | 0.52 | 0.70 | 0.62 | 0.73 | 0.51 | 0.44 | 0.42 | - | - | 1.05 | 0.42 | - | 0.62
K Ozs Au | - | 0 | 0 | 13 | 2 | 10 | 9 | 4 | 2 | - | - | 4 | 1 | - | 45
Recovery | 0% | 78% | 78% | 78% | 78% | 78% | 78% | 78% | 78% | 0% | 0% | 78% | 78% | 0% | 78%
K Ozs Au Rec | - | 0 | 0 | 10 | 1 | 8 | 7 | 3 | 1 | - | - | 3 | 1 | - | 35
Oxidized Ore | K Tonnes | - | 36 | 4 | 98 | 9 | 239 | 321 | 205 | 85 | - | - | - | 65 | - | 1,061
g-Au/t | - | 0.72 | 0.53 | 0.84 | 0.53 | 0.77 | 0.52 | 0.44 | 0.42 | - | - | - | 0.42 | - | 0.58
K Ozs Au | - | 1 | 0 | 3 | 0 | 6 | 5 | 3 | 1 | - | - | - | 1 | - | 20
Recovery | 0% | 78% | 78% | 78% | 78% | 78% | 78% | 78% | 78% | 0% | 0% | 0% | 78% | 0% | 78%
K Ozs Au Rec | - | 1 | 0 | 2 | 0 | 5 | 4 | 2 | 1 | - | - | - | 1 | - | 16
Total Ore | K Tonnes | - | 17,750 | 17,750 | 17,799 | 17,750 | 17,750 | 17,750 | 17,799 | 17,750 | 17,750 | 17,750 | 17,799 | 12,291 | - | 207,687
g-Au/t | - | 1.19 | 0.83 | 0.95 | 1.00 | 0.95 | 0.56 | 0.55 | 0.61 | 0.78 | 1.24 | 0.84 | 0.46 | - | 0.84
K Ozs Au | - | 682 | 473 | 543 | 570 | 542 | 317 | 317 | 349 | 448 | 709 | 481 | 183 | - | 5,616
Recovery | 0% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 0% | 85%
K Ozs Au Rec | - | 582 | 404 | 463 | 487 | 462 | 270 | 270 | 298 | 382 | 605 | 411 | 156 | - | 4,790

16.6Equipment Selection and Productivities

Mt Todd has been planned as an open-pit mine using large haul trucks, hydraulic shovels, and front-end loading equipment. Primary mine production is to be achieved using 31-cubic meter hydraulic shovels along with 226-tonne haul trucks, though final equipment selection may differ.

Secondary mine production is to be achieved using 18-cubic meter loaders along with the 226-tonne trucks. Loaders will be used mostly to mine ore from the pit to the crusher, and for reclamation of ore from stockpiles. Some waste production from the loader is anticipated as well.

Table 16‑6 shows the maximum shovel productivity estimate based on scheduled time, availability, and truck and material parameters. This maximum productivity would require that trucks are always available and the shovels are always digging; however, that is not always the case.

In-pit and ex-pit centerlines were drawn for each of the pits and destinations, including the waste dump, crusher, and ore stockpile. As the dump is very large, it was divided into 20 smaller volumes to account for haulage requirements during the life of mine. Truck speeds for each profile were calculated based on published rim-pull curve data. Maximum speed limits were also applied to ensure that safe operating conditions were adhered to and that productivities were achievable.

Bench haulage routes were also drawn for each bench to ensure proper travel on the benches and that truck requirements are properly accounted for. Bench travel speed limits were applied to the profiles for both loaded and empty trucks.

Mine production schedules were run using MineSched (version 9.1) mine scheduling software. The profiles and truck parameters were supplied to MineSched to calculate the productive truck hours required. An efficiency of 83% was used to derive operating hours from the productive hours. This accounts for inefficiencies in the operations that are found between the loading units and the dumping locations. This is similar to a 50-minute working hour.

Incremental truck hours were added to waste haulage to account for waste material hauled to TSF 1 and TSF 2 for construction purposes. Haulage requirements for sorter tailings were estimated within cost sheets using a constant cycle time. The material would be loaded into a truck from a silo and the silo bin is sized to use the mine fleet. It was determined that a single truck would be able to take care of the haulage needs for the sorter.

Loading-unit hours were estimated using 83% efficiency and the production rate for loading equipment. The schedule was constrained using tonnage on a period basis to balance the use of loading and haulage equipment.

Availability was estimated dependent on the age of the piece of equipment. Availabilities start out at 90% and decrement 1% per year until they reach 85%, and then they are kept constant. Availabilities, efficiencies, operating hours, and load and haul equipment requirements are shown in Table 16‑7 for the Base Case.

Table 16‑6: Maximum Loader Productivity Estimate

Description | Unit | All Rock |
MATERIAL PROPERTIES
Material SG (BCM) | t/cm (Wet) | 2.70 |
Material SG (Loose) | t/cm (Wet) | 1.93 |
Material SG (BCM Dry) | t/cm (Dry) | 2.50 |
Material SG (LCM Dry) | t/cm (Dry) | 1.79 |
Swell Factor |  | 1.4 |
DAILY SCHEDULE
Shifts per Day Hours per Shift | shift/day hr/shift | 212 |
Theoretical Hours per Day | hrs/day | 24 |
Shift Startup / Shutdown Lunch Breaks Operational Standby | hrs/shift hrs/shift hrs/shift hrs/shift | 0.50.50.250.25 |
Total Standby / Shift Total Standby / Day | hrs/shift hrs/day | 1.503.00 |
Available Work Hours | hrs/day | 21.00 |
Schedule Efficiency | % | 87.5 |

|  | 31 cm Hyd 226 T Trks | 18 cm FEL 226 T Trks
LOADING PARAMETERS |  |  |
Shovel Mech. Avail. | % | 85% | 85%
Operating Efficiency | % | 83% | 83%
Bucket Capacity | m3 | 31 | 18
Bucket Fill Factor | % | 95% | 95%
Avg. Cycle Time | Sec | 34 | 50
TRUCK PARAMETERS |  |  |
Truck Mech. Avail. | % | 85% | 85%
Operating Efficiency | % | 83% | 83%
Volume Capacity | m3 | 176 | 176
Tonnage Capacity | lt (Wet) | 227 | 227
Truck Spot Time | Sec | 24 | 24
SHOVEL PRODUCTIVITY |  |  |
Effective Bucket Capacity | Cyd | 29.45 | 17.10
Tonnes per Pass – Wet | lst (Wet) | 56.8 | 33.0
Tonnes per Pass – Dry | lst (Dry) | 52.6 | 30.5
Theoretical Passes – Vol | passes | 5.98 | 10.29
Theoretical Passes – Wt | passes | 4.00 | 6.88
Actual Passes Used | passes | 4.0 | 7.0
Truck Tonnage – Wet | wmt/load | 226 | 226

Truck Tonnage – Dry | dmt/load | 210 | 210
Truck Capacity Utilized – Vol | % | 67% | 67%
Truck Capacity Utilized – Wt | % | 100% | 100%
Load Time | min | 2.67 | 6.23
Theoretical Productivity | dst/hr | 4,729 | 2,023
Tonnes per Operating Hour | dst/hr | 3,930 | 1,680
Tonnes Per Day | dst/day | 70,200 | 30,000
Potential – 355 days/year | t/year | 24,921,000 | 10,650,000

Table 16‑7: Annual Load and Haul Equipment Requirements – Base Case

| Units | Pre-Prod | Yr 1 | Yr 2 | Yr 3 | Yr 4 | Yr 5 | Yr 6 | Yr 7 | Yr 8 | Yr 9 | Yr 10 | Yr 11 | Yr 12 | Yr 13 | Yr 14 | Total
HAULAGE REQUIREMENTS |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Productive Hours | Hrs | 37,225 | 66,596 | 109,379 | 160,274 | 198,158 | 197,784 | 194,365 | 191,732 | 193,794 | 193,159 | 195,693 | 29,093 | 14,979 | 8,246 | 30,033 | 1,820,516
Operating Efficiency | % | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 1743%
Operating Hours | Hrs | 44,850 | 80,236 | 131,782 | 193,101 | 238,745 | 238,294 | 234,175 | 231,002 | 233,486 | 232,721 | 235,774 | 35,052 | 18,047 | 9,935 | 36,184 | 2,193,393
Number of Trucks | # | 9 | 17 | 23 | 36 | 38 | 38 | 38 | 38 | 38 | 38 | 38 | 6 | 4 | 4 | 8 | 374
Truck Availability | % | 90% | 89% | 89% | 89% | 88% | 87% | 86% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% |
Available Operating Hours | Hrs | 48,893 | 93,792 | 148,805 | 216,043 | 242,532 | 239,767 | 238,193 | 236,820 | 235,063 | 235,063 | 235,063 | 37,220 | 24,744 | 24,744 | 49,487 | 2,306,227
Use of Available Hours | % | 94% | 86% | 89% | 89% | 98% | 99% | 98% | 98% | 99% | 99% | 100% | 94% | 73% | 40% | 73% | 95%
Tonnes per Operating Hour | t/Hr | 543 | 595 | 452 | 511 | 380 | 497 | 380 | 311 | 264 | 243 | 251 | 688 | 1,048 | 944 | 775 | 393
HYDRAULIC SHOVEL USAGE |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Number of Shovels | # | 2 | 3 | 3 | 4 | 4 | 4 | 3 | 3 | 3 | 3 | 2 | 2 | - | - | 1 | 3.2
Availability | % | 90.0% | 89.3% | 88.6% | 88.1% | 87.3% | 86.3% | 85.7% | 85.3% | 85.0% | 85.0% | 85.0% | 85.0% | 0.0% | 0.0% | 85.0% | 86.5%
Operating Efficiency | % | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 0% | 0% | 83% | 83.0%
Available Operating Hrs | Op Hrs | 11,439 | 15,629 | 19,347 | 24,084 | 25,400 | 25,109 | 18,703 | 18,683 | 18,558 | 18,558 | 12,372 | 12,407 | - | - | 12,372 | 232,659
Tonnes Mined | K Tonnes | 27,287 | 37,943 | 52,425 | 83,427 | 85,519 | 100,732 | 73,080 | 63,772 | 58,765 | 54,312 | 48,698 | 2,942 | - | - | 21,018 | 709,921
Operating Hours | Op Hrs | 3,122 | 9,667 | 13,357 | 21,255 | 21,788 | 25,664 | 18,619 | 16,247 | 14,972 | 13,837 | 12,407 | 750 | - | - | 5,355 | 177,038
Use of Available Operating Hours | % | 27% | 62% | 69% | 88% | 86% | 102% | 100% | 87% | 81% | 75% | 100% | 6% | 0% | 0% | 43% | 76%
FRONT END LOADERS |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Number of Loaders | # | 1 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 1 | 2 | 2 | 2 | 2 | 1 | 1.9
Availability | % | 85% | 90% | 89% | 88% | 87% | 86% | 86% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 86%
Operating Efficiency | % | 78.7% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0%
Available Operating Hrs | Op Hrs | 3,210 | 10,314 | 13,015 | 12,900 | 12,718 | 12,519 | 12,445 | 12,407 | 12,372 | 12,372 | 12,372 | 12,407 | 12,372 | 12,372 | 12,372 | 176,164
Tonnes Mined | K Tonnes | 525 | 9,779 | 8,130 | 15,172 | 5,322 | 17,785 | 15,988 | 8,013 | 2,785 | 2,263 | 10,464 | 21,181 | 18,921 | 9,376 | 7,006 | 152,710
Operating Hours | Op Hrs | 377 | 7,013 | 5,831 | 10,880 | 3,817 | 12,755 | 11,466 | 5,747 | 1,997 | 1,623 | 7,504 | 15,190 | 13,569 | 6,724 | 5,025 | 109,516
Use of Available Operating Hours | % | 12% | 68% | 45% | 84% | 30% | 102% | 92% | 46% | 16% | 13% | 61% | 122% | 110% | 54% | 41% | 62%

16.7Mine Personnel

Mine personnel estimates include both operating and mine-staff personnel. Operating personnel are estimated as the number of people required to operate trucks, loading equipment, and support equipment to achieve the production schedule. Mine staff is based on the people required for supervision and support of mine production. The mine staff organizational chart is shown in Figure 16‑1. The estimated number of mine personnel required to execute the mine plan is shown in Table 16‑8 for the Base Case.

Salaries for each position were estimated based on information received from Vista. Salaries include an allowance for benefits at a factor of 27% of the base salary for each position. Note that the mine personnel do not include contractors. Vista anticipates using a Maintenance and Repair Contract (“MARC”) to maintain the mining fleet during the first three years. After that time, Vista will operate all maintenance crews. For the purpose of costing, the MARC costs were reduced to take into account savings by lowering contractor’s overhead. Maintenance foremen were added to personnel along with another planner starting in year 3 as part of the maintenance responsibility takeover. However, since the maintenance cost used includes labor, the mechanics are not reflected in the total count for personnel. This would add approximately 80 mechanics, servicemen, and welders to the Base Case.

Figure 16‑1: Mine Organizational Chart

Table 16‑8: Mine Personnel Requirements – Base Case

| Pre-Prod | Yr 1 | Yr 2 | Yr 3 | Yr 4 | Yr 5 | Yr 6 | Yr 7 | Yr 8 | Yr 9 | Yr 10 | Yr 11 | Yr 12 | Yr 13 | Yr 14
MINE OVERHEAD |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Mine Manager | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | - | -
Mine Clerk | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | - | -
Mine Shift Foremen | 6 | 6 | 6 | 6 | 6 | 6 | 6 | 6 | 6 | 6 | 6 | 6 | 6 | 6 | 6
Mine Trainer | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1
Blaster | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | - | -
Blaster's Helper | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | - | -
MINE PRODUCTION |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Loading Operators | 6 | 10 | 13 | 15 | 16 | 18 | 15 | 14 | 12 | 11 | 11 | 9 | 6 | 6 | 6
Haul Truck Operators | 30 | 44 | 69 | 101 | 114 | 114 | 114 | 114 | 114 | 114 | 114 | 45 | 45 | 45 | 45
Drill Operators | 16 | 20 | 29 | 44 | 46 | 48 | 34 | 29 | 29 | 30 | 33 | 3 | - | - | -
Support Equipment Operators | 15 | 15 | 17 | 17 | 17 | 17 | 21 | 21 | 21 | 21 | 21 | 21 | 18 | 6 | 6
Total Mine Operating | 80 | 102 | 141 | 190 | 206 | 210 | 197 | 191 | 189 | 189 | 192 | 91 | 82 | 64 | 64
MINE MAINTENANCE |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Maintenance Superintendent | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1
Maintenance Foremen | - | - | - | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 1 | 1
Light Vehicle Mechanics | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 1 | 1
Tiremen | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2
Shop Laborers | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2
Maintenance Planner | 1 | 1 | 1 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 1 | 1
Service, Fuel, & Lube | 6 | 6 | 6 | 6 | 6 | 6 | 6 | 6 | 6 | 6 | 6 | 6 | 6 | 3 | 3
Total Mine Maintenance | 14 | 14 | 14 | 18 | 18 | 18 | 18 | 18 | 18 | 18 | 18 | 18 | 18 | 11 | 11

ENGINEERING |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Chief Engineer | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | - | - | -
Mine Surveyors | 1 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 1 | 1 | 1
Surveyor Helper | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 1 | 1 | 1
Mine Engineer | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 1 | 1 | 1
Total Engineering | 7 | 8 | 8 | 8 | 8 | 8 | 8 | 8 | 8 | 8 | 8 | 8 | 3 | 3 | 3
MINE GEOLOGY |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Chief Geologist | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | - | - | - | -
Ore Control Geologist | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | - | -
Sampler | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | - | -
Total Geology | 5 | 5 | 5 | 5 | 5 | 5 | 5 | 5 | 5 | 5 | 5 | 4 | 4 | - | -
TOTAL MINE OPERATIONS WORKFORCE
Mine Operations | 80 | 102 | 141 | 190 | 206 | 210 | 197 | 191 | 189 | 189 | 192 | 91 | 82 | 64 | 64
Mine Maintenance | 14 | 14 | 14 | 18 | 18 | 18 | 18 | 18 | 18 | 18 | 18 | 18 | 18 | 11 | 11
Engineering | 7 | 8 | 8 | 8 | 8 | 8 | 8 | 8 | 8 | 8 | 8 | 8 | 3 | 3 | 3
Geology | 5 | 5 | 5 | 5 | 5 | 5 | 5 | 5 | 5 | 5 | 5 | 4 | 4 | - | -
Total | 106 | 129 | 168 | 221 | 237 | 241 | 228 | 222 | 220 | 220 | 223 | 121 | 107 | 78 | 78

17.0RECOVERY METHODS

The Base Case is identified as a 50,000 tpd operation, as presented in this section. Another option, defined as the Alternate Case (33,000 tpd) is presented in Section 24.7 – Alternate Case.

The key criteria used in the process design of the Process Plant have been largely derived from metallurgical testwork and, where appropriate, have been provided by Vista, or based on TTP experience and industry norms. The design criteria and flowsheet development are discussed in this section.

17.1Process Design Criteria

Detailed process design criteria have been developed for the Project. The nominal headline design criteria are listed as follows:

Table 17‑1: Headline Design Criteria

| Unit | 50,000 tpd
Annual Ore Feed Rate | Mt/a | 17.75
Operating Days per Year | d/a | 355
Daily Ore Feed Rate | t/d | 50,000
Crushing Rate (6,637 hours per year availability) | tph | 2,674
HPGR Rate (7,838 hours per year) | tph | 2,264
Ore Sorting Rate (7,838 hours per year) | tph | 408
Milling Rate (7,838 hours per year) | tph | 2,055
Gold Head Grade | g/t | 0.82
Copper Head Grade | % | 0.055
Cyanide Soluble Copper | % | 0.0024
Bulk Density | t/m 3 | 2.76
Primary Grind P80 to Secondary Grind | µm | 250
Grind P80 to Leach | µm | 60
Gold Recovery | % | 85
Gold Production (average) | oz/d | 1,120
Gold Production (average) | oz/a | 397,761

The testwork results collated from the 2011 and 2012 testing campaigns and additional metallurgical and process test work conducted in 2016/2017, together with the process design criteria, were used to develop the process flow sheet and mass balance.

17.2Flow Sheet Development

A schematic diagram of the process flowsheet is presented in Figure 17‑1.

17.2.1Crushing Modeling

Impact crush work index (CWi) tests were performed on eighty individual samples from the 2011 drill cores. The CWi values ranged from 3.2 kilowatt hours per tonne (kWh/t) to 26.5 kWh/t. For design purposes, a CWi of 20 kWh/t was selected, 75% of the maximum.

Unconfined compressive strength (UCS) was measured on 16 samples. The values ranged from 14 megapascals (MPa) (med strong) to 183 MPa (very strong). Eighty percent of the results were in the strong to very strong designation of ore hardness.

The run of mine ore from the pit has a maximum particle size of 1000 mm and F80 to the primary crusher of 400 mm. Two stages of crushing, primary and secondary are required to reduce particle size to a P80 of 31.5 mm, required as feed to the HPGR tertiary crushers. A single gyratory crusher is sized for the primary duty reducing ore size to a nominal P80 of 130 mm. Two secondary cone crushers operating in parallel and in closed circuit with two sizing screens cutting at 40 mm, are used to produce the feed to the HPGRs at product size P80 of 31.5 mm.

Figure 17‑1: Simplified Process Flow Diagram

17.2.2Primary Crusher

The primary crusher power was calculated using both the FL Smidth gyratory calculation model and the Metso Bruno model. Using the CWi of 20 kWh/t and a fall through percentage of zero to simulate peak conditions, both models provided peak primary crushing power requirements of a nominal 574 kW and 576 kW respectively to reduce a feed F80 of 400 mm to a product P80 of 130 mm.

17.2.3Secondary Crushers

Secondary crushing with closed circuit screening was modelled by FL Smidth. Two Raptor 1300 cone crushers operating in parallel are used to reduce the primary crusher product to a final product P80 of 31.5 mm, for the 50,000 tpd Base Case.

17.2.4HPGR

HPGR power requirements to reduce the HPGR feed to a final product P80 of 3.25 mm was shown by the Polysius testwork to be 1.9 kWh per tonne of feed to the HPGR. The feed to the HPGR is the sum of new feed plus the recirculating load screen oversize material, less Ore Sorting reject. The total feed to the HPGR is two times the fresh feed rate. HPGR testwork supported vendor recommendations for two HPGR Polycom PM8‑24/17, each equipped with 2 x 2,650 kW drives.

17.2.5Ore Sorting

The coarse fraction (plus 16mm) from the HPGR will be sent to the ore sorting equipment for separating the gold-bearing sulfide minerals and quartz veining from non-gold bearing waste material.

17.2.6Grinding Modeling

A variety of internal models were used to provide the initial baseline ball mill power requirements and vendors were approached for proposals. The most price competitive and technically acceptable submission was then selected for further interaction, with the vendor calculations compared against internal calculations. The circuit incorporates two dual pinion drive ball mills.

17.2.7Thickener / Leach / CIP Design

THICKENER

Thickener design was performed by Outotec based on the testwork results. Outotec’s recommendation included a 45 m thickener for the Base Case.

Maximum compaction of 63% solids was achieved at a bed depth of 240 mm, there would be no difficulty achieving the target of 62% solids in the underflow and with a clear overflow.

Additional thickening test work conducted by Pocock Industrial in 2017 produced similar results with the Pre‑Leach Thickener requiring a nominal 45m for the Base Case. The Tails Thickener is of a similar size, however requires less flocculant at 20 g/t with the Pre-Leach Thickener requiring 45 g/t.

LEACH AND ADSORPTION

The optimum leach / adsorption density as determined by SPX testwork was 55% solids for the previous grind size of 90 um. This is changed to 45% solids with recycled cyanide rich process liquor for two main reasons:

1)Current grind size of 60 µm would result in an excessive viscosity in 55% slurry; and

2)To maximize recovery of Cyanide.

The leach and adsorption circuits were modelled. A six stage adsorption is required to minimize solution losses. Dissolved gold in residue solution will be ≤0.010 ppm.

At the planned gold head grade, the system will produce a loaded carbon head grade of approximately 1,250 g/t, and carbon movement requirements will be in the order of 30 tpd for the Base Case.

17.3Description of Process Areas

17.3.1Area 3100 – Crushing Circuit Availabilities

The crushing circuit availabilities coupled with the ore crusher work index are the two predominant factors in sizing crusher circuits. Rather than assuming a standard availability of between 70% and 75%, a review of the previous primary crusher operations at Mt Todd was conducted. When removing the downtime periods when the crushing system was not required, the average availability for the remaining duration was 59%.

Additionally, TTP reviewed results from a two-year study and dynamic simulation of a large scale crusher operation in the tropics, which indicated the downtime was apportioned as follows:

Dump hopper empty 19.2% (mining not keeping up)

Cannot discharge 15.6% (downstream equipment interruptions)

Operating Breakdown 0.6% (crusher specific)

Mechanical breakdown 1.2% (crusher specific)

Electrical breakdown 2.3% (crusher specific)

Planned maintenance 2.5% (crusher specific)

The combination of this data coupled with the historical Mt Todd crusher downtime led to an initial crusher circuit availability of 60% being selected, with first pass crushing equipment initially being selected on this basis.

Subsequently it was agreed with the mining design consultant MDA that the costs of an extra loader and build of an emergency stockpile on the ROM pad be included and to remove the downtime attributable to mining lack of supply in its entirety.

This resulted in an availability of 75.8%, 6,637 operating hours per year.

17.3.1.1Crushing Circuit Design

The crushing circuit was chosen based on reliability and similarity to existing mining operations. It consists of a single Primary Crusher in an open loop configuration and two Secondary Crushers in parallel in a closed loop configuration with sized output conveyed to a buffer stockpile, providing three days live capacity. The primary and the secondary crushers discharge onto a common conveyor that feeds the Coarse Ore Screens. This configuration allows reduced conveyor footprint and maximum plant productivity.

The Coarse Ore Screens will be fed by vibrating feeders, which regulate the flow from the feed bins. This arrangement maximizes the efficiency of the screens by ensuring full coverage of screen decks at controlled bed depth. Crusher area dust is controlled by dust collection at the screens and dust suppression in all other dust generating areas.

17.3.2Area 3200 – Coarse Ore Stockpile, Reclaim, HPGR and Ore Sorting

A plant availability factor of 89.5% of 365 days/year has been used for the HPGRs and subsequent downstream processes, which is 7,838 operating hours per year. HPGR availability in large hard rock applications ranges from 89% to 92%, with some operations reporting periods of 95% availability when roll change has not been required (Boddington). It is considered appropriate to use a conservative availability factor of 89.5% of the annual 8,760 hours for Mt Todd ore due to its ore hardness.

The coarse ore stockpile will have approximately three days of total capacity between the secondary crushers and the HPGRs, with 23% of that total capacity representing the live volume. Ore will be removed from beneath the Coarse Ore Stockpile by two Apron Feeders.

Two HPGRs will operate in parallel to process for the Base Case. The HPGRs are protected from tramp metal by installation of metal detectors on feed conveyors.

A common HPGR Product Conveyor will receive the discharge from the HPGRs and convey the material to the Fines Screens Feed Bins. The HPGR fines screens are double decked, cutting at nominal 4.5 mm to produce an underflow product at P80 of 3.2 mm and 16mm to produce a mid and an oversize. The screens operate as wet screens with high pressure spray water applied to the decks to assist with screen efficiency. The screen mid material (+4.5mm-16mm), <3-5% moisture, will be conveyed back to the HPGR feed bins and the screen oversize (+16mm) material will be conveyed to Ore Sorting.

Ore Sorting receives a nominal 408 t/h for the Base Case. Ore Sorting comprises two stages, XRT and Laser sorting. The two stages together reject 210 t/h for the Base Case, representing approximately 10% of plant feed.

The above reject performance and nominal gold loss was derived from Outotec bulk ore sorting test work. Gold lost to Ore Sorting reject is minor at a nominal 1.3% of gold entering plant.

17.3.3Area 3300 – Grinding and Classification

Two Ball mills will be used for the Base Case. The parallel Ball Mill circuits are in a conventional configuration. Fresh feed from the fines screens underflow will gravitate to the mill discharge hopper and will be pumped together with the mill discharge slurry to the Cyclones. The cyclone underflow will gravitate to the ball mill feed. The overflow will gravitate to the Secondary Grind feed sump. The cyclone overflow will be pumped to the pre-leach thickener and the underflow will be sent to the ISA mills for secondary grinding.

An automated ball charging system will be provided to deliver approximately 15 tonnes of balls per day to each mill.

17.3.4Area 3400 – Pre-Leach Thickening, Leach Conditioning, Leach and CIP

In order to achieve the required 45% solids feed to the leach and CIP tanks, a Pre-Leach Thickener will be used to thicken to 62% solids with cyanide rich liquor from the tails thickener being used to dilute to 45% solids, thereby minimizing cyanide usage.

Two Leach Conditioning stages will be incorporated ahead of Leach. These tanks will be sized to deliver a total residence time of 4 hours. In these stages, the ore is treated with lime which inhibits reaction of cyanide with pyrites and pyrrhotites by way of forming a lime coating around these gangue components.

The leach and adsorption tanks will be sized to deliver a total residence time of 24 hours for Leach and 6 hours for adsorption, as determined by the test work. For both cases, leach and adsorption will consist of eighteen mechanically agitated tanks in total. There will be twelve leach tanks and six adsorption tanks.

In order to maximize the gold adsorption kinetics, lead nitrate will be added, oxygen will be provided by sparging compressed air into the leach tanks.

Each Leach and CIP Tank can be bypassed for maintenance purposes. Carbon will be regularly pumped upstream from downstream CIP Tanks in a conventional counter-current configuration. The adsorption tanks will be equipped with Kemix interstage carbon screens. The pumping screens will be used to generate the overflow head required for downstream slurry advance.

Carbon Safety Screens will catch any fugitive carbon from the tails slurry. Usable carbon will be returned to the circuit, undersize carbon will report directly out of the circuit via detoxification and tails.

17.3.5Area 3500 – Desorption, Goldroom and Carbon Regeneration

Loaded carbon will be acid washed in an Acid Wash Column, then stripped of copper and gold in an Elution Column. Cold cyanide wash will be used to strip adsorbed copper prior to hot caustic cyanide wash to strip gold. Acid wash effluent and copper wash effluent will be pumped to the detox tanks. The elution and electrowinning process will be the Anglo American Research Laboratories configuration. Eluant will be pumped through the column, heated to 120 deg°C, and collected as loaded eluate in one of two eluate tanks. The desorption circuit will be batch and will take up to 8 hours. The columns are sized to ensure that two elution batches can be performed in a day. After the elution is completed and the carbon is stripped of its gold to about 10 g/t Au, the eluate will be processed through the electrowinning circuit for deposition of gold onto cathodes. The electrowinning circuit will be batch and take up to 8 hours, or until the gold in solution reduces to less than 10 ppm.

The Goldroom consisting of Electrowinning, Drying and Smelting facilities will be supplied as a vendor package. Stripped carbon will be regenerated using an indirect heated horizontal rotary kiln.

17.3.6Area 3600 – Detoxification and Tailings

CIP tailings will be thickened to 62% to maximize recovery of cyanide. This liquor will be used to dilute the Leach circuit to 45% solids. Thickener underflow is then pumped to detoxification.

Two Detoxification Tanks in series will be used to minimize short-circuiting and sized to ensure the required residence time of one hour is achieved.

The second Detox Tank will cascade overflow to a Tailings Pump Hopper from where the tailings will be pumped to the Tailings Storage Facility. Future booster pumps will be required once the second tailings facility is operational. A duty/standby configuration of pumps will be used to ensure continuous operation.

17.3.7Area 3700 – Reagents

Sodium Meta Bi-Sulfite (SMBS) will be delivered to site as a 95% pure solid powder in sea containers. A container tipper and solids handling equipment will transfer the powder from the storage containers to the mixing tank. SMBS will be mixed to 20% w/v in solution and dosed to the detoxification tanks via duty/stand-by dosing pumps. Dust extraction equipment is present at all transfer points of the solids handling and the area where solids handling takes place will be well ventilated. The SMBS solution will have storage for 3 days of nominal usage.

The Sodium Cyanide for Leaching and Elution will be delivered as briquettes in a bulk tanker. For the Base Case, the sodium cyanide will be consumed at a rate of 21 tpd. The solids will be dissolved in the tanker and cyanide solution will be transferred into a mixing tank to ensure full dissolution. The cyanide solution will then be transferred into four storage tanks allowing seven days nominal capacity. There will be a secured and covered facility to store cyanide briquettes in bulk bags on site as emergency storage.

The Hydrochloric Acid (HCL) for the Acid Wash Column will be delivered as a 33% HCl solution and will have storage for 20 days of nominal usage.

Lime will be delivered as 92% activity quick lime powder in road tankers. The lime will be pneumatically transferred to storage silos approximately 200 tonnes capacity. Lime will be slaked on a daily basis. Milk of lime will be distributed from a lime surge tank to Leach.

Sodium Hydroxide (NaOH) will be delivered as a powder in bulk bags and mixed to produce a 50% NaOH solution. Sodium hydroxide is only consumed periodically and therefore does not require an additional storage tank beyond the mixing tank. A nominal 20 day dry solids storage capacity was included into the design.

The lead nitrate for the leach circuit will be delivered as a powder in bulk bags and mixed to produce a 20% solution. The lead nitrate solution will have storage for seven days of nominal usage.

17.3.8Area 3800 – Process Plant Services

Approximately 800 m 3/h of medium pressure process air will be used to service the air requirements for leach and adsorption for the Base Case. Detoxification will be serviced by medium pressure air blowers at a consumption rate of 4,813 m 3/h for the Base Case. High pressure compressors are used to provide plant and instrument air.

Raw water will be supplied via the Raw Water Dam and will service the process water, fire water and gland seal water requirements. Raw water will also service the water treatment plant for potable water required at the mining facilities, process plant, powerhouse and camp. For the Base Case, the nominal raw water consumption will be ~800 m 3/h and will occasionally peak at 2,200 m 3/h during the dry season.

Process plant water will be predominantly made-up of tailings decant return water and raw water. Process water will be used for dilution and density control in the grinding circuit.

17.4Process Water

The water reticulation system for the process plant will consist of the following:

Raw water supply;

Potable water supply;

Fire water supply;

Gland service water supply; and

Process water supply.

Raw water will be delivered from the raw water dam (RWD) to the 9,600 m 3 process plant raw water tank. This water will be used as make-up water for the process water supply, emergency firefighting supply, gland seal, dust suppression, plant clean-up hosing stations, powerhouse, mining facilities and water for the reagents make-up.

The fire water supply will be drawn from the reserve in the raw water tank providing water to the plant site fire water distribution system.

Gland service water for the main plant site will be drawn from the raw water tank. It will be used to supply gland service water for slurry pumps in the plant.

The process water system will include a 9,500 m 3 storage tank. Process water will be supplied to the plant via centrifugal pumps, one operating and one stand-by unit. This water supply will be used for process stream dilution and for use as spray water for the screens. The pre-leach thickener, tailings dam decant water and raw water all report to the process water tank.

17.4.1Process Compressed Air

The plant and instrument air supply systems for the process plant will consist of high pressure compressed air units in the following locations:

Primary Crushing (duty only);

Reclaim Tunnel (duty only);

HPGRs (duty only);

Grinding and Classification (duty/stand-by); and

Leach and CIP (duty/stand-by).

Twin-screw compressors at each location will supply plant air and instrument air to the buildings in which they are located. The air discharging from each compressor will be fed to a plant air receiver and distributed throughout the building. An off-take from the discharge of the plant air receiver will be dedicated to instrument air which will pass through a refrigerant dryer with pre and post filters to an instrument air receiver. This air will be used for instrument air purposes with the required air quality achieved. The remainder of the air generated by the compressors will be used for general plant air duties. The dry areas of the plant will only have a single duty compressor due to the limited requirement of plant and instrument air whereas the wet plant areas will have a duty/standby arrangement.

A dedicated low pressure compressed air system in a duty/stand-by arrangement will be located in the CIP area of the plant for process air in the leach and CIP tanks. The CIP process compressors will deliver air at the required pressure and flow for injection into the leach and CIP tanks.

Similarly, a dedicated low pressure blower air system in a duty/stand-by arrangement will be located in the cyanide detoxification area of the plant for process air in the cyanide detoxification tanks.

17.5Plant Mobile Equipment

The plant mobile equipment will be as follows:

Table 17‑2: Mobile Equipment for Process Plant

Light Vehicles | Quantity
Landcruiser wagon | 2
Dual cab Ute | 10
Tray top Ute | 9
Troop carrier (ambulance) | 1
Bus/troop carrier (15-seat) | 1
Coach | 3
Subtotal | 26
Process Plant Mobile Equipment | Quantity
Loader – Cat 966G | Allowed for in mining
Tool Carrier – Cat IT28 | 1
Bob Cat – Mustang Case | 1
Crane – 15-t Franna | 1
Hiab Truck – 7-t | 1
Service Truck – 2-t | 1
2-t Forklift – allowance | 2
25-t Container Forklift | 1
80-t Crane | 1
Subtotal | 9

18.0PROJECT INFRASTRUCTURE

The Base Case is identified as a 50,000 tpd operation, as presented in this section. Another option, defined as the Alternate Case (33,000 tpd) is presented in Section 24.7 – Alternate Case.

18.1Facility 2000 – Mine

The following section provides a description of the Mine Support Facilities and Mine Support Services that have been developed to support the mining activities.

18.1.1Area 2300 – Mine Support Facilities

Area 2300 Mine Support Facilities consists of the buildings and services for the maintenance and repair of the mine vehicle fleet including Heavy Vehicles (HV). The area is located along the haul road adjacent to the proposed stockpile, between the Pit and the existing Tailings Storage Facility.

18.1.1.1Sub-Area 2305 – Support Facilities – HV Workshop/Warehouse

The Workshop Facility will consist of six Dome Shelter structures mounted on sea containers with concrete floors. The sea containers come equipped as Site Offices, Store Services, Store Consumables, Equipment Repair and Lube Storage and Dispensing facilities for the maintenance and servicing of HVs that are used for mining operations.

The Workshop will be approximately 85.6 m by 24.4 m and sized to service Caterpillar 793F mining trucks.

The Warehouse Facility will consist of one Dome Shelter structure mounted on sea containers with a concrete floor. The Warehouse Facility will be approximately 21.7 m by 24.4 m in size. The sea containers come equipped as Site Offices, Rigging Container, Equipment Repair Workshop and Stores Consumable Container for the storage of parts, components, spares and the like, used by the HV Workshop for vehicle repair.

The HV Workshops and Warehouse Facilities will be complete with all services including power, lighting, communications, lubes, compressed air, water, specialist equipment and other services necessary for the maintenance of the mine vehicle fleet.

The Dome Shelters will be constructed of steel frame and tensile fabric with a fabric life expectancy of 10 years.

A mobile crane will be used externally to the Dome Shelters for the lifting and removal of vehicle parts.

18.1.1.2Sub-Area 2310 Support Facilities – Bulk Fuel Storage

The Bulk Fuel Storage will consist of the relocated 600 kL tank complete with six new bowsers for dispensing into the HV fleet and one new 110 kL self-bunded diesel fuel tank complete with one bowser for dispensing into the LV’s.

The new 110 kL tank will be utilized for refueling the Mine fleet when the 600 kL tank is not in service and will be located adjacent to the HV Workshop.

18.1.1.3Sub-Area 2315 – Support Facilities – HV Washdown

The HV Washdown Facility will primarily be used for washing down the body and undercarriage of heavy vehicles prior to entering the HV Workshop. The facility will consist of a single bay with raised platforms with stair access to four manually operated high pressure water cannons. The run-off water will be connected to the oily water separator and will include drive-in concrete sumps and pits for waste water storage and recovery. The entire facility’s footprint is 18 m by 21.8 m and is sized to service Caterpillar 793F mining trucks.

18.1.1.4Sub-Area 2320 – Support Facilities – Crib / Ablutions / Lockers

The Crib / Ablutions / Lockers Facilities will be a transportable building used by mining personnel and is located adjacent to the HV Workshop. The building will include the necessary system furniture. The Crib Area will also double as a Pre Start Area.

The building will be sized to serve 75 people per shift and cover approximately 19.8 m by 14.4 m.

18.1.1.5Sub-Area 2325 – Support Facilities – HV Tire Change

The Tire Change Facility will consist of one Dome Shelter mounted on sea containers with a concrete floor. The sea containers come equipped as Tire Change Workshop and Store Consumables for the maintenance and changing of HV tires.

The Tire Change Facility will be approximately 26.9 m by 18.1 m and sized to service Caterpillar 793F mining trucks.

The Tire Change Facility will be complete with services including power, lighting, communications, compressed air, water, specialist equipment and other services necessary for the changing of tires.

The Dome Shelter will be constructed of steel frame and tensile fabric with a fabric life expectancy of 10 years.

18.1.1.6Sub-Area 2335 – Support Facilities – Lube Storage

The Lube Storage Facility will consist of a bunded concrete slab for the storage of Intermediate Bulk Containers (IBCs) containing oils and lubricants for the servicing of HVs. The Lube Storage Facility will be located in-between the HV Workshop and the Fuel Storage Facility. Full IBCs will replace containerized IBCs within the Workshops. Lube will be distributed manually. Used oil will be collected in a designated area for approved recycle/disposal.

18.1.1.7Sub-Area 2340 – Support Facilities – ANFO / Magazine Facility

The Ammonium Nitrate Fuel Oil (ANFO) facility is capable of distribution of 10,000 tpa. It is a secure compound for the Ammonium Nitrate (AN), Ammonium Nitrate Emulsion (ANE) and diesel fuel.

The facility includes an area for AN storage, concrete hardstand for AN transfer to a Mobile Process Unit (MPU) and containment pond for spill material.

The ANE is tank stored on concrete plinths with air compressor and pumps for in-loading and out-loading of emulsion.

The diesel is stored in a 110 kL self-bunded tank and includes a spill containment unit.

Magazine storage will consist of two secured modified shipping containers for the storage of detonators, accessories and explosives. The magazines are located adjacent to the ANFO Facility and are surrounded by earth bunding and secure fencing.

The MPU will be used to transport, mix and deliver ANFO to the mine.

A transportable building will be provided to include Office / Crib / Ablution facilities at the Site for driver and delivery personnel.

The ANFO facility footprint is approximately 84.7 m by 128.5 m, excluding the diesel tank.

18.1.1.8Sub-Area 2345 – Support Facilities – Mining Offices

The Mining Offices will be a transportable building used by mining personnel and is located adjacent to the HV Workshop. The building will include a kitchen, ablutions, cellular and open planned offices, meeting rooms, training spaces and necessary system furniture.

The footprint of the Mining Offices is approximately 24 m by 16.5 m and will be sized to account for 25 people.

18.1.1.9Sub-Area 2355 – Support Facilities – Core Shed

The Core Storage Facility will consist of one Dome Shelter mounted on sea containers with a sealed asphalt floor for the storage of core samples at the mine support area.

The sea containers will be equipped with racking for additional storage.

The Core Storage facility will be complete with power and lighting and located north-east of the mining offices.

The Dome Shelter will be constructed of steel frame and tensile fabric with a fabric life expectancy of 10 years.

The Core Storage Facility has a footprint of approximately 48.8 m by 47.3 m.

18.1.2Area 2400 – Mine Support Services

Mine Support Services consists of the services for the Mine Support Facilities.

18.1.2.1Sub-Area 2410 – Support Services – Potable Water

Potable water will be provided to the Mine Support Facilities from the Process Plant Area via pipework in common services trenching.

18.1.2.2Sub-Area 2420 – Support Services – Raw Water

Raw water will be provided to the HV Washdown storage tank at the Mine Support Facilities via a connection from the raw water pipework running along the existing haul road to the Process Plant Area.

18.1.2.3Sub-Area 2430 – Support Services – Fire Water

The Fire Water Main will be provided to the Mine Support Facilities and camps from the Process Plant Area via pipework in common services trenching. Fire hydrants will be provided at required locations.

18.1.2.4Sub-Area 2440 – Support Services – Air

Compressed air will be provided at the HV Workshop and HV Tire Change facilities via suitably sized standalone air compressors and receivers.

18.1.2.5Sub-Area 2450 – Support Services – Power

Power will be provided to the Mines Support Facilities via a connection from the 33 kV overhead power line running past the site into a kiosk substation. From the kiosk, 400V/230V power will be reticulated to all required buildings and services in common services trenches.

18.1.2.6Sub-Area 2450 – Support Services – Communications

Communications will be provided to the Mine Support Facilities from the Process Plant Area via a fiber optic cable in the overhead power lines and will terminate into a server room within the Mine Offices. Cat 6 cables will be reticulated to required building and services.

18.2Facility 4000 – Project Services

This section details the supply and distribution of services outside the process plant.

18.2.1Area 4100 – Water Supply

Area 4100 covers the water supply to the process plant and between facilities.

18.2.1.1Sub-Area 4110 – Water Treatment Plant (WTP)

A Water Treatment Plant will be fed with a combination of decant return, runoff pond water and pit dewatering discharge at a nominal rate of 500 m³/hr.

18.2.1.2Sub-Area 4120 – Raw Water

The raw water requirement for the Base Case will be approximately 30,000 m 3/day, fluctuating due to current operations and weather. The existing line from the Raw Water Dam will be supplemented with an additional 250 mm poly line approximately 4 km in length in order to handle the increased raw water requirements of the higher throughput. This would run parallel to the existing 400 mm poly line.

Raw water will be supplied to the mine support facilities via a one km supply line to a storage tank in that facility. Raw water will be supplied to the power plant via a two km supply line and to the construction camp via a 5 km supply line.

Supply of water to the construction camp via tanker was investigated and it was deemed that a supply pipeline was the most cost efficient method for transferring water to the construction camp.

18.2.1.3Sub-Area 4130 – Potable Water

Potable water will be produced by a Potable Water Treatment Plant within the processing facility, and will be distributed to the process plant, construction camp, residual operating camp, mining, administration offices and laboratory facilities. For the Base Case there will be nominally 100 m 3 of potable water consumed per day.

18.2.2Area 4200 – Power Supply

18.2.2.1Sub-Area 4210 – Power Generation

Refer to Section 18.8 – Electric Power Plant for the discussion on power supply design.

18.2.2.2Sub-Area 4230 – High Voltage Electrical Distribution

A Main 33kV Switchroom will be the main point of connection for incoming power from the Power Station, as well as fiber optic communications between Telstra and the plant. The Switchroom includes the main plant 33 kV switchgear feeders, metering, and an allowance for process plant power quality equipment.

33 kV power distribution is via the Main 33 kV Switchroom, which will feed 33 kV buried cables supplying the Process Plant 33 kV Substations, as well as the site wide overhead power line.

18.2.2.3Sub-Area 4231 – Power Distribution

It is not desirable to install overhead power lines close to the Process Plant, where it may cause a hazard to over-height vehicular traffic such as cranes. Therefore, in order to keep the overhead power lines away from these areas, 33 kV power to and from the Process Plant will be connected by buried cables. The buried cables will be connected to the Main 33 kV Switchroom.

18.2.2.4Sub-Area 4232 – Overhead Power Lines

33 kV power will be provided from the Power Station via a single feeder. New 33 kV overhead power lines will be required to connect the Power Station to the Process Plant, which are approximately 1.2 km apart. These will be installed along a similar route as the main access road.

The 33 kV power line will also need to be distributed around site to the following facilities:

ANFO Facility

Heap Leach Pad (existing)

Construction Camp/Residual Accommodation Camp

Waste Water Treatment Plant (WWTP)

Pit Dewatering

Mine Services

Site Radio Communication Tower (depending on final location)

Gatehouse

Future Tailings Storage/Decant

The total length of overhead power line required to reach the Process Plant and the above locations from the Power Station is 7.1 km.

The overhead power line will incorporate a fiber optic cable into the earth conductor. Overhead power lines will be suitably rated for a high dust and lightning strike region.

The Accommodation Camp is assumed to be within 2 km of either the existing 22 kV power lines or the new site 33 kV power lines.

18.2.3Area 4300 – Communications

18.2.3.1Sub-Area 4310 – Fiber Optic

Two fiber optic cable ring mains will be installed around the Process Plant to form redundant topology. These cables will generally be installed on cable ladders within the plant, although sections of the cable will be buried where cable ladder access is not available. The second cable is to provide redundancy within the Process Plant in case of damage to the first cable and will follow a separate route where this is practical.

The plant fiber optic cables will contain up to 72 cores and will incorporate separate networks for data communications including those for the Plant Process Control System, the site IT system, a site Voice over Internet Protocol (VoIP) phone system, site Closed Circuit TV (CCTV) and security network, and fire detection system.

Outside of the Process Plant, the fiber optic cables will be incorporated into the earth conductor of the overhead power lines. Optical Ground Wire (OPGW) is a dual functioning cable. It is designed to replace traditional earth wires on overhead power lines with the added benefit of containing optical fiber cores that can be used for communications purposes. These will connect communications equipment from locations such as the Power Station, Water Treatment Plant, Gatehouse and ANFO Facility to the plant communications network.

A Telstra communications hut will be provided (by others) outside the gatehouse. A fiber optic cable will be installed underground between this communications hut and the site overhead power line network at the first overhead power line pole from the Power Station.

As it utilizes the OPGW, the fiber optic cable between the Telstra Hut and Process Plant will not have a second redundant cable, although some redundancy will be provided by using additional fiber cores in the OPGW. Redundancy requirements will be investigated and implemented during the detail design stage.

18.2.3.2Sub-Area 4311 – Phones

Telephone communications will be via digital VoIP technology. This allows telephone calls to be made over an Internet Protocol (IP) network rather than through a separate copper network. Calls can traverse the company’s Information Technology (IT) network or an external portal.

18.2.3.3Sub-Area 4312 – Radios

Refer to Section 18.3.5 – Area 5800 – Communications.

18.2.3.4Sub-Area 4313 – Telemetry

A Radio Telemetry System will be used to communicate to remote locations that require data exchange between the Process Plant and the remote location. Radio Telemetry will be provided to communicate with the decant water return pump station, ANFO Facility and pit dewatering pump station.

The system will incorporate a Master Telemetry Station, located in a switchroom of the Process Plant, and a number of remote Telemetry Stations, located in remote equipment switchboards.

The Master Telemetry Station will communicate with the Plant Process Control System via the preferred communications network, and will communicate with the remote locations via radio. Suitable antennas will be installed at each location.

Control of the remote equipment will be made by the Plant Process Control System, with sufficient data exchange to ensure correct operation of the remote equipment.

18.2.4Area 4400 – Tailings Dam

A total of 202 Mt of process tailings will be stored in two separate tailings storage facilities (TSFs) over a design operating life of 13 years at a nominal ore processing rate of 50,000 tpd. The starter embankments for the existing TSF 1 were constructed during active mining operations between 1996 and 2000. A total of approximately 9 Mt of ore was processed during this period (MWH, 2006). Approximately 87 Mt of additional tailings will be stored in the existing TSF 1 through staged raises of the existing facility constructed using a combination of centerline and upstream construction techniques. TSF 2 will be constructed east of the Process Plant and raised in stages using upstream construction techniques. A total of approximately 114 Mt of tailings will be deposited in TSF 2. The embankments for TSF 1 and TSF 2 will be constructed using non-acid generating waste rock from the open pit operations.

Table 18‑1: 50 ktpd TSF 1 and TSF 2 Parameters

TSF 1 Design Parameter | Value
TSF 1 EXPANSION
Design Tailings Storage Capacity | 87.4 million tonnes
Average Tailings Dry Density | 1.5 t/m 3
Design Life | 12 years
TSF 2
Design Tailings Storage Capacity | 114.7 Mt
Average Tailings Dry Density | 1.5 t/m 3
Design Life | 13 years

The design storage capabilities for TSF 1 and TSF 2 were based on an assumed average in-place dry density of 1.5 t/m 3. Conventional thickened slurry tailings will be pumped to the TSFs at a nominal rate of 50,000 tpd. Tailings will be deposited within the TSF using subaerial deposition techniques through multiple spigot points along the perimeter embankment crest of the TSF.

The existing TSF 1 is a side-hill type conventional slurry tailings storage with perimeter embankments constructed using mine waste and select borrow materials. The existing TSF 1 embankment is referred to as the Stage 1 embankment. The existing facility incorporates an extensive underdrainage system and decant towers with gravity drainage pipes that penetrate the perimeter embankment and connect to an external water collection pond. The existing embankment will be initially raised by the centerline method using mine waste and select borrow material. This approach provides for a robust platform for future raising construction. Subsequent embankment raises will be constructed using mine waste and upstream methods. The TSF 1 raises will be constructed in an alternating sequence with construction of TSF 2 starter and raises. This alternating sequence was adopted to provide adequate time for tailings consolidation and strength gain to permit upstream raising construction. The installation of wick drains in the foundation of each tailings raise

is planned to improve the tailings consolidation rate, reduce risks associated with upstream embankment raising construction, and improve water recovery from the deposited tailings.

The TSF 2 starter embankment will be constructed using mine waste and select borrow material after the TSF 1 Stage 2 raise is completed and operational. The TSF 2 embankment will be raised by upstream methods and using mine waste. TSF 2 raises will be constructed in an alternating sequence with construction of TSF 1 raises. Similar to TSF 1, this alternating sequence was adopted to provide adequate time for tailings consolidation and strength gain to permit upstream raising construction. The installation of wick drains in the foundation of each tailings raise is planned to improve the tailings consolidation rate, reduce risks associated with upstream embankment raising construction, and improve water recovery from the deposited tailings.

18.2.5Area 4500 – Waste Disposal

Sewage waste disposal will be via a Waste Water Treatment Plant (WWTP) installed at the Process Plant Area. The Mine Support and Process Plant buildings will be connected to the WWTP via the sewer pipework reticulation system.

18.2.6Area 4600 – Plant Mobile Equipment

The plant mobile equipment to be purchased for the Base Case process plant will be as follows:

Table 18‑2: Mobile Equipment for Process Plant

Light Vehicles | Quantity
Landcruiser wagon | 2
Dual cab Ute | 11
Tray top Ute | 9
Troop carrier (ambulance) | 1
Bus/troop carrier (15 seater) | 1
Coach | 3
Subtotal | 27
Loader – Cat 966G | Allowed for in mining
Tool Carrier – Cat IT28 | 1
Bob Cat – Mustang Case | 1
Crane – 15t Franna | 1
Hiab Truck – 7t | 1
Service Truck – 2t | 1
2t Forklift – allowance | 2
25t Container Forklift | 1
80t Crane | 1
Subtotal | 9

18.3Facility 5000 – Project Infrastructure

This section provides a description of the Project infrastructure required for the construction and operation of the process plant.

18.3.1Area 5100 – Site Preparation

Bulk earthworks for the Process Plant will be designed to minimize the import of fill material. Where fill material is required to be imported, material from the existing RoM Pad ramp and from the existing stockpile located adjacent to the Tollis and Golf Pits will be utilized.

The site will be prepared such that there is a mono slope fall from the proposed boundary of the pit toward the existing drainage channel on the east side of the proposed process plant. To minimize the extent of stormwater run-off across the plant site, cut-off drainage channels will be installed to divert stormwater run-off around the plant. This will also minimize underground drainage and depth of open channels required on the plant site. A settling pond will be located north of the stockpile and is designed to minimize solids overflowing into the drainage channel.

Stormwater channels will be designed to collect water alongside the unsealed plant roads and direct them beneath the roads via corrugated steel culverts to prevent scouring of plant roads. All stormwater run-off will be directed toward the existing drainage channel on the east side of the proposed process plant. Rip-rap protection to earthwork embankments adjacent to the existing drainage channel on the east side of the proposed process plant will also be installed for flood protection.

18.3.2Area 5200 – Support Buildings

The Support Buildings consist of the building infrastructure for the Process Plant. The support building sizes and number of operations personnel has been developed for the Base Case.

18.3.2.1Sub-Area 5210 – Administration Offices

The Administration Offices will be complexed with multiple transportable buildings and used by plant management and administration personnel and is located at the northern end of the Process Site. The building will include necessary system furniture and provide cellular and open planned offices along with conference and meeting spaces.

The footprint of the Administration Offices is approximately 14.4 m by 29.7 m and will be sized to accommodate 30 people.

18.3.2.2Sub-Area 5211 – Process Plant Offices

The Process Plant Offices will be complexed with multiple transportable buildings located within the existing Flotation building. The buildings will include the necessary system furniture and provide cellular and open planned offices.

The Process Plant Offices will be sized to accommodate 17 people per shift.

18.3.2.3Sub-Area 5220 – Workshop / Warehouse

The Workshop / Warehouse will be incorporated into the existing Flotation Building along with the Process Plant Offices, Main Control Room, Crib and Ablutions and the Light Vehicle Workshop. The Offices / Ablutions / Crib facilities will be transportable building located within the annex of the building.

The existing Flotation building will require modifications to steelwork and replacement of the concrete floors. The building will be complete with services including overhead traveling crane, power, lighting, communications, compressed air, water, specialist equipment and other services necessary for the maintenance of process plant equipment and the LV fleet.

The Workshop / Warehouse will be sized to accommodate 25 people per shift.

18.3.2.4Sub-Area 5230 – Reagent Store

The Reagent Store will consist of one Dome Shelter mounted on sea containers with a concrete floor. The reagent store will be sized approximately 16.7 m by 24.4 m, which includes four sea containers that will act as additional space for the storage of reagents.

The Reagent Store will be complete with all services including power and lighting.

The Dome Shelter will be constructed of steel frame and tensile fabric with a fabric life expectancy of 10 years.

The Reagent Yard will cover an area of 1,750 m 2 and contain a secured fenced hardstand area for the storage of sea containers used at the Reagent Store.

18.3.2.5Sub-Area 5240 – Crib / Ablutions

The Crib / Ablutions facilities will be complexed with transportable buildings located within the existing Flotation Building. The buildings will include the necessary system furniture fixtures and fittings and will be suitable for operations and periodic shutdown personnel.

18.3.2.6Sub-Area 5250 – Emergency Services

The Emergency Services Facilities will be a transportable building used by the First Aid and Fire and Emergency Services personnel. It will be located adjacent to the Administration Offices in the Process Plant and will be sized 14.4 m by 9.9 m. This area will include an undercover area for an ambulance bay and an area for additional services.

18.3.2.7Sub-Area 5255 – Helipad

An allowance has been made for a bitumen helipad to be located close to the Process Plant. The helipad landing zone will be in a fenced-off enclosure and contain a wind sock. The helipad location is not confirmed at this stage.

18.3.2.8Sub-Area 5260 – Sample Preparation and Laboratory

The Sample Preparation and Laboratory facility will be a structural steel shed with insulated metal clad walls and roof and concrete floor for the receipt and storage of samples and a transportable building containing the preparation areas, laboratory and offices for processing samples. The Sample Preparation and Laboratory

building and equipment has been sized to process 450 samples/day. Sampling will be taken from various points throughout the process plant. Samples will be assayed for composition and gold loading.

18.3.2.9Sub-Area 5270 – Gatehouse / Security

The Gatehouse / Security Facilities will be a single transportable building used by security personnel for recording movement to and from the Site and drug and alcohol testing of contractors and employees. The facility will include a boom gate, pedestrian turnstile and swipe card access. The Gatehouse will be located along the access road to the Process Plant.

18.3.2.10Sub-Area 5280 – Control Building – Crushing

The Crushing Control Room will be a single transportable building located at the Primary Crusher. The buildings will include the necessary system furniture for one operator.

18.3.2.11Sub-Area 5281 – Control Building – Main Control Building

The Main Control Room will be a single transportable building located within the Flotation Building. The building will be sized 11 m by 3 m and include the necessary system furniture for one supervisor and three operators.

18.3.2.12Sub-Area 5282 – Control Building – CIP

The CIP Control Room will be a single transportable building located on top of the Leach Tanks and subdivided into a Control Room and a Titration Room. The building will be sized 9.6 m by 3 m and include the necessary system furniture for one supervisor and two operators.

18.3.2.13Area 5300 – Access Roads, Parking and Laydown

The existing Plant Access Road is suitable for the Base Case. Miscellaneous road repairs will be to the existing Plant Access Road.

The existing Process Plant Retention Pond corrugated steel culvert crossing the drainage channel on the east side of the proposed Process Plant is suffering from corrosion. These corrugated steel culverts will be replaced.

18.3.3Area 5400 – Heavy Lift Cranage

Heavy lift cranage covers the cranage that will be needed on site during the construction period for the heavy lifts on site, approximated as follows:

Table 18‑3: Heavy Lift Cranage Requirements

Crane | Duration (Hours Per Year)
600 t | 270
450 t | 470
200 t | 540
180 t | 540
100 t | 810
80 t | 3090
50 t | 1610

18.3.4Area 5600 – Bulk Transport

Bulk transport in and out of site will be weighed on a weighbridge near the gatehouse. The weighbridge will be located on a dedicated off take from the main road. The site weighbridge will be capable of weighing a triple trailer tanker or truck.

18.3.5Area 5800 – Communications

18.3.5.1Sub-Area 5810 – Site-wide Radio Communications

The site will require radio communication for both individual division usage and also across all site personnel for emergencies. Some divisional usage will be localized, but coverage across the site will generally be required.

To cover all radio communications requirements across the site, there will be a suitably located, approximately 50 m tall, communications tower complete with appropriate antenna arrays and ancillary equipment. A communication hut will be located at the base of the tower. This hut will house the repeaters, servers, communications equipment and back-up batteries to provide a robust radio communications system. A maximum of eight individual radio channels will be provided.

Depending on the final location, the communications hut will either be connected to the overhead power line network or, in the case where this is not practicable, a solar powered power supply will be provided. The communication hut back-up battery life will last for a minimum of 10 hours on loss of incoming power.

The radio system will include the following radio quantities for individual personnel and vehicle usage:

320 hand held radios and spare batteries

50 mobile (vehicle) radios complete with battery charger, remote speaker/microphone and antennas

10 base station radios complete with battery charger, remote speaker/microphone and antennas

50 multi-bay chargers for portable radios.

18.4Facility 6000 – Permanent Accommodation

Permanent accommodation for plant operating staff will be in the town of Katherine at the discretion of operators. A portion of the camp will remain after the construction period for temporary accommodation for staff, fly-in maintenance teams and shutdown personnel. Refer to Section 18.5.1 – Area 7300 – Construction Camp for the permanent camp details.

18.4.1Area 6100 – Personnel Transport

A bus transit area consisting of three bus shelters will be constructed in the town of Katherine for transport of operators to and from site. This is to ensure staff will not be driving from the Mt Todd mine site to Katherine after 12-hour shifts.

18.5Facility 7000 – Site Establishment and Early Works

The site establishment will occur prior to the operation of the Construction Camp with the hire / purchase of EPCM Contractor and Client Offices / Crib / Ablutions for the duration of the project. The facilities will be located at the Process Plant Area.

The early works will require a ‘Fly Camp’ for bulk earthworks and services Contractors. This accommodation has been allowed for at the town of Katherine for 40 people for three months to complete the early work at the Construction Camp Facilities and Process Plant Area.

18.5.1Area 7300 – Construction Camp

The Construction Camp will be sized for 390 construction workers for the Base Case based on the manning histogram developed for the Project. The Construction Camp will be located approximately 2 km from the access road. The final location of the camp will be determined during the Feasibility Study.

The Construction Camp will be hired for the 24 month construction duration with the exception of 60 rooms which will be purchased from the outset. Bulk earthworks and all services including power, communications, water and sewerage will be completed prior to the arrival of the hire buildings.

The accommodation village will consist of the following building and services:

390 rooms certified in accordance with the Building Code of Australia

First Aid

Laundry Buildings

Male / Female Ablutions

Dry Mess including Kitchen / Dining / Crib Facilities

Wet Mess

Ice Rooms

Administration Building

Covered Outdoor Area

Gymnasium Building

Power Supply and Distribution

Communications Nodes and Distribution

Potable Water and Reticulation

Fire Services

Organic Materials Waste Dump

Waste Water Treatment Plant

LV Parking Area and Bus Drop Off / Pick Up

Unsealed Access Road

18.6Facility 8000 – Management, Engineering, EPCM Services

Facility 8000 will cover the indirect costs associated with the management of the project from detailed design through to handover to operations. Included within this section will be the EPCM team, external consultants, commissioning team, owner’s team and any costs for licenses, fees, legal costs and insurances.

18.6.1Area 8100 – EPCM Services

This area includes the costs for engaging the services of one or more contractors to perform the engineering, procurement and construction management for the project. The costs in this area have been derived by way of a bottom-up estimate.

18.6.2Area 8200 – External Consultants/Testing

This area is a Prime Cost (PC) Sum that allows for the engagement of any environmental, Human Resources/Industry Relations or Health, Safety, Environment and Community (HSEC) consultants that might be required through the execution phase of the project.

18.6.3Area 8300 – Commissioning

Area 8300 is concerned with the costs for the management and engineering associated with commissioning and was derived, for the Process Plant, as 3% of the total mechanical equipment supply costs.

18.6.4Area 8400 – Owners Engineering/Management

Area 8400 contains costs associated with the owner’s team located either on site and or in the project office.

18.6.5Area 8800 – License, Fees and Legal Costs

This area contains a PC Sum for the costs of licenses, fees and legal costs that would need to be expended throughout the execution phase of the project. Additional costs to this area may need to be incorporated by Vista based on information that is not yet known.

18.6.6Area 8900 – Project Insurances

Project insurances are a PC Sum included to allow Vista to take out any insurances that are deemed necessary to ensure project success. The amount of funds to be included in this area will be dependent on Vista’s criteria for an acceptable risk profile and, as such, is subject to interpretation by Vista.

18.7Facility 9000 – Preproduction Costs

Facility 9000 will cover the indirect costs associated with direct labor during commissioning, the purchase of spare equipment and replacement of equipment damaged during commissioning. The Areas 9600 to 9900 are sums of money associated with working capital, escalation and exchange rate fluctuation, contingency and management reserve.

18.7.1Area 9100 – Preproduction Labor

Preproduction labor covers the costs that are not part of Construction Contracts, not part of Commissioning, not part of post-handover operating ramp up costs but are for costs that may arise prior to operations taking over the Project in an operating context. This area is proposed for minor plant modifications and additions deemed necessary to achieve Project handover status.

18.7.2Area 9200 – Commissioning Expenses

Commissioning expenses is intended to cover the power, materials, labor and spare parts that are associated with making plant modifications, additions and operations during the commissioning period.

18.7.3Area 9300 – Capital Spares

Capital spares are all spares which are non-consumables. These are large items that are not expected to be used; however, these items must be kept in spare for the project due to long lead times, high cost and process importance. These items include, but are not limited to, spare mill motors, HPGR motors, HPGR rolls, intertank screens, conveyor drives, pulleys and belts.

18.7.4Area 9400 – Stores and Inventories

Stores and Inventories allows for a first fill of the primary warehouse for smaller items that are replaced frequently, including but not limited to valves, flanges, pipe fittings, pulleys and ‘V’ belts.

18.7.5Area 9600 – Working Capital and Finance

Working Capital and Finance will be an allowance for a sum of money to be left for use after the plant is operational before the revenue stream is stable. Costs for this have not been included in the estimate as provision for this will be made Vista.

18.7.6Area 9700 – Escalation and Foreign Currency Exchange

Escalation and Foreign Currency Exchange will be an allowance to cover the fluctuation of foreign currencies and inflation from the time of this study until purchase date. Costs for this have not been included in the estimate as provision for this will be made Vista.

18.7.7Area 9800 – Contingency Provision

The contingency provision covers those items within the scope that are known to exist but have not yet be defined.

18.7.8Area 9900 – Management Reserve Provision

The management reserve provision is a measure of the accuracy of this cost estimate and is a portion of additional money that would not be available to the project manager but is held in reserve by Vista to cover unforeseeable and uncontrollable events including, but not limited to: strikes, unusual weather conditions, premium payments arising from accelerated construction programs to recover lost time.

18.8Electric Power Plant

The mine’s electrical power demands are estimated to be 70 MW for the Base Case based upon the load list provided by TTP dated November 27, 2017. Electrical demand will be met through the installation of seven Jenbacher J920 reciprocating gas engines to meet the power demand for the Base Case. Water consumption in this power plant arrangement is very small and primarily for makeup of the closed loop engine cooling system and general housekeeping washdown. It is estimated that intermittent water use will be up to 11.5m 3/hr (50 gpm).

Two potential locations are considered for the location of the 68.2MW power station. Option 1 is locating the power station near the entrance to the mine and includes connection to the existing natural gas pipeline spur with the shortest powerline connection to the mine. This pipeline spur requires a tolling fee of AUD0.60 per gigajoule (GJ) of gas on top of the wholesale price of AUD7.00 per GJ.

Near the main gas transmission pipeline is an alternate Option 2 location for the power station that will avoid the tolling fee for the spur pipeline but requires the upgrade of the 10 km electrical powerlines with new conductors and as well as new towers designed to carry the additional weight of conductors rated for a higher megavolt-ampere (MVA) class. The Option 2 location will also require additional infrastructure for non-potable service water and fire protection, which would entail a non-potable service/fire water storage tank, firewater pumps, and non-potable service water pumps.

Figure 18‑1: Power Station Location

The plant required net output is the mine’s required electrical demand plus the power stations auxiliary loads and is based on the preferred generation technology.

Table 18‑4: Power Station Location Budgetary Comparison

Mine Output | Base Case
Generator Manufacturer | GE-Jenbacher
Model | J920 (7)
Plant Required Net Output | 69.6
Estimated Fuel Demand (GJ/yr) | 5,253,349

For the purpose of this Preliminary Feasibility Study, the Option 1 location will be carried forward.

18.8.1 Generation Option Selection

POWER Engineers (POWER) performed an evaluation of generation options based upon providing reliable power for a steady load demand with minimum onsite personnel requirements and low life cycle operating costs. The evaluation has concluded that natural gas-fueled reciprocating engines provide the most economical and reliable means of power generation at the mine site. POWER recommends that the Base Case electrical demand be met with seven reciprocating engines with backup power being provided by the electric utility grid. Gas reciprocating engines will provide excellent reliability, increased operational flexibility and redundancy required by the Project. If startup loads exceed the capacity of the initial engines, the utility grid connection is available for supplemental power.

The technology configuration for this study uses the General Electric Jenbacher J920 engine but there are a number of commercially available gas engine models in the 10MW range with operating plants in-country to provide a technically sound field to select a reliable equipment supplier with competitive pricing and local technical support.

18.8.2Mt Todd Electrical

18.8.2.1Conceptual Design

A conceptual electrical one-line diagram, Figure 18‑2, has been created to show the electrical distribution system from the 33kV utility interconnect down to the 400V power distribution bus. The equipment ratings are preliminary and based on generator ratings provided by General Electric and budgetary quotes for the balance of plant equipment.

The equipment ratings shown are for cost estimating purposes only. The actual equipment ratings will be determined using detailed load flow and short circuit studies during detailed design.

18.8.2.2Plant Arrangement

The auxiliary electrical equipment is included on the mechanical general arrangement drawing (Figure 18‑3). All the equipment physical sizes are based upon similar equipment from reference projects.

18.8.2.3Step Load Capability

Reciprocating engine generators are designed for fast startup times, fast ramp rates and flexibility. For best performance when starting and fast ramp times, the engine must be pre-heated or operating at low loads with the cooling water at 55°C (by onboard electric heater) to avoid engine damage. A single 10 megawatt (MW) engine can accelerate from start to 100% load in 5 minutes with ramp up as fast as 100 kilowatts per second (kW/sec). If a large motor at the mine is brought online, the starting load can be spread across all engines in the power station for a ramp rate as high as 700 kW/sec for the Base Case with the connection to the electric utility grid available for supplementary starting support.

Figure 18‑2: Conceptual Electrical Line Diagram

Figure 18‑3: General Plant Arrangement

19.0MARKET STUDIES AND CONTRACTS

19.1Markets

Gold metal markets are mature, with many reputable refiners and brokers located throughout the world. The advantage of gold, like other precious metals, is that virtually all production can be sold in the market. As such, market studies, and entry strategies are not required.

Metallurgical process studies confirm that the Project will produce doré of a specification comparable with existing operating mines.

Demand is presently high with prices showing remarkable increases during recent times. The 36-month average London PM gold price fix through December 31, 2017 is US$1,222/oz.

19.2Contracts

Currently there are no contracts in place. However, Vista has obtained budgetary quotes, as is common for PFS level studies, for future service needs. The following contracts will be in place upon project commencement:

Secure doré transportation to market;

Doré refining; and

Supplier and service contracts including:

Diesel and fuel oil;

Natural gas for the power plant;

Process reagents;

Equipment preventive maintenance and repair (MARC) services;

Site security services; and

Camp management, catering and support services.

20.0ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT

The Base Case is identified as a 50,000 tpd operation, as presented in this section. Another option, defined as the Alternate Case (33,000 tpd) is presented in Section 24.7 – Alternate Case.

This section discusses the environmental permitting and social impact aspects of the Project. The EIS was submitted in June 2013. The Northern Territory Environmental Protection Authority (NTEPA), as the responsible government authority to advise on the environmental impact of development proposals, provided its final assessment of the Project in June 2014.

In January 2018, the “authorization of a controlled activity” was received for the Project as required under the Australian Environmental Protection and Biodiversity Conservation Act of 1999 (EBPC) as it relates to the Gouldian Finch, and as such has received approval from the Australian Commonwealth Department of Environment and Energy.

20.1Environmental Studies

A number of environmental studies have been conducted at the Project in support of development of EISs and as required for environmental and operational permits. Studies conducted have investigated soils, climate and meteorology, geology, geochemistry, biological resources, cultural and anthropological sites, socio-economics, hydrogeology, and water quality.

The Mt Todd Project Environmental Impact Statement (EIS) submitted June 28, 2013 to the Northern Territory Environment Protection Authority (NTEPA), approved in September 2014, provides an understanding of the existing environmental conditions and an assessment of the environmental impact of the Project.

Key issues of concern regarding the Project impacts that were addressed in the EIS include:

Acid and metalliferous drainage (AMD) seepage and runoff from the waste rock dump, ore stockpiles and tailings storage facilities potentially contaminating surface and ground waters continuing long after the mine has ceased operation;

Potential contamination of surface water from AMD causing adverse impacts on downstream water quality, aquatic environment and downstream users;

Potential contamination of groundwater from AMD causing groundwater quality impacts outside of the Mineral Lease or release of contaminated groundwater expressing to surface water;

Potential groundwater drawdown impacts to groundwater flows in the Edith River and the potential to impact production bores in the region;

Management and treatment of a large quantity of acidic and metal laden water currently existing on the site;

The proposed WRD covers an approximate area of 217ha with an estimated height of 350m. Final design of the WRD must ensure the structure is safe, stable, not prone to significant erosion, minimizes AMD seepage and runoff and meets stakeholder expectations as a final land use structure;

Biodiversity impacts, including matters of environmental significance, associated with disturbance footprint of mining activities and infrastructure requirements;

The challenges of successful mine closure and rehabilitation; and

Potential social, economic, transport and heritage impacts.

The Project is located in the Pine Creek Bioregion and part of the Yinberrie Hills Site of Conservation Significance (SOCS). Each of these potential impacts were assessed and mitigation or management measures were outlined in the EIS.

20.1.1Flora and Vegetation

Eight vegetation types covering 5,462.56ha were mapped in the Mineral Leases. Eucalyptus tectifica, E. latifolia, E. tintinnans, E. spp. Woodland; E. phoenicea, Corymbia latifolia low woodland – woodland (scattered E. tintinnans); and C. dichromophloia, E. tintinnans, Erythrophleum chlorostachys Woodland covers 80% of the site. The Project is not expected to significantly impact vegetation in the area.

Eight-hundred and forty species of flora are known to occur within 10km of the leases. Recent surveys identified 226 taxa, of which 67 were not recorded from previous surveys. The total number of species known from the area is now 959. The only threatened plant species recorded from the area is the bladderwort, Utricularia singeriana. This species as vulnerable to extinction. The closest known record is 6km west of the Mineral Leases. The Project is not expected to have an impact on any threatened flora.

20.1.2Nationally Threatened Species

Ten EPBC Act listed threatened species potentially occur within 10km of the project area. Two of these, the Gouldian finch and crested shrike-tit, were recorded on-site. The EPBC considered potential mitigation for the Gouldian finch. The partridge pigeon was not recorded, but may be present on-site occasionally. The remaining seven species are unlikely to occur on-site.

20.1.3Migratory and / or Marine Species

Fourteen EPBC Act listed migratory bird species potentially occur within 10km of the project area. Ten have been recorded from the leases. Seven EPBC listed marine species potentially occur with 10km of the project area. This includes six bird species and one reptile species. The freshwater crocodile was recorded in the leases. None of the listed marine species is likely to have a high risk of impact from the proposed development.

20.1.4National Heritage Places

The Yinberrie Hills is a Site of Conservation Significance and was placed on the Interim Register of the National Estate for its natural values. However in 2007 the Register of the National Estate was declared no longer a statutory list.

Surveys located 20 archaeological sites. The most significant was Mt Todd 26 – an extensive greywacke quarry, extraction and reduction site, one of the largest recorded in the Northern Territory. The remainder were lithic scatters or quarry and reduction sites with low to medium heritage significance.

With respect to Jawoyn Resource Knowledge, 62 animal, 63 plant and one fungal taxa were identified and the associated Jawoyn knowledge recorded. Amongst the Jawoyn, the mine site is not considered a notably productive environment. Plants and animals encountered and discussed during the ecological knowledge consultation are widespread and not unique to the mine site. Vista employs Jawoyn Rangers for reviewing and potentially clearing any heritage sites prior to disturbance.

20.2Waste and Tailings Disposal, Site Monitoring and Water Management

20.2.1Waste Rock Disposal

Waste rock will be disposed of in a WRD constructed as an expansion of the existing WRD. All waste rock will be analyzed to identify the rock as potentially acid generating (PAG) or non-PAG material before being hauled to the WRD. Non-PAG material will be stockpiled for use in reclamation covers or placed in the WRD. Construction of the WRD is described in Section 16.0 – Mining Methods. Reclamation and closure of the WRD is described in Section 20.5 – Mine Reclamation and Closure.

20.2.2Tailings Disposal

Tailings will be disposed of in two tailings storage facilities, TSF 1 and TSF 2. TSF 1, an existing tailings storage facility, will be expanded with eight additional raises to the embankment and construction of two new saddle dams at the west end of the impoundment. A second tailings storage facility, TSF 2, is to be constructed after re-commissioning of TSF 1. The engineered containment system for the TSF 2 impoundment includes a 60‑mil linear low-density polyethylene (LLDPE) textured (double sided) liner and a tailings overdrainage collection network to mitigate the risk of seepage. Tailings decant water and water collected in the TSF seepage interception network will be treated in the water treatment plant or used for the process plant. Construction of the tailings storage facilities is described in Section 18.2 – Facility 4000 – Project Services.

Reclamation and closure of the TSFs is described in Section 20.5 – Mine Reclamation and Closure.

20.2.3Site Monitoring

Currently, surface water monitoring is conducted at various locations at the site. A comprehensive site monitoring plan will be developed as part of future work on the Project.

20.2.4Water Management

The primary existing environmental issue at the site is water management resulting from the project shutdown without implementation of closure or reclamation activities. The pit and existing water RPs (excluding the raw water pond) contain acidic water with elevated concentrations of regulated constituents. This water has been managed through evaporation, pumping to the Batman Pit for containment, micronized lime treatment of the pit lake, and controlled discharge of treated water to the Edith River in accordance with the approved WDL. Historically, wet season rainfall resulted in short-term uncontrolled overflow from retention ponds to the Edith River due to the high amount of precipitation received in short periods of time coupled with insufficient pumping capabilities. Current water management strategies employed by Vista appear to be successful at preventing recurrence of historic uncontrolled discharges and are minimizing impacts on the Edith River downstream of the Project Site.

Prior to, during, and following resumed mining operations, water management at the site involves distinct water management components including in-pit treatment, seepage management, treatment of acid rock drainage and metal laden leachates (ARD/ML), and surface water management. Each of these components is discussed in the subsections below:

20.2.4.1In-situ Pit Treatment

In-situ treatment of the Batman Pit (RP3) was conducted by use of limestone and quicklime. Treatment has been undertaken to produce water to be discharge at rates protective of water quality in the Edith River in a suitable timeframe to meet project requirements. The treatment methodology included raising the pH of the water within the pit lake to greater than pH 7.0 using limestone and quicklime in succession to capitalize on the capabilities of the low-cost limestone and minimize the quantity of quicklime required to attain a pH sufficient to precipitate additional metals. Raising the pH to greater than 7.0 will result in the precipitation of key metals of concern including iron, aluminum, chromium, copper, lead, nickel, cadmium, cobalt and zinc. On an ongoing basis, quicklime is used to buffer the pH as required on an annual basis.

20.2.4.2Seepage Management

A thorough assessment of the infiltration and seepage conditions of the WRD, HLP, TSF 1, ore stockpiles, and other site facilities has not been well characterized at the current time but will be foundational to developing the site water management plan. The infiltration and seepage assessment will be included in the comprehensive site environmental system model (hydrogeologic, geologic, seepage, and geochemical conceptual models) to understand the solute-transport processes at the site and possible impacts to the aquifer from mine operation. Numeric modeling will be used for the infiltration and seepage assessment.

20.2.4.3Ongoing ARD/ML Water Treatment

Water treatment for the project will involve active water treatment for ARD/ML. Active water treatment will occur prior to operations, as part of rehabilitation of the site necessary to restart mining, during mining operations, and for a period following cessation of operations. Passive water treatment will be conducted at the site following closure in addition to use of the active water treatment plant as required.

Active water treatment at the site has been described in Section 24.0 – Other Relevant Data and Information.

Passive water treatment will be conducted in four separate passive treatment systems which include (in total) one biochemical reactor (BCR), four aerobic polishing wetlands (APW) and three aeration/settling ponds (AP). The goals of the passive/semi-passive water treatment at Mt Todd are to:

Eliminate or drastically curtail the costs and continual inputs (e.g. reagents, power, staff) required to operate and maintain the active WTP;

Eliminate sludge disposal operations and maintenance associated with active water treatment;

Collect, contain, and treat ARD/ML prior to effluent release year-round; and

Ensure that treated ARD/ML complies with the WDL numeric water quality standards.

The passive water treatment technology recommended for treatment of WRD seepage, which is predicted to be net-acidic ARD/ML, is primarily metal-sulfide and metal-hydroxide precipitation via sulfate-reduction and the concomitant rise in solution alkalinity. The passive water treatment technology recommended for treatment of seepage from the TSFs, which is predicted to be net-alkaline ML, is aeration (oxidation) in

aeration/settling ponds (Aps) to allow metals to precipitate and settle. Effluent from the APs will be further aerated and treated prior to release to the environment in aerobic polishing wetlands (APWs) where the concentration of dissolved metals should be further reduced through complexation to plant-derived organic substrate, and potentially, accumulation in plant tissue.

The treatment capacity of the four separate passive water treatment systems range from 10 to 50 m 3/hour, which should be adequate to treat the anticipated rate of seepage from the WRD and TSFs following closure. The quantity of seepage from the WRD and TSFs following closure was estimated by simply multiplying the predicted infiltration of daily precipitation through the proposed WRD and TSF closure covers by the ultimate two-dimensional surface area of each facility. Using stochastic precipitation developed in the water balance model from site and Katherine gage data statistics, 1000 simulations (realizations) of daily precipitation were calculated in GoldSim at the following probabilities: 0%, 1%, 5%, 15%, 25%, 35%, 45%, 50%, 55%, 65%, 75%, 85%, 95%, 99%, and 100%. The mean of these precipitation probabilities was then calculated to represent daily precipitation. To estimate the daily seepage rate from each facility the calculated mean daily precipitation was multiplied by the ultimate facility surface area and the estimated rate of infiltration through the closure cover.

Estimating flows and water quality 20 years in the future is wrought with uncertainty. These and other uncertainties inherent to passive water treatment are magnified by changes in mine plans and changes in closure plans and designs, which occur during normal operations, as well as unpredictable circumstances such as changes in climatic conditions, unforeseen material characteristics, etc. Therefore, the estimates and recommendations provided at this time should be considered preliminary and design parameters such as: hydraulic retention time; biochemical oxygen demand removal rate; metals and metal-precipitates removal and settling rate; and reactive substrate type, quantities, depletion rate and permeability overtime must be checked and updated or entirely modified as the project progresses and more information becomes available.

20.2.4.4Surface Water Management

Surface water at the site is well-documented and its management has been the object of study by both Vista and the NT Government in recent years. Surface water management is described further in Section 24.4 – Surface Water Hydrology.

20.3Permitting and Authorizations

On January 1, 2007, Vista became the operator of the Project Site and accepted the obligation to operate, care for, and maintain the assets of the NT Government on the site. Vista developed an Environmental Management Plan (EMP) for the care and maintenance of the Mt Todd mine site in accordance with the provisions of the Mineral Leases 1070, 1071, and 1127 granted under the Mining Act (Vista, 2007a). The EMP identified the environmental risks found at the Project Site at its then present state of operations and defined the actions for Vista to take to control, minimize, mitigate, and/or prevent environmental impacts originating at the Project Site. As part of the agreement, the NT Government acknowledged its commitment to rehabilitate the site and that Vista has no obligations for pre-existing conditions until it submits and receives approval of an MMP for resumption of mining operations.

The Project requires approvals, permits and licenses for various components of the Project. Table 20‑1 includes a list of approvals, permits, and licenses required for the project and their current status.

Table 20‑1: Mt Todd Permit Status

Approval/ Permit/ License | Current Status | Approval/ Permit License Date | Expiration Date
Environmental Impact Statement | The NT Environmental Protection Authority provided its final assessment of the Project in June 2014. | Approved Sep. 2014 | NA
Mining Management Act (or Plan) Approval from NT Department of Mines & Energy | Mine operating permit request has not been submitted. Interim MMP submitted in November 2017 regarding current operations. | Prior to commencing mine operations | NA
Heritage Act permit to destroy or damage archeological sites and scatters/ Aboriginal Areas Protection Authority Clearances | Authority Certificate Number 2011/15538 issued. This certificate defined restricted works areas and granted select clearances to allow for initial investigations. Additional clearances will be required for further investigations as well as prior to disturbance associated with mine development. | Aboriginal Areas Protection Authority dated Jul. 31, 2012 | NA
Dangerous Goods Act (1988) permit for blasting activities | Waiting on final mine plan | NA | NA
Extractive Permit (under DME Guidelines) for development of borrow pits outside of approved mining areas | Would be required for PGM or LPM borrow areas. Permit application not yet in progress pending final selection of borrow areas | NA | NA
Waste Discharge License (under Section 74 of the Water Act 1992) for management of water discharge from the site | WDL 178-5 licensing discharge of waste water into the Edith River from the Mt Todd mine site, granted with conditions | Jan. 16, 2017 | Aug. 31, 2018
Waste water treatment system permits under Public Health Act 1987 and Regulations | May be required for the waste water treatment system for the construction and operations accommodation village. Permit application not yet in progress pending design and siting of accommodation village. | NA | NA
Approval to Disturb Site of Conservation Significance (SOCS) | Batman pit expansion will disturb SOCS as breeding / foraging habitat for the Gouldian finch, pending determination on EIS. | Jan. 22, 2018 | NA

In addition, permits that are required to commence construction works will be obtained prior to any construction activity.

20.4Social or Community Requirements

The Jawoyn people have strong involvement in the planning for the future of the Project. Vista has a good relationship with the Jawoyn. Areas of aboriginal significance have been designated, and the mine plan has avoided development in these restricted works areas.

Those parts of the JAAC agreement that are within the public domain are presented in this report; the remaining part of the agreement, which is confidential, is not presented in this Technical Report.

20.5Mine Reclamation and Closure

A reclamation plan for the Project was developed in support of the PFS for renewed mining operations. This reclamation plan evaluates the reclamation activities that will be conducted for the landforms planned as part of mining commencement. Reclamation plans and strategies for each major facility at Mt Todd are briefly summarized in Table 20‑2. Costs associated with reclamation and closure are provided in Sections 21.0 and 22.0.

Table 20‑2: Reclamation Approach

Task | Facility
Batman Pit | WRD | HLP | TSF 1&2 Impounded Surface | TSF 1&2 Dams (Embankments) | Process Plant and Pad | LGOS 2
Surface of Facility at Cessation of Production Composed of Non-PAG Material |  | X |  |  | X |  |  |
Final Overall Slopes > 3H:1V* | X | X |  |  |  |  |  |
Final Overall Slopes < 3H:1V* |  |  | X | X | X | X | X | X
Benches Created During Construction | X | X |  |  | X |  |  |
Install minimum 1.0 m-Thick non-PAG Material |  | X |  | X |  |  |  |
Install 0.8 m-Thick Store and Release Cover |  |  |  | X | X |  |  | X
Install 0.2 m-Thick Plant Growth Medium (PGM) Cover |  |  | X | X | X | X | X | X
Revegetate with Native Seed Mix |  |  | X | X | X | X | X | X
Install geosynthetic liner (with under and overlayer of fines) |  | X |  |  |  |  |  |
Install Erosion and Sediment Controls |  | X |  | X | X | X | X | X
Construct Access Restriction Bund | X |  |  |  |  |  |  |
Additional Remedial Measures (as necessary) | X | X | X | X | X | X | X | X

* > and < indicates slopes are steeper and less steep, respectively.

“X” denotes where the task or characteristic is applicable to the landform

Costs associated with reclamation and closure are provided in Sections 21.0 and 22.0. In accordance with regulatory requirements, a reclamation bond will be required for the site. Calculation of bond amounts will be conducted with the NT Security Calculation excel-based worksheet periodically throughout the mine life in accordance with regulatory requirements. Costs associated with reclamation bonding have been included in the technical economic model.

20.5.1Batman Pit

Based on a preliminary regional groundwater flow model that included enlargement of the Batman pit and post-mining recovery of the groundwater system (outlined in Section 24.5 – Regional Groundwater Model and Mine Dewatering), a terminal-sink pit lake may result during the post-closure phase, making active dewatering and treatment of pit water unnecessary following closure. All water inflow to the pit lake, including precipitation, storm-water runoff and groundwater, will leave the pit lake only via evaporation. No surface water or groundwater drainage from the pit lake is expected to occur.

A berm (also termed bund) will be constructed around the perimeter of the Batman pit to impede human access and reduce the inflow of surface water to the pit. The safety berm will be offset 30 m from the pit perimeter per the requirements outlined in the guidelines “Safety Bund Walls around Abandoned Open Pit Mines” from the Department of Industry and Resources in Western Australia.

20.5.2Waste Rock Dump

The existing WRD will be slightly enlarged based on plans for the resumption of mining. The WRD will be constructed at an angle of repose slope of 1.5 vertical to 1.0 horizontal, with catch benches of 8.0 meters every 30 meters in height. Each lift will be constructed with 8 m wide benches at 30 m vertical intervals on the face of the WRD.

As described in Section 16.0 – Mining Methods, the WRD will be constructed with an encapsulating non-PAG material outer shell on each lift. Concurrent installation of a low permeability geosynthetic liner (i.e., LLDPE or GCL) following attainment of final grades will serve to reduce infiltration of precipitation into the WRD core. This liner system will include a 0.3 m thick bedding layer of fine material to serve as liner bedding, followed by placement of the liner material, and capped with a 0.3-m thick protecting layer of fine material placed over the liner. The liner will span approximately 52 m on top of each lift, covering the 8 m bench, and running below the subsequent lift. The liner will be installed at five percent slopes toward the outside of the WRD, and will be constructed with a 0.5-m tall berm with 1:1 side slopes at the interior edge of the liner. At least a 1-m thick layer of non-PAG waste rock will cover all surfaces of the WRD to aid in erosion control.

Prior to WRD grading, a seepage collection system will be constructed along the down-gradient toe of the WRD and subsequently covered with waste rock from grading activities. ARD/ML collected by the WRD seepage collection system will initially be pumped to the New WTP for treatment prior to release until it is feasible to treat this and other ARD/ML on-site using a passive treatment system.

20.5.3Tailings Disposal Facility

The TSF embankment and impoundment surfaces will be reclaimed at closure by installing and revegetating a 1-m thick store and release cover. The 1-m thick store and release cover will consist of a 0.8-m thick layer of blended non-PAG waste rock (40%) and low-permeability material (60%), overlain by a 0.2-m thick layer of plant growth medium (PGM). Following placement, the cover surface will be roughened and revegetated with native species. The store and release cover will serve to effectively reduce percolation of precipitation into waste rock, PAG, and/or metaliferous materials.

The majority of the impounded surface of the TSF at closure will be primarily composed of thixotropic tailings (thick like a solid but flows like a liquid when a sideways force is applied) which will maintain a

high degree of saturation for many years unless actively dewatered and consolidated, covered with material, or chemically treated to increase their strength. A crowned cover constructed using non-PAG and PAG waste rock and sorter reject material will result in a final tailing surface that drains and does not impound water. This crowned cover is assumed to adequately bridge the thixotropic tailings and allow for equipment to place the 1-m thick store and release cover.

To the degree possible, store and release covers will be installed concurrently during construction when portions of facilities reach final grade. Storm water drainage, erosion, and sediment controls will be constructed to minimize erosion and scour of active reclamation areas.

20.5.4Processing Plant and Pad Area

A new process plant will be built for renewed mining. Once ore processing ceases, the process plant will be decommissioned, decontaminated, demolished and any reusable equipment and materials will be salvaged and resold. Material that cannot be treated in-situ will be excavated and disposed of in the WRD, TSF, or an off-site facility that is certified to accept and dispose of contaminated soil. Concrete foundations, building walls, and other inert demolition waste will be broken up and either:

Placed in the WRD;

Buried in-place; and/or

Backfilled against cut banks and highwalls throughout the process plant and pad area, as well as other areas that will be reclaimed at Mt Todd.

Surface and large shallow pipes will be removed and pipes at depth will be plugged with concrete or other suitable materials.

The process plant area will be graded to blend into the surrounding topography and drain towards Batman Creek. The process plant area and pad will be covered with a 0.2-m thick layer of plant growth medium (PGM) and revegetated. Storm water drainage, erosion, and sediment controls will be constructed to minimize erosion.

The WTP and equalization pond (EQP) will be left in place, up-graded if necessary, and used to treat acid rock drainage and metal-laden leachates (ARD/ML) during the closure and post-closure phases. These facilities will be closed when it is feasible to treat ARD/ML in passive treatment systems.

20.5.5Heap Leach Pad and Pond

The HLP and Pond will be left in place and reprocessed following processing of ore and low grade ore. Following reprocessing of the heap material, the pad and pond footprint will be reclaimed by cutting and removing the liner for consolidation in TSF 2. It is anticipated that the integrity of the heap liner will have been compromised and removal of 0.5-m thick of impacted soils below the liner will be necessary. These materials would be removed and consolidated in TSF 2. The area will then be regraded to prevent ponding of water, and will be covered with a 0.2-m thick layer of PGM and revegetated.

20.5.6Low Grade Ore Stockpile

The existing LGOS1 will be eliminated during the expansion of the Batman Pit and it is assumed that no reclamation is required for the closure of this facility.

The LGOS2 will be located near the pit and the process plant area. Closure of LGOS2 will include removal of residual ore from the stockpile areas, regrading, covering the material with a 0.2-m thick layer of PGM and revegetating the area. In addition, storm-water drainage, erosion, and sediment controls will be constructed to minimize erosion. It is assumed that RP2 will be closed during the closure phase and that the LGOS will no longer be a source of ARD/ML following closure.

Any potential ARD generated during operations reports to the process water pond, and therefore the WTP.

20.5.7Mine Roads

Mine access roads will remain in place to provide post-closure access to the area. All haul roads will be closed by grading into surrounding topography, ripping subgrade materials, placing 0.2 m of PGM (when applicable), and revegetating the areas.

20.5.8Water Storage Ponds

Prior to construction of the active WTP, a process water pond (PWP) will be constructed for mixing of ARD/ML from various on-site sources prior to treatment and to temporarily store ARD/ML in case of system upset. All proposed and existing ponds at Mt Todd will be maintained for the collection of seepage, storm water and ARD/ML until long-term quality of water collected by the WRD seepage collection system meets applicable standards, flows to the collection system cease, or an alternative passive water treatment system is installed.

The return water, polishing and overdrain ponds for the TSFs shall remain post-closure and be incorporated into the passive water treatment system. These and potentially other ponds maybe be used post-closure as backup water storage in case treatment upset occurs.

To decommission and close ponds, residual standing water will be pumped to the PWP for processing by the WTP, and sediments and foundation materials will be tested to determine their chemical characteristics with acidic, PAG and metaliferous materials treated in-situ or buried in place. Following sediment testing and removal, pond liners will be cut and folded in place. Pond berms will be pushed into the pond void to cover the liners and until the area no longer impounds water. The top 0.6 m of graded material is assumed to have physical and chemical properties to support plant growth. Storm water drainage, erosion, and sediment controls will be constructed to minimize erosion and channel scour, and the areas will be revegetated.

20.5.9Low Permeability Borrow Area

A low permeability borrow area will be developed to provide low permeability material for use in project feature construction and for use in reclamation. As portions of the low permeability borrow area are taken out of service and are no longer used to generate material, they will be reclaimed by ripping and amending the remaining soils with organic matter, constructing channels to route drainage within the borrow area footprint and revegetating the area. Some portions of the low permeability borrow area may also be used as stock water ponds.

20.5.10Closure Cost Estimate

Costs for reclaiming major facilities at the Project were estimated using closure material quantities based on the Base Case ultimate designs and following the closure plans discussed above. Closure costs are accrued and contained in the financial model.

21.0CAPITAL AND OPERATING COSTS

The Base Case is identified as a 50,000 tpd operation, as presented in this section. Another option, defined as the Alternate Case (33,000 tpd) is presented in Section 24.7 – Alternate Case.

For the purposes of understanding how the mine will operate, Table 21‑2 details the Project based on the principal operating time periods.

Table 21‑1: Operating Periods

Principal Assumptions | Unit | Parameter
Pre-Production Period | Years | 2
Mine Life | Years | 13
Closure Period | Years | 4
Operating Days | Days / Year | 355

Estimated capital and operating costs are summarized in this section and are prepared by Vista’s engineers and consultants as follows:

Open Pit Mining: MDA;

Process Plant: TTP;

Tailings Dam: Tetra Tech;

Infrastructure: TTP;

Raw Water Dam & Water Treatment: Tetra Tech;

Reclamation: Tetra Tech; and

Owner’s Costs: Vista.

Costs are presented in Q4 2017 US dollars and are based on an US$0.80:AUD1.00 exchange rate, unless otherwise noted.

Section 21.0 presents costs as provided to Tetra Tech for incorporation into the Technical Economic Model (TEM). These costs are based on their source data and in some cases use different foreign exchange rates or unit rates for fuels, etc. The cash flow results presented in Section 22 are all tied to the same foreign exchange and unit costs rates. These costs are summarized using the listed foreign exchange rate provided in Section 22.0 – Economic Analysis.

21.1Capital Cost

LoM capital cost requirements are estimated at US$1,191 million as summarized in Table 21‑2. Initial capital of US$839 million is estimated to be required to commence operations. At the end of operations, the Project will receive a US$142 million credit for remaining asset sales and salvage (reference Table 22‑13).

Table 21‑2: Estimated Capital Cost Summary (US$000s)

Area | Description | Cont. (%) | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
Estimate | Contingency | Total | Estimate | Contingency | Total | Estimate | Contingency | Total
2000 | Mining | 0.3% | $144,278 | $1,204 | $145,482 | $385,207 | $237 | $385,445 | $529,485 | $1,441 | $530,927
3000 | Process Plant | 20.7% | $340,768 | $72,386 | $413,154 | $22,992 | $2,951 | $25,943 | $363,760 | $75,337 | $439,097
4000 | Project Services | 3.5% | $117,676 | $7,085 | $124,760 | $82,857 | $0 | $82,857 | $200,533 | $7,085 | $207,617
5000 | Project Infrastructure | 12.9% | $23,263 | $3,001 | $26,264 | $0 | $0 | $0 | $23,263 | $3,001 | $26,264
6000 | Permanent Accommodation | 10.0% | $68 | $7 | $75 | $0 | $0 | $0 | $68 | $7 | $75
7000 | Site Establishment & Early Works | 11.4% | $19,723 | $2,248 | $21,972 | $0 | $0 | $0 | $19,723 | $2,248 | $21,972
8000 | Management, Engineering, EPCM Svcs | 9.5% | $85,941 | $8,139 | $94,080 | $0 | $0 | $0 | $85,941 | $8,139 | $94,080
9000 | Pre-Production Costs | 17.2% | $11,127 | $1,918 | $13,045 | $0 | $0 | $0 | $11,127 | $1,918 | $13,045
10000 | Asset Sale | 0.0% | $0 | $0 | $0 | ($141,788) | $0 | ($141,788) | ($141,788) | $0 | ($141,788)
| Capital Cost | 9.1% | $742,844 | $95,988 | $838,832 | $349,269 | $3,189 | $352,458 | $1,092,113 | $99,177 | $1,191,290

21.1.1Mining (MDA)

Table 21‑3 shows the estimated mine capital requirements for the Base Case by year. The initial mine capital is estimated to be US$132 million, with a LoM capital of US$372 million. This includes capitalized operating costs of US$63 million for construction, US$44 million for pre-stripping, and US$26 million for reclamation.

Table 21‑3: Estimated Mine Annual Capital Costs (US$000s) – Base Case

| Pre-Prod | Yr 1 | Yr 2 | Yr 3 | Yr 4 | Yr 5 | Yr 6 | Yr 7 | Yr 8 | Yr 9 | Yr 10 | Yr 11 | Yr 12 | Yr 13 | Yr 14 | Total
PRIMARY MINING EQUIPMENT |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Atlas Copco PV235 | $ 17,121 | $ 6,420 | $ - | $ 8,560 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 32,102
165mm Rotary Blast Hole Drill | $ 714 | $ - | $ - | $ - | $ - | $ - | $ 714 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 1,428
28m3 Hyd. Shovel (PC 5500) | $ 13,753 | $ 6,877 | $ - | $ 6,877 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 27,507
18m3 Front End Loader (994) | $ 4,154 | $ 4,154 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 8,309
250t Haul Truck | $ 37,701 | $ 49,011 | $ - | $ 56,551 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 143,263
Total Primary Equipment | $ 73,443 | $ 66,462 | $ - | $ 71,988 | $ - | $ - | $ 714 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 212,607
SUPPORT EQUIPMENT |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
300 Kw Dozer (D9) | $ 1,716 | $ - | $ - | $ - | $ - | $ - | $ 1,716 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 3,433
230 Kw Dozer (D8) | $ 531 | $ - | $ - | $ - | $ - | $ - | $ 531 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 1,062
4.9 m Motor Grader (16H) | $ 1,788 | $ - | $ - | $ - | $ - | $ - | $ 1,788 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 3,575
Water Truck - 70,000 Liter | $ 3,008 | $ - | $ - | $ - | $ - | $ - | $ 1,504 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 4,512
RTD Dozer (834H) | $ 1,036 | $ 1,036 | $ - | $ - | $ - | $ - | $ 2,073 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 4,145
Rock Breaker - Impact Hammer (691 Kg m) | $ 43 | $ - | $ - | $ - | $ - | $ - | $ 43 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 86
Backhoe/Loader (1.5 cu m-446D) | $ 252 | $ - | $ - | $ - | $ - | $ - | $ 252 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 503
2 cm excavator (Cat 392) | $ 315 | $ - | $ - | $ - | $ - | $ - | $ 315 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 631
Low Boy | $ 994 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 994
Flatbed | $ 56 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 56
Manlift | $ - | $ - | $ - | $ - | $ 21 | $ 21 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 42
Total Support Equipment | $ 9,738 | $ 1,036 | $ - | $ - | $ 21 | $ 21 | $ 8,221 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 19,038
BLASTING |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Skid Loader | $ 52 | $ - | $ - | $ - | $ 52 | $ - | $ - | $ 52 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 157
Total Blasting Equipment | $ 52 | $ - | $ - | $ - | $ 52 | $ - | $ - | $ 52 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 157
MINE MAINTENANCE |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Lube/Fuel Truck | $ 560 | $ - | $ - | $ - | $ - | $ - | $ 280 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 840
Mechanics Truck | $ 187 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 187
Tire Truck | $ 137 | $ - | $ - | $ - | $ 137 | $ - | $ 137 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 411
Total Mine Maintenance | $ 884 | $ - | $ - | $ - | $ 137 | $ - | $ 417 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 1,438
OTHER MINE CAPITAL |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Light Plant | $ 44 | $ 29 | $ - | $ 29 | $ - | $ 29 | $ - | $ 29 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 161
Mobile Radios | $ 55 | $ 19 | $ - | $ 40 | $ 1 | $ - | $ 28 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 143
Shop Equipment | $ 491 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 491
Engineering & Office Equipment | $ 200 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 200
Water Storage (Dust Suppression) | $ 98 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 98

Base Radio & GPS Stations | $ 105 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 105
Unspecified Miscellaneous Equipment | $ 150 | $ - | $ - | $ 2,000 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 2,150
Access Roads - Haul Roads - Site Prep | $ 175 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 175
Light Vehicles | $ 540 | $ - | $ - | $ 591 | $ - | $ - | $ 456 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 1,587
Total Other Mine Capital | $ 1,858 | $ 48 | $ - | $ 2,661 | $ 1 | $ 29 | $ 484 | $ 29 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 5,110
CAPITALIZED MINE OPERATING COSTS |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Pre-Stripping Mining Cost | $ 44,235 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 44,235
Construction Costs | $ 2,189 | $ 1,123 | $ 6,857 | $ 5,082 | $ 4,549 | $ 2,469 | $ 5,083 | $ 4,531 | $ 10,658 | $ - | $ 10,671 | $ 8,249 | $ 1,525 | $ - | $ - | $ 62,986
Reclamation (Occurs in Years 13 and 14) | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 25,958
Total Capitalized Mining Costs | $ 46,425 | $ 1,123 | $ 6,857 | $ 5,082 | $ 4,549 | $ 2,469 | $ 5,083 | $ 4,531 | $ 10,658 | $ - | $ 10,671 | $ 8,249 | $ 1,525 | $ - | $ - | $ 133,178
CAPITAL SUMMARY |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Primary Mining Equipment | $ 73,443 | $ 66,462 | $ - | $ 71,988 | $ - | $ - | $ 714 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 212,607
Support Equipment | $ 9,738 | $ 1,036 | $ - | $ - | $ 21 | $ 21 | $ 8,221 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 19,038
Blasting | $ 52 | $ - | $ - | $ - | $ 52 | $ - | $ - | $ 52 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 157
Mine Maintenance | $ 884 | $ - | $ - | $ - | $ 137 | $ - | $ 417 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 1,438
Other Mine Capital | $ 1,858 | $ 48 | $ - | $ 2,661 | $ 1 | $ 29 | $ 484 | $ 29 | $ - | $ - | $ - | $ - | $ - | $ - | $ - | $ 5,110
Capitalized Mine Operating Costs | $ 46,425 | $ 1,123 | $ 6,857 | $ 5,082 | $ 4,549 | $ 2,469 | $ 5,083 | $ 4,531 | $ 10,658 | $ - | $ 10,671 | $ 8,249 | $ 1,525 | $ 4,354 | $21,603- | $ 133,178
Total - All Mining Capital | $ 132,401 | $ 68,670 | $ 6,857 | $ 79,731 | $ 4,760 | $ 2,519 | $ 14,919 | $ 4,613 | $ 10,658 | $ - | $ 10,671 | $ 8,249 | $ 1,525 | $ 4,354 | $ 21,603 | $ 371,529

21.1.1.1Drilling and Blasting

Primary drilling equipment capital is based on equipment quotations for a total of 15 Atlas Copco Pit Viper 235 blast-hole drills required through the life of mine. Eight of the drills will be purchased at the start of mining in Year 1, an additional three drills purchased in Year 1, and then four additional drills will be purchased in Year 3 at a cost of US$2,140,100 each (including shipping and assembly). The cost of the drills was provided by EMG LLC.

In addition to the production drills, smaller 45K pull-down drills will be used for pre-split drilling. These will use 165mm bits and will cost approximately US$713,900 each. One drill is purchased in Year -1 and a replacement drill has been planned for Year 6.

Quotes for explosives trucks, powder magazines, and bulk ANFO storage has been obtained by TTP. These capital costs are included in the infrastructure costs. Additional, capital expense for a skid loader is provided to be used by the blasting crew for stemming holes. The skid loader would be purchased at an estimated cost of US$52,300 during Year -1 and then two additional units would be purchased in Year 4 and Year 6.

21.1.1.2Loading

Capital costs for loading equipment have been quoted by EMG LLC and include four Hitachi Ex5500 hydraulic shovels and two Caterpillar 994 Loaders. Two of the hydraulic shovels would be purchased during Year -1, with a third being purchased during Year 1. The fourth shovel is purchased in Year 3. The estimated cost for each shovel is US$6,876,700, which includes freight and assembly.

The cost of the 18-cubic meter loaders is based on a quote for a Caterpillar 994 loader, with the first one being purchased at the start of production in Year 1, and the second purchased in Year 2, at a cost of US$4,154,300 each.

21.1.1.3Haulage

The 226-tonne haulage truck costs are based on CAT 793F trucks and were quoted by EMG LLC. Nine trucks are purchased during Year -1, with another 8 trucks purchased in Year 1. Trucks are purchased as they are required through the mine life. The trucks are staged in to allow ramp up of production through each year as they are needed to meet production requirements. The total number of trucks required by year is shown as follows:

Number of

Year Trucks Added Trucks in Use

-1 9 9

1 8 17

3 6 23

4 13 36

5 2 38

Throughout the mine life, a total of 38 trucks are purchased. The number of operating trucks is reduced toward the end of the mine life as haulage requirements are decreased. The cost of each truck is estimated at US$3,770,100, including freight and assembly.

21.1.1.4Mine Support

Capital estimates for mine support equipment include freight and erection. The initial support equipment to be purchased in Year -1 is as follows:

Two Caterpillar D9 track dozers (US$858,200 each quoted by EMG LLC);

One Caterpillar D8 track dozer (US$531,000 each quoted by EMG LLC);

Two Caterpillar 16H motor graders (US$893,800 quoted by EMG LLC);

One Caterpillar 777 truck with a 70K liter water tank (US$1,504,000 quoted by EMG LLC);

Two Caterpillar 834H rubber tire dozer (US$1,036,300 quoted by EMG LLC);

One Caterpillar 392DL excavator (US$315,300 quoted by EMG LLC);

One low-boy trailer complete with a used 60t haul truck to tow it (US$993,600);

One flatbed truck (US$56,700);

Two pit pumps (US$25,500 each);

One rock breaker to be attached to the 392DL excavator as needed (US$39,100); and

Five light plants (US$14,700).

Replacements are purchased for most units in Year 6.

21.1.1.5Maintenance

Capital for mine maintenance equipment includes three fuel/lube trucks (US$279,800 each), one mechanic’s truck (US$187,000 each), and three tire trucks (US$137,000 each). Note that requirements for mechanic’s trucks are reduced due to the assumption of MARC for maintenance. The mechanic’s truck is intended for support of a small number of owner-operated equipment.

An additional US$491,000 has been included for shop equipment / tooling. Shop facilities were estimated by TTP and included in facility capital.

21.1.1.6Mine Facilities

Mine facility capital has been estimated by TTP and is included in facility capital.

21.1.1.7Light Vehicles

Initial capital for light vehicles is estimated to be US$540,000 while sustaining light vehicle capital is US$1,047,200. Initial and sustaining light vehicle capital is shown in Table 21‑4.

Table 21‑4: Estimated Mine Light Vehicle Capital (US$)

| Type | Initial Capital (US$) | Sustaining Capital (US$)
Quantity | Unit Cost | Ext. Cost | Quantity | Unit Cost | Ext. Cost
MINE DEPARTMENT |  |  |  |  |  |  |
Mine Superintendent | 3/4 ton 4wd Pickup | 1 | $ 36,000 | $ 36,000 | 2 | $ 36,000 | $ 72,000
Shift Foreman | 4wd Pickup | 2 | $ 25,600 | $ 51,200 | 4 | $ 25,600 | $ 102,400
Trainer | 4wd Pickup | 1 | $ 22,400 | $ 22,400 | - | $ 22,400 | $ -
Blasting | 4wd Pickup | 1 | $ 25,600 | $ 25,600 | 2 | $ 25,600 | $ 51,200
Blasting | 1 ton 4wd Pickup | 1 | $ 25,600 | $ 25,600 | 2 | $ 25,600 | $ 51,200
Crew Vans | 3/4 ton Passenger Van | 2 | $ 56,000 | $ 112,000 | 4 | $ 56,000 | $ 224,000
ENGINEERING |  |  |  |  |  |  |
Chief Engineer | 4wd Pickup | 1 | $ 25,600 | $ 25,600 | 2 | $ 25,600 | $ 51,200
Short Range Planning | 4wd Pickup | 1 | $ 22,400 | $ 22,400 | 2 | $ 22,400 | $ 44,800
Survey | 4wd Pickup | 1 | $ 25,600 | $ 25,600 | 3 | $ 25,600 | $ 76,800
GEOLOGY |  |  |  |  |  |  |
Chief Geologist | 4wd Pickup | 1 | $ 25,600 | $ 25,600 | 3 | $ 25,600 | $ 76,800
Ore Control | 4wd Pickup | 1 | $ 22,400 | $ 22,400 | 3 | $ 22,400 | $ 67,200
Samplers | 4wd Pickup | 1 | $ 22,400 | $ 22,400 | 2 | $ 22,400 | $ 44,800
MINE MAINTENANCE |  |  |  |  |  |  |
Maintenance Superintendent | 4wd Pickup | 2 | $ 36,000 | $ 72,000 | 3 | $ 36,000 | $ 108,000
Mechanics / Labor | 4wd Pickup | 2 | $ 25,600 | $ 51,200 | 3 | $ 25,600 | $ 76,800
Total |  | 18 |  | $ 540,000 | 35 |  | $ 1,047,200

21.1.1.8Other Mine Capital

Other miscellaneous capital includes mobile radios for mobile equipment (US$1,000 per unit), engineering and office equipment ($200,000 US), water storage for dust suppression (US$97,900), GPS stations and surveying equipment (US$105,000), and other unspecified miscellaneous equipment (US$150,000). At the end of year three, Mt Todd personnel will take over the maintenance of equipment. Accordingly, US$2,000,000 as unspecified equipment has been added in year three for additional maintenance equipment.

21.1.2CIP Process and Infrastructure (TTP)

Please note that this Section describes costs in Australian Dollars (AUD).

TTP’s capital cost estimates (CCEs) are based on an Enhanced Factored Cost Estimate (EFCE) methodology, which features higher confidence levels around Contingency Provision and Management Reserve. The capital estimates are supported by the design work carried out throughout the study including process documentation, schematics, general arrangement drawings, 3D models and calculations. Note all currencies are in Australian Dollars. All capital estimated by TTP is summarized in Table 21‑5.

Table 21‑5: Estimated Capital Cost Summary (AUD000s)

Capital Cost | Initial Capital (AUD000s)
Facility 1000 – Geology | 0.000
Facility 2000 – Mine Infrastructure | 13,872
Facility 3000 – Process Plant | 443,227
Facility 4000 – Project Services | 14,608
Facility 5000 – Project Infrastructure | 26,036
Facility 6000 – Permanent Accommodation | 0,086
Facility 7000 – Site Establishment & Early Works | 24,654
Facility 8000 – Management, Engineering, EPCM Services | 80,925
Facility 9000 – Preproduction Costs | 13,909
Subtotal | 617,316
Direct | 497,828
Indirect | 119,488
Subtotal | 617,316
Contingency Provision (14.6%) | 88,343
TOTAL | 705,659

The total capital cost, base cost plus contingency provision, represents the expected cost for the project, with approximately a 55% confidence level of completion within cost. This estimate has an accuracy range of approximately -0 to 15% based on the expected cost. At the upper limit of the accuracy range, there is an 85% confidence level of completion within cost.

Typically, the EPCM Project Manager would initially receive Owner’s approval for expenditure up to the expected cost (i.e., this is the initial project budget).

However, funding arrangements would also need to be in place for expenditure up to the 85% confidence level. This additional funding is commonly referred to as Management Reserve. The selection of Management Reserve quantity will rest with Vista, and will be determined by Vista’s attitude to risk.

21.1.2.1Exclusions

The TTP scope of work is a significant part of the overall Project scope, although other parties have compiled capital costs for other areas on behalf of Vista. The potential impacts of possible price or labor rate fluctuations or currency exchange rate fluctuations are the role of a qualified actuary and should be covered by Vista in its standard business practices.

21.1.2.2Capital Cost Estimating Methodology

A PFS design has been developed for a 50,000 tpd plant as the basis for an EFCE. The EFCE approach uses a combination of bottom-up calculations and factoring methods for each area in the estimate. The methods used to estimate capital in the CCE are summarized in the following sections.

The EFCE for the process plant features the methodology shown in the table below.

Table 21‑6: CCE Methodology for Facility 3000 – Process Plant

Bulk Commodity | Base Case
Mechanical Equipment | A detailed mechanical equipment list, with supply and installation pricing based on budget quotations and internal body of knowledge
Concrete | MTOs based on 3D model and unit rates and factored up from 33,000 tpd option and unit rates
Structural Steel | MTOs based on 3D model and unit rates and factored up from 33,000 tpd option and unit rates
Platework | MTOs from previous projects a unit rates
Tankage | MTOs based on preliminary design calculations and unit rates
Piping | Percentage factor of the supplied mechanical equipment supply price, assessed on an area by area basis.
Electrical | Percentage factor based on total mechanical equipment supply price
Instrumentation and Control | Costs factored up from 33,000 tpd case

Subsequently estimate factors, by area, were back calculated for each bulk commodity as a percentage of the mechanical equipment supply cost estimate. In turn, the resultant estimate factors were critiqued against published data and industry experience.

21.1.2.3Other Area Capital Cost Estimates

The EFCE for all areas outside of the process plant features the methodology shown in Table 21‑7.

Table 21‑7: Methodology for Other Areas of the Capital Cost Estimate

Area / Sub Area | Base Case
FACILITY 2000 – MINE
Area 2300 – Mine Support Facilities | Drawings developed for the buildings and priced largely on a building per square meter basis
Area 2400 – Mine Support Services | Drawings developed for the buildings and priced largely on a building per square meter basis
FACILITY 4000 – PROJECT SERVICES
Area 4100 – Water Supply | Sub-area 4110 – Water Supply WTP estimated by Tetra Tech Golden Office Sub-Area 4120 – Raw Water Distribution was estimated by mechanical equipment costs and factoring of bulk commodities
Area 4200 – Power Supply | Based on length of power distribution cables and trenching, and overhead power lines using rates developed from previous projects. Power – Generation (7.3.1) is by Power Engineers
Area 4300 – Communications | Based on MTOs for the fiber optic cables, phones and telemetry using budget pricing and rates developed from previous projects.
Area 4200 – Power Supply | Based on length of power distribution and overhead power lines and rates developed from previous projects
Area 4300 – Communications | Based on MTOs and provisional sums for the fiber optics, phones and telemetry
Area 4400 – Tailings Dam | Estimated by Tetra Tech Golden Office
Area 4500 – Waste Disposal | Based on provisional sums for sewerage services from previous project designs
Area 4600 – Plant Mobile Equipment | Vendor pricing of the proposed fleet for plant operation
FACILITY 5000 – PROJECT INFRASTRUCTURE
Area 5100 – Site Preparation | Based on MTOs from preliminary drawings and rates developed from first principles
Area 5200 – Support Buildings | Drawings developed for the buildings and priced largely on a building per square meter basis
Area 5300 – Access Roads, Parking and Laydown | Provisional sums based on miscellaneous road and culvert repairs
Area 5400 – Heavy Lift Cranage | Based on the proposed fleet for plant construction and rates from previous project experience
Area 5600 – Bulk Transport | Based on the mechanical equipment cost for the weigh bridge and MTO for concrete
Area 5800 – Communications | Based on budget pricing and quantities provided by Vista
FACILITY 6000 – PERMANENT ACCOMMODATION
Area 6100 – Personnel Transport | Based on unit rates for bus shelters with an allowance for the small amount of concrete required
FACILITY 7000 – SITE ESTABLISHMENT AND EARLY WORKS
Area 7300 – Construction Camp | Based on MTOs for the access roads and site works and vendor quotes for the camp and operation

FACILITY 8000 – MANAGEMENT, ENGINEERING, EPCM SERVICES
Area 8100 – EPCM Services | Factored up from 33,000 tpd Case based on the Total Direct Costs
Area 8200 – External Consultants/Testing | Provisional Sums based on previous project experience
Area 8300 – Commissioning | Process Plant commissioning costs based on 3% of the total mechanical equipment costs. Provisional Sums allowed for Mine, Project Services and Infrastructure commissioning.
Area 8400 – Owners Engineering/Management | Based on 2% of the project direct costs.
Area 8800 – License, Fees and Legal Costs | Based on 0.5% of the project direct costs.
Area 8900 – Project Insurances | Based on 0.5% of the project direct costs.
FACILITY 9000 – PRE-PRODUCTION COSTS
Area 9100 – Preproduction Labor | Based on 0.25% of the project direct costs.
Area 9200 – Commissioning Expenses | Based on 0.5% of the project direct costs.
Area 9300 – Capital Spares | Based on 5% of the mine and process plant mechanical equipment costs.
Area 9400 – Stores and Inventories | Based on 1% of the mine and process plant mechanical equipment costs
Area 9800 – Contingency Provision | Priced based on a weighted average of the contingency of each facility
Area 9900 – Management Reserve Provision | A weighted average Management Reserve of 20% was allowed for, the selection of Management Reserve quantity will rest with Vista, and will be determined by Vista’s attitude to risk.

21.1.2.4Construction Labor Rates

A detailed calculation of composite, direct man-hour site rates has been carried out using TTP standard templates. The calculation is based upon current ordinary time wages for various classes of labor including direct supervision, to which the following factor may apply; site allowance, tool allowance, leave provisions, taxes and insurances, overtime, etc. This develops a gang rate that is combined with costs of incumbent support equipment (such as light vehicles, light mobile cranes, small tools, consumables, first-aid facilities and accommodation) and management support to arrive at an all-purpose site gang rate.

The construction labor rates developed for the CCE include the following construction contractors:

Concrete;

Structural, Mechanical and Piping (SMP); and

Electrical and Instrumentation (E&I).

BASE LABOR RATES

The base labor rate includes the direct labor allocated for the installation of equipment and bulk commodities. Base pay rates were derived from award rates for similarly sized projects currently underway in the North West of Western Australia and in the Northern Territory. These are considered to be the benchmark for the area, including Mt Todd. Allowances were made for overtime loadings above a 36-hour week including time and a half for the initial 12 hours overtime, followed by double time for the final 17 hours overtime, to provide for a 65-hour working week. The rates were averaged over a standard mix of trades, to produce a composite rate per man per hour.

The base labor rates were developed to include items listed below:

All direct payments including the site allowances and special project allowances for straight time and overtime worked for personnel;

Overtime at penalty rates;

Provision for holiday leave and loadings thereon;

Provision for sick leave;

Provision for cost of travel time to site and return travel on job completion;

Provision for additional manpower turnover, bereavement leave and miscellaneous paid non-work days;

Payroll tax;

Workers compensation insurance;

Superannuation considerations; and

Industry redundancy payments.

A rest and recreation (R&R) loading was also added to the composite rate, together with a 15% contractors allowance for overheads and margin to produce the base labor rate for each contractor type.

CONTRACTOR INDIRECT RATES

The contractor indirect rate is a combination of costs associated with indirect contractor personnel, contractor vehicles, contractor overheads and construction plant equipment. An estimate of construction contract duration and installation hours was based on the EPCM schedule and bulk quantity development.

The contractor indirect rates were developed to include the items listed below:

Project Management personnel;

Construction Supervision personnel;

Site Quality Assurance and Control personnel;

Site Health, Safety, Environmental personnel;

Other indirect labor (stores officer, surveyor, etc.);

Contractor vehicles for the Project Management team;

Office accommodation;

Workshop and stores facilities;

Staff travel including airfares ;

Office overheads; and

Vehicle consumables.

Provisions for the accommodation and messing are also not included in the indirect contractor rates. This is allowed for in the construction camp cost estimate to supply and operate the camp.

Although they are considered indirect costs, construction plant equipment rates are estimated separately to include the following:

Construction plant equipment mobilization / demobilization;

Construction plant management support; and

Construction plant and equipment.

The provision for task specific heavy lift cranes >50 tonnes were not included in the indirect contractor rates build-up; instead it was allowed for in a separable line item in the chart of accounts (CoA).

CONSTRUCTION GANG RATES

The overall site construction gang rates were developed by summing the base labor rate, contractor indirect rates and construction plant rates to provide an overall site construction gang rate for Concrete, SMP and E&I contractors as shown in Table 21‑8 below.

Table 21‑8: Estimated Construction Gang Rate Development (AUD)

Contractor | Base Labor Rate (AUD/hr) | Contractor Indirect Rate (AUD/hr) | Construction Plant Rate (AUD/hr) | Construction Gang Rate (AUD/hr)
Concrete | $ 88.37 | $ 23.98 | $ 24.36 | $ 136.71
SMP | $ 97.29 | $ 45.47 | $ 20.91 | $ 163.68
E&I | $ 102.49 | $ 56.27 | $ 24.07 | $ 182.83

21.1.2.5Mechanical Equipment

The supply costs comprise the direct mechanical equipment cost plus the cost for freight to site. Installation costs are estimated based on an evaluation of installation hours multiplied by the SMP contractor gang rate. These estimating methods are discussed in the following sections.

EQUIPMENT COSTS

The basis for estimating the mechanical equipment supply costs was largely based on budgetary pricing from vendors. The vendors were provided with preliminary specifications and/or data sheets for major equipment items. The budget quotations received from vendors are expected to have an accuracy equal to ±10%.

All other minor equipment items were priced from a TTP’s database of costs from recent similar sized projects. The basis of the supply cost estimate for each mechanical equipment line item is documented in the process plant CCE.

FREIGHT COSTS

Several methods were used to determine and validate the allowance for delivery costs of mechanical equipment to site. These methods included:

Quotes provided by the manufacturer;

Estimates based on the weight and volume of the load;

Estimates based on published and in-house guides for similar installations; and

Estimates based on a validated percentage of the mechanical equipment cost (determined to be 9% of the supply price).

INSTALLATION HOURS

Several methods were used to determine and validate the installation hour allowance for mechanical equipment. These methods included:

Quotes provided by the manufacturer;

Estimates based on the weight of the equipment; and

Estimates based on published and in-house guides for similar installations.

The installation hour estimates for large process equipment (>3000 man hours/ equipment) including the crushers, HPGRs, ball mills, ISAmills and thickeners were reviewed in detail against historical records and published guidelines.

21.1.2.6Quantity Development and Unit Rates

The basis for the development of supply and installation costs of bulk commodities is discussed in the following section. Bulk commodities include civil, concrete, structural steel, etc. which will be used in the construction of the process plant. These costs were largely derived based on an estimate of material take-off (MTO) quantities which were multiplied by a unit rate for each type of material. The unit rates were calculated using TTP standard methods including obtaining current market rates from contractors, historical data and reference books (Rawlinsons) and comprise of allowances for supply of the raw material, fabrication, freight and erection.

CIVIL

Preliminary bulk earthwork quantities were estimated using civil 3D modelling software 12D Model. The 12D Model accurately calculates earthworks volumes utilizing the existing topography and proposed design levels. Structural excavation and backfill required for concrete structures are included in the concrete quantities. Trenching requirements for underground utilities distribution were determined from service plans. Stormwater drainage quantities were determined from the civil site plan with vee-drains alongside plant roads directing surface run-off beneath roads via corrugated steel culverts. All quantities were categorized by standard type of work classification.

Unit rates for this work classification were developed from the in-house rates database. This rates database is constantly maintained so as to be current and has proven to be sufficiently accurate over several recent projects. The availability of water and local earthworks materials was taken into account in to the development of unit rates.

CONCRETE

Concrete quantities for foundations and ground slabs for all equipment and structures in the process plant were calculated by 3D models. Concrete quantities were categorized by standard classes of concrete including spread/pad footings, strip footings, raft footings, ring beams, ground slabs, walls, sumps and pits, etc.

Unit pricing was obtained from industry sources by standard classification, each having an assessment of formwork, props, bracing reinforcing, embedment’s, joints in slabs plus a miscellaneous allowance for curing, formwork hardware and other sundries. Concrete supply was costed at a rate deemed to include plant control testing, some admixtures and out of hours pouring. A wastage factor was included in the rates. A Contractor’s mark-up was also applied to all materials. Direct labor unit man-hours were sought from industry sources and checked against historical data and various published references.

STRUCTURAL STEEL

Steel quantities were categorized by standard classes of steel including light, medium, heavy and very heavy. There are also provisions made for grating, handrailing and stair treads.

Unit rates for the supply and installation of structural steel were calculated using TTP Standard templates. Supply of steel was based on rates quoted from Thai fabricators from a similar project. The supply rate includes provisions for steel supply, shop drawings, shop fabrication, painting and freight to site. Estimates for the installation costs of structural steel are based on estimates of erection hours and the SMP gang rate.

PLATEWORK

Quantities of steel required for custom designed platework was calculated using the 3D models. The cost items for platework includes plate thicknesses of <10mm, 12-20mm and floor plate of 6mm and also allowances for Bisalloy or rubber lining where applicable.

Unit rates for platework were provided by Thai fabricators as is described for structural steel.

TANKAGE

Quantities of steel required for custom designed tankage were quantified by structural engineers using TTP standard spread sheets to determine the required tank shell and base thickness in accordance with the provisions of API 650. This includes allowances for the mass of steel for shell plates, top rings, and base plates.

Unit rates for supply and installation of tankage were based on Thai fabricated steel using the same methodology as is described for structural steel. There are two classes of tankage allowed for in the CCE including shop fabricated and site erected tanks (assumed to be greater than 7m in diameter).

PIPING

The estimate for the supply and installation of process piping was factored based on a percentage of the supplied mechanical equipment price, assessed on an area by area basis. These percentages were based on in-house and industry typical piping allowances for similar gold plants. These factors were validated with values reported in published guidelines.

ELECTRICAL

The estimate for the supply and installation of electrical components for the process plant was factored based on a percentage of the total Mechanical Equipment supply and installation cost. These percentages were based on in-house and industry typical electrical allowances for similar gold plants.

INSTRUMENTAL AND CONTROL

The estimate for the supply and installation of the Instrumentation and Process Control System for the Alternate Case process plant was estimated based on preliminary P&IDs and equipment lists based on a highly automated gold plant with all field instruments marshalled to remote Input / Output (I/O) cabinets. The estimate of Instrumentation and Control costs for the Base Case was up-scaled based on an expected 30% increase in equipment, I/O, programming and instrumentation.

21.1.2.7Indirect Costs

CONSTRUCTION CAMP

An estimate of the construction facilities was developed from previous project experience for the various scopes of work. This includes a breakdown of costs for contractor preliminaries, transportable building (supply and install) and establishing the infrastructure, power supply, communications and water supply. It also includes allowances for removal of the infrastructure following completion. Provisions were made for the operation of the camp based on a man-day rate. The man-day rate applied is based on budget enquiries for appropriate contractors.

EPCM SERVICES

Engineering, drafting and documentation functions are task and deliverable related. Hence, their estimates were based on task and deliverable identification with time estimates based in industry experience. Procurement activities were estimated from hours related to purchasing, expediting, inspection and transport functions derived by time involvements, and then checked against industry experience. Management, administrative and project engineering functions are mostly time-related and were assessed by title, rate and man-months of key personnel and other staff proposed.

To the extent possible, site office items were detailed and estimated on an item-by-item basis. Management, supervisory and administrative staffing were estimated on an hours basis.

EXTERNAL CONSULTANTS AND TESTING

Cost allowances for Environmental, Human Resources and Industrial Relations, and Health and Safety consultants are based on industry experience of required manning and market contract values.

OTHER INDIRECT COSTS

The following costs were calculated based on industry validated percentages of the total direct costs of the Project:

Owners engineering / management;

License, fees and legal costs;

Project insurances; and

Pre-production labor.

The following costs were calculated based on industry validated percentages of the mechanical equipment supply cost for the Project:

Commissioning Expenses;

Capital Spares; and

Stores and Inventories.

21.1.2.8Contingency Provision

The contingency provision is an allowance added to an estimate to provide for costs which cannot be estimated due to inadequate information, but which are known to be implicit in the scope.

The contingency provision represents costs which are expected to be incurred to complete the project and must be regarded as part of the total funds placed under the direct control of the project manager.

The contingency provision includes an allowance for:

Unidentified items not included in the quantity calculations or equipment lists, due to lack of knowledge, but implicit in the scope.

Small changes, arising from detailed design, which normally occur during the course of the project, as knowledge becomes firmer.

Design Omissions.

Changes in concept, scope or production rates which depart from those on which the estimate has been based require a new estimate. These changes are not allowed for in the contingency.

21.1.3Power Plant (POWER Engineers)

Please note that this Section describes costs in Australian Dollars (AUD).

Estimated capital costs compiled for this study cover direct and indirect costs of power station construction including equipment quoted from suppliers, material quantities estimated from the preliminary design, and installation labor and supervision with local material and labor costs applied to those estimates. Power plant estimated capital costs are shown in Table 21‑9 in Australian dollars.

Table 21‑9: Estimated Power Station Installed Capital Cost Summary (AUD)

| 69.6 MW (Seven Reciprocating Engines) (AUD)
Reciprocating Engines | $ 30,189,744
BOP Equipment | $ 9,644,816
Mechanical, Civil, & Electrical Direct Costs | $ 47,870,887
Engineering Fees | $ 6,410,256
Contractor’s Fees | $ 5,540,385
Taxes, and Other Indirect Costs | $ 10,171,381
Total Installed Costs AUD | $ 109,827,468
Capital Cost per Net Installed Capacity AUD/kW | $ 1,541

21.1.4Mine Dewatering (Tetra Tech)

Mine dewatering capital costs are based on direct vendor quotes or Tetra Tech in-house estimates and include 5% indirect costs.

Table 21‑10: Estimated Mine Dewatering Capital Cost Summary (US$000s)

WBS No. | Description | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
2501 | PPD Dewatering | $608 | $0 | $608
2502 | Self-Priming Pump on Pontoon | $0 | $0 | $0
2503 | Pump | $0 | $395 | $395
2504 | Piping in Pit | $0 | $187 | $187
2505 | Piping from Pit to PWP | $0 | $285 | $285
2506 | Electrical | $0 | $67 | $67
2507 | Indirects | $0 | $37 | $37
| 2500 Mine Dewatering/Drainage | $608 | $972 | $1,580

21.1.5Reclamation and Closure (Tetra Tech)

Costs for reclaiming major facilities at the Project were estimated using closure material quantities based on ultimate designs and following the closure plans discussed above. Capital costs for reclamation are estimated at US$144 million for LoM.

Table 21‑11: Estimated Reclamation Capital Cost Summary (US$000s)

WBS No. | Description | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
2901 | Heap Leach Pad | $0 | $987 | $987
2902 | Low Grade Ore Stockpile | $112 | $705 | $818
2903 | TSF 1 | $0 | $29,579 | $29,579
2904 | TSF 2 | $0 | $34,379 | $34,379
2905 | WRD (GCL Cover) | $0 | $35,347 | $35,347
2906 | Process Plant Area | $0 | $12,228 | $12,228
2907 | Soil Stockpiles | $0 | $250 | $250
2908 | Mine Roads | $0 | $451 | $451
2909 | Batman Pit | $0 | $1,238 | $1,238
2910 | Passive Treatment Systems | $0 | $3,769 | $3,769
2911 | Indirect Costs | $17 | $24,945 | $24,961
| 2900 Mine Closure | $129 | $143,877 | $144,007

21.1.6Water Treatment Plant (Tetra Tech)

Water treatment plant capital costs are based on direct vendor quotes or Tetra Tech in-house estimates, initial capital costs are estimated at US$14.4 million and no sustaining capital improvements are expected.

Table 21‑12: Estimated Water Treatment Plant Capital Cost Summary (US$000s)

WBS No. | Description | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
4111 | Earthwork | $490 | $0 | $490
4112 | Concrete | $225 | $0 | $225
4113 | Building | $1,400 | $0 | $1,400
4114 | Equipment | $8,466 | $0 | $8,466
4115 | Mechanical | $1,460 | $0 | $1,460
4116 | Electrical and Instrumentation | $2,000 | $0 | $2,000
4117 | Construction Management | $400 | $0 | $400
| 4110 Water Treatment Plant | $14,441 | $0 | $14,441

21.1.7Raw Water Dam (Tetra Tech)

Raw water dam capital costs are based on direct vendor quotes or Tetra Tech in-house estimates, initial capital costs are estimated at US$2.4 million and no sustaining capital improvements are expected.

Table 21‑13: Estimated Raw Water Dam Capital Cost Summary (US$000s)

WBS No. | Description | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
4121 | Site & Foundation Prep | $601 | $0 | $601
4122 | Embankment Construction | $858 | $0 | $858
4123 | Relocate Outlet Downstream | $11 | $0 | $11
4124 | Spillway Construction | $271 | $0 | $271
4125 | Other Construction Costs | $279 | $0 | $279
4126 | Pump Operation Cost | $0 | $0 | $0
4127 | Engineering Procurement | $214 | $0 | $214
4128 | Construction Management | $87 | $0 | $87
4129 | Temporary Construction Facilities | $53 | $0 | $53
| 4120 Raw Water Dam | $2,374 | $0 | $ 2,374

21.1.8Tailings Storage Facilities (Tetra Tech)

Tailings storage facility capital costs are based on direct vendor quotes or Tetra Tech in-house estimates. Initial capital costs are estimated at US$8 million with sustaining capital of US$83 million.

Table 21‑14: Estimated Tailings Storage Facility Capital Cost Summary (US$000s)

WBS No. | Description | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
4410 | TSF 1 |  |  |
1 | Site & Foundation Preparation | $563 | $2,165 | $2,728
2 | Embankment Construction | $315 | $5,500 | $5,815
3 | Downstream Embankment Toe Drain | $1,023 | $0 | $1,023
4 | Tailings Delivery & Return Pipelines | $2,237 | $211 | $2,448
5 | Return Water Ponds | $175 | $0 | $175
6 | Diversion Channels | $0 | $0 | $0
7 | Equipment Purchase | $1,927 | $0 | $1,927
8 | Mobilization | $851 | $1,074 | $1,925
9 | EPCM | $936 | $1,181 | $2,117
10 | Instrumentation | $0 | $200 | $200
| 4410 TSF 1 | $8,028 | $10,330 | $18,359
4420 | TSF 2 |  |  |
1 | Site & Foundation Preparation | $0 | $14,643 | $14,643
2 | Underdrain Construction | $0 | $1,398 | $1,398
3 | Downstream Toe Drain | $0 | $651 | $651
4 | Embankment Construction | $0 | $8,449 | $8,449
5 | Impoundment Liner | $0 | $18,433 | $18,433
6 | Overdrain & Reclaim Sump/Pond Construction | $0 | $1,968 | $1,968
7 | Tailings Delivery & Return Pipelines | $0 | $7,164 | $7,164
8 | Surface Water Management | $0 | $1,499 | $1,499
9 | Equipment Purchase | $0 | $1,927 | $1,927
10 | Mobilization | $0 | $7,654 | $7,654
11 | EPCM | $0 | $8,420 | $8,420
12 | Instrumentation | $0 | $320 | $320
| 4420 TSF 2 | $0 | $72,527 | $72,527
| 4400 Tailings Dam | $8,028 | $82,857 | $90,886

21.2Operating Costs

LoM operating costs requirements are estimated to be US$14.18/t-milled as summarized in Table 21‑15.

Table 21‑15: Estimated LoM Operating Costs (US$)

Description | US$/t-milled
OPEN PIT MINE-OWNER |
Mine General Service | $0.07
Mine Maintenance | $0.12
Engineering | $0.05
Geology | $0.03
Drilling | $0.73
Blasting | $1.20
Loading | $0.62
Hauling | $2.89
Mine Support | $0.37
Mine Dewatering | $0.01
Open Pit Mine-Owner | $6.08
CIP PROCESS PLANT |
Labor | $0.89
3100-Crush/Screen/Stockpile | $0.19
3200-Reclaim & HPGR | $0.49
3300-Classification & Grinding | $2.66
3400-Pre-Leach,Thick/Aeration/CIP | $0.15
3500-Desorption, Gold Room | $0.02
3600-Detox & Tailings Pumping | $0.08
3700-Reagents | $2.11
3800-Plant Services | $0.04
Mining, Infrastructure & Misc | $0.05
General Consumables | $0.01
Plant Mobile Equipment | $0.01
Plant Gas Consumption | $0.02
CIP Process Plant | $6.71
Project Services | $0.16
G&A | $1.22
Operating Costs | $14.18

21.2.1Mining (MDA)

Annual mine operating costs have been estimated based on personnel requirements and equipment hourly costs. Table 21‑16 summarizes annual mine operating costs before the allocation of capitalized mining (capitalized mining is included in Section 21.1.1 – Mining (MDA)). Costs are provided based on functionality (drilling, blasting, loading, hauling, support, mine general services, mine maintenance, engineering, and geology).

The following subsections describe the operating cost estimate by functionality. Note that these costs are described before the allocation of costs to capital for pre-stripping, mining of waste for construction purposes, and mining and waste re-handle of material to be used for reclamation.

The total average mining cost for the Base Case is estimated to be US$2.02/t mined. The average LOM cost per tonne moved is US$1.71/t, which includes mined tonnage plus ore and waste rehandle. After the allocations of capital and rehandle, the net mine operating cost is US$1.84/tonne based on a net operating cost of US$1.341 million and 731 million tonnes.

21.2.1.1Drilling Costs

The average life-of-mine drilling cost is estimated to be US$0.23/t mined after allocation of drilling costs for pre-stripping and tailings construction. This includes maintenance allocations based on MARC cost assumptions.

21.2.1.2Blasting Costs

The average life-of-mine blasting cost is estimated to be US$0.38/t mined.

21.2.1.3Loading Costs

The average life-of-mine loading cost is estimated to be US$0.21/t mined, or US$0.18/t moved. The cost per tonne moved includes the re-handle of ore and waste from stockpiles at the end of the mine life. Maintenance costs assume the use of MARC costs provided by EMG LLC.

21.2.1.4Haulage Costs

The average life-of-mine haulage cost is estimated to be US$0.97/t mined or US$0.83/t moved. The cost per tonne moved includes re-handling of stockpiled ore and waste at the end of the mine life. Maintenance costs assume the use of MARC costs provided by EMG LLC.

21.2.1.5Mine Support Costs

Mine-support costs include the operation of all of the mine-support equipment. The average life-of-mine support cost is estimated to be US$0.13/t mined or US$0.11/t moved. The cost per tonne moved includes support during re-handling of stockpiled ore and waste at the end of the mine life. Maintenance costs assume the use of MARC costs provided by EMG LLC.

Support costs also include the costs to reinforce the eastern high wall in the ultimate pit with bolts and mesh as recommended by Call & Nicholas.

21.2.1.6Mine Maintenance Costs

Most maintenance will be done under a MARC cost structure for the first two years. Beyond this it was assumed that Vista would take over all maintenance tasks. The vendor with the contract will be expected to supply mechanics and maintenance parts for major equipment repair. Costs associated with the contract have been included in the equipment hourly cost. During the first two years when the contractor will provide MARC services, Vista will employ one maintenance planner.

After the first two years, the MARC costs for the parts and labor were still used for maintenance cost estimates of mining equipment, but the anticipated overhead and profit of the contractor would be removed. For this reason, during Year 3 and beyond, the MARC costs were multiplied by 85%. MDA has assumed that this will require hiring of a maintenance foreman and an additional maintenance planner.

Owner mine-maintenance costs have been included to cover items not covered by the MARC costs, as well as supervision. This includes salaries for a Maintenance Superintendent and Maintenance Planner to track costs associated with the contract. Tiremen will be hired by the owner to maintain all equipment tires, and servicemen will be hired to keep equipment fueled and lubricated. An allocation for shop laborers has been included for light maintenance of facilities.

The average life-of-mine mine-maintenance cost is estimated to be US$0.04/t mined or US$0.04/t moved. This does not include the specific parts and labor allocations to individual equipment, as those costs are allocated to the equipment and the cost center for which the equipment is used.

21.2.1.7Mine General Services, Engineering and Geology Costs

Mine General costs include salaries for a Mine Manager, Mine Clerk, Shift Foremen, and trainers. Mine general costs also include an allocation for various supplies and office costs.

Engineering and geology services are provided to maintain surveying, mine planning, and ore control for the operations. The average life-of-mine general services, Engineering, and Geology costs are estimated to be US$0.06/t mined or US$0.04/t moved.

Table 21‑16: Estimated Annual Mine Operating Costs (US$)

| Units | Pre-Prod | Yr 1 | Yr 2 | Yr 3 | Yr 4 | Yr 5 | Yr 6 | Yr 7 | Yr 8 | Yr 9 | Yr 10 | Yr 11 | Yr 12 | Yr 13 | Yr 14 | Yr 15 | Total
MINED TONNAGE |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Ore to Mill | k tonnes | - | 10,486 | 15,882 | 13,532 | 15,991 | 9,004 | 6,427 | 10,429 | 14,965 | 17,750 | 17,750 | 2,895 | - | - | - | - | 135,113
Ore to Stkpl | k tonnes | 10,437 | 2,688 | 7,797 | 6,580 | 18,157 | 1,840 | - | - | - | 4,883 | 20,193 | - | - | - | - | - | 72,574
Total Ore Mined | k tonnes | 10,437 | 13,174 | 23,679 | 20,112 | 34,149 | 10,843 | 6,427 | 10,429 | 14,965 | 22,633 | 37,943 | 2,895 | - | - | - | - | 207,687
Re-handle Ore | k tonnes | - | 7,264 | 1,868 | 4,266 | 1,759 | 8,746 | 11,323 | 7,369 | 2,785 | - | - | 14,903 | 17,750 | 7,895 | - | - | 85,928
Re-handle Waste | k tonnes | 525 | - | 1,383 | - | - | - | - | - | - | - | 6,229 | 6,278 | 1,171 | 1,480 | 7,256 | - | 24,326
Re-handle Sorter Rejects | k tonnes | - | - | - | - | - | - | - | - | - | - | - | - | - | - | 20,769 | - | 20,769
Waste to Dumps | k tonnes | 16,850 | 27,284 | 32,692 | 74,220 | 54,933 | 98,928 | 71,318 | 53,987 | 43,800 | 33,942 | 14,990 | 47 | - | - | - | - | 522,990
Total Tonnes Mined | k tonnes | 27,287 | 40,458 | 56,371 | 94,332 | 89,082 | 109,771 | 77,745 | 64,416 | 58,765 | 56,575 | 52,933 | 2,942 | - | - | - | - | 730,677
Total Tonnes Moved | k tonnes | 27,812 | 47,722 | 59,622 | 98,599 | 90,841 | 118,517 | 89,068 | 71,785 | 61,550 | 56,575 | 59,162 | 24,123 | 18,921 | 9,376 | 28,025 | - | 861,700
Strip Ratio | w:o | 1.61 | 2.07 | 1.38 | 3.69 | 1.61 | 9.12 | 11.10 | 5.18 | 2.93 | 1.50 | 0.40 | 0.02 |  |  |  |  | 2.52
Mined Waste to Construction | k tonnes | 825 | 1,250 | 5,380 | 5,804 | 4,176 | 2,697 | 4,275 | 3,218 | 6,592 | - | 109 | - | - | - | - | - | 34,325
Mined Material for Pre-Production | k tonnes | 14,274 | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | 14,274
Mined Waste for Closure | k tonnes | - | - | - | - | - | - | - | - | - | - | - | - | - | 1,480 | 7,256 | - | 8,740
Net Tonnage Mined | k tonnes | 12,713 | 39,208 | 52,375 | 88,528 | 84,906 | 107,074 | 73,470 | 61,198 | 52,173 | 56,575 | 59,053 | 9,220 | 1,171 | - | - | - | 697,664
MINING COSTS |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Mine General Service | K US$ | $ 1,169 | $ 1,395 | $ 1,395 | $ 1,399 | $ 1,395 | $ 1,395 | $ 1,395 | $ 1,399 | $ 1,395 | $ 1,395 | $ 1,395 | $ 1,399 | $ 1,395 | $ 254 | $ 763 | $ - | $ 18,941
Mine Maintenance | K US$ | $ 1,764 | $ 1,770 | $ 1,770 | $ 2,361 | $ 2,354 | $ 2,354 | $ 2,354 | $ 2,361 | $ 2,354 | $ 2,354 | $ 2,354 | $ 2,361 | $ 2,354 | $ 461 | $ 1,079 | $ - | $ 30,406
Engineering | K US$ | $ 815 | $ 1,020 | $ 1,020 | $ 1,023 | $ 1,020 | $ 1,020 | $ 1,020 | $ 1,023 | $ 1,020 | $ 1,020 | $ 1,020 | $ 1,023 | $ 395 | $ 48 | $ 144 | $ - | $ 12,629
Geology | K US$ | $ 688 | $ 700 | $ 700 | $ 702 | $ 700 | $ 700 | $ 700 | $ 702 | $ 700 | $ 700 | $ 700 | $ 523 | $ 522 | $ 13 | $ 38 | $ - | $ 8,785
Drilling | K US$ | $ 6,923 | $ 9,968 | $ 14,753 | $ 20,243 | $ 21,627 | $ 21,600 | $ 15,158 | $ 13,315 | $ 13,129 | $ 13,954 | $ 15,764 | $ 1,112 | $ - | $ - | $ - | $ - | $ 167,548
Blasting | K US$ | $ 10,954 | $ 15,695 | $ 22,372 | $ 34,619 | $ 34,664 | $ 38,593 | $ 27,306 | $ 23,361 | $ 22,052 | $ 22,284 | $ 23,036 | $ 1,818 | $ 25 | $ - | $ - | $ - | $ 276,778
Loading | K US$ | $ 4,980 | $ 9,305 | $ 11,534 | $ 17,132 | $ 15,322 | $ 20,495 | $ 15,769 | $ 12,518 | $ 10,476 | $ 9,546 | $ 10,525 | $ 6,097 | $ 4,827 | $ 2,220 | $ 5,082 | $ - | $ 155,830
Hauling | K US$ | $ 15,056 | $ 27,143 | $ 44,309 | $ 62,064 | $ 75,651 | $ 75,530 | $ 74,425 | $ 73,605 | $ 74,240 | $ 74,035 | $ 74,854 | $ 13,991 | $ 9,415 | $ 3,809 | $ 13,139 | $ - | $ 711,270
Mine Support | K US$ | $ 4,075 | $ 5,107 | $ 5,476 | $ 5,544 | $ 5,530 | $ 6,390 | $ 10,240 | $ 9,482 | $ 9,325 | $ 8,731 | $ 8,749 | $ 6,304 | $ 5,730 | $ 453 | $ 1,358 | $ - | $ 92,491
Total Mine Cost | K US$ | $ 46,425 | $ 72,103 | $ 103,330 | $ 145,085 | $ 158,264 | $ 168,079 | $ 148,368 | $ 137,764 | $ 134,691 | $ 134,019 | $ 138,397 | $ 34,627 | $ 24,664 | $ 7,257 | $ 21,603 | $ - | $ 1,474,678
MINE COST PER TONNE MINED |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Mine General Service | $/t | $ 0.04 | $ 0.03 | $ 0.02 | $ 0.01 | $ 0.02 | $ 0.01 | $ 0.02 | $ 0.02 | $ 0.02 | $ 0.02 | $ 0.03 | $ 0.48 | $ - | $ - | $ - | $ - | $ 0.03
Mine Maintenance | $/t | $ 0.06 | $ 0.04 | $ 0.03 | $ 0.03 | $ 0.03 | $ 0.02 | $ 0.03 | $ 0.04 | $ 0.04 | $ 0.04 | $ 0.04 | $ 0.80 | $ - | $ - | $ - | $ - | $ 0.04
Engineering | $/t | $ 0.03 | $ 0.03 | $ 0.02 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.02 | $ 0.02 | $ 0.02 | $ 0.02 | $ 0.35 | $ - | $ - | $ - | $ - | $ 0.02
Geology | $/t | $ 0.03 | $ 0.02 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.18 | $ - | $ - | $ - | $ - | $ 0.01
Drilling | $/t | $ 0.25 | $ 0.25 | $ 0.26 | $ 0.21 | $ 0.24 | $ 0.20 | $ 0.19 | $ 0.21 | $ 0.22 | $ 0.25 | $ 0.30 | $ 0.38 | $ - | $ - | $ - | $ - | $ 0.23
Blasting | $/t | $ 0.40 | $ 0.39 | $ 0.40 | $ 0.37 | $ 0.39 | $ 0.35 | $ 0.35 | $ 0.36 | $ 0.38 | $ 0.39 | $ 0.44 | $ 0.62 | $ - | $ - | $ - | $ - | $ 0.38
Loading | $/t | $ 0.18 | $ 0.23 | $ 0.20 | $ 0.18 | $ 0.17 | $ 0.19 | $ 0.20 | $ 0.19 | $ 0.18 | $ 0.17 | $ 0.20 | $ 2.07 | $ - | $ - | $ - | $ - | $ 0.21
Hauling | $/t | $ 0.55 | $ 0.67 | $ 0.79 | $ 0.66 | $ 0.85 | $ 0.69 | $ 0.96 | $ 1.14 | $ 1.26 | $ 1.31 | $ 1.41 | $ 4.76 | $ - | $ - | $ - | $ - | $ 0.97
Mine Support | $/t | $ 0.15 | $ 0.13 | $ 0.10 | $ 0.06 | $ 0.06 | $ 0.06 | $ 0.13 | $ 0.15 | $ 0.16 | $ 0.15 | $ 0.17 | $ 2.14 | $ - | $ - | $ - | $ - | $ 0.13
Total Mine Cost | $/t | $ 1.70 | $ 1.78 | $ 1.83 | $ 1.54 | $ 1.78 | $ 1.53 | $ 1.91 | $ 2.14 | $ 2.29 | $ 2.37 | $ 2.61 | $ 11.77 | $ - | $ - | $ - | $ - | $ 2.02

MINE COST PER TONNE MOVED |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Mine General Service | US$ | $ 0.04 | $ 0.03 | $ 0.02 | $ 0.01 | $ 0.02 | $ 0.01 | $ 0.02 | $ 0.02 | $ 0.02 | $ 0.02 | $ 0.02 | $ 0.06 | $ 0.07 | $ 0.03 | $ 0.03 | $ - | $ 0.02
Mine Maintenance | US$ | $ 0.06 | $ 0.04 | $ 0.03 | $ 0.02 | $ 0.03 | $ 0.02 | $ 0.03 | $ 0.03 | $ 0.04 | $ 0.04 | $ 0.04 | $ 0.10 | $ 0.12 | $ 0.05 | $ 0.04 | $ - | $ 0.04
Engineering | US$ | $ 0.03 | $ 0.02 | $ 0.02 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.02 | $ 0.02 | $ 0.02 | $ 0.04 | $ 0.02 | $ 0.01 | $ 0.01 | $ - | $ 0.01
Geology | US$ | $ 0.02 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.01 | $ 0.02 | $ 0.03 | $ 0.00 | $ 0.00 | $ - | $ 0.01
Drilling | US$ | $ 0.25 | $ 0.21 | $ 0.25 | $ 0.21 | $ 0.24 | $ 0.18 | $ 0.17 | $ 0.19 | $ 0.21 | $ 0.25 | $ 0.27 | $ 0.05 | $ - | $ - | $ - | $ - | $ 0.19
Blasting | US$ | $ 0.39 | $ 0.33 | $ 0.38 | $ 0.35 | $ 0.38 | $ 0.33 | $ 0.31 | $ 0.33 | $ 0.36 | $ 0.39 | $ 0.39 | $ 0.08 | $ 0.00 | $ - | $ - | $ - | $ 0.32
Loading | US$ | $ 0.18 | $ 0.19 | $ 0.19 | $ 0.17 | $ 0.17 | $ 0.17 | $ 0.18 | $ 0.17 | $ 0.17 | $ 0.17 | $ 0.18 | $ 0.25 | $ 0.26 | $ 0.24 | $ 0.18 | $ - | $ 0.18
Hauling | US$ | $ 0.54 | $ 0.57 | $ 0.74 | $ 0.63 | $ 0.83 | $ 0.64 | $ 0.84 | $ 1.03 | $ 1.21 | $ 1.31 | $ 1.27 | $ 0.58 | $ 0.50 | $ 0.41 | $ 0.47 | $ - | $ 0.83
Mine Support | US$ | $ 0.15 | $ 0.11 | $ 0.09 | $ 0.06 | $ 0.06 | $ 0.05 | $ 0.11 | $ 0.13 | $ 0.15 | $ 0.15 | $ 0.15 | $ 0.26 | $ 0.30 | $ 0.05 | $ 0.05 | $ - | $ 0.11
Total Mine Cost | US$ | $ 1.67 | $ 1.51 | $ 1.73 | $ 1.47 | $ 1.74 | $ 1.42 | $ 1.67 | $ 1.92 | $ 2.19 | $ 2.37 | $ 2.34 | $ 1.44 | $ 1.30 | $ 0.77 | $ 0.77 | $ - | $ 1.71

21.2.2Mine Dewatering (Tetra Tech)

Operating costs are related to power consumption of pumps; labor is excluded, as supervision of the dewatering system is planned for existing mine or environmental staff and will not require dedicated personnel. Mine dewatering estimated operating costs average US$0.011/t-milled.

21.2.3CIP Process and G&A (TTP)

Please note that this Section describes costs in Australian Dollars (AUD).

Overall the approach taken for the PFS operating cost estimate establishment was to perform the estimates at a FS level of detail, leading to a higher than usual level of detail presented for the PFS. This approach was deliberated adopted to minimize rework during the FS stage, with additional information expected to be limited to the use of improved accuracy quotes for the FS cost estimate.

Final plant operating cost estimates issued for the PFS were AUD166 million per year, giving a cost of AUD9.38/t treated as shown in Table 21‑17.

21.2.3.1Cost Distribution

The distribution of operating costs was not unexpected for large scale gold operations, with the five main operating cost expenditures in descending order being:

Reagents and Consumables;

Power;

Labor;

Maintenance; and

G&A.

Items of expenditure higher than normally expected for gold mining operations related specifically to:

Ore hardness, and included consumables (mill balls) and power consumption; and

High volume / low grade ore treatment schedule and related predominantly to reagents.

Table 21‑17: Estimated Plant Operating Costs (@ Steady State) (AUD)

Cost Center | Operating Cost |
AUD/ a | AUD/ t | AUD/oz
LABOR |  |  |  |
Total | 28,210,000 | 1.59 | 68.42 | 16.9%
TRANSPORT & ACCOMMODATION |  |  |  |
Total | 1,780,000 | 0.10 | 4.32 | 1.1%
POWER |  |  |  |
Processing Plant | 44,610,000 | 2.51 |  |
Miscellaneous | 550,000 | 0.03 |  |
Bores | 0 |  |  |
Total | 45,160,000 | 2.54 | 109.53 | 27.1%

FUEL |  |  |  |
Bores | 0 | 0.00 |  |
Vehicles | 360,000 | 0.02 |  |
Plant Gas | 500,000 | 0.03 |  |
Total | 860,000 | 0.05 | 2.09 | 0.5%
MAINTENANCE |  |  |  |
Fixed Plant | 10,600,000 | 0.60 |  |
Mobile Equipment | 150,000 | 0.01 |  |
Total | 10,750,000 | 0.61 | 26.07 | 6.5%
REAGENTS & CONSUMABLES |  |  |  |
Reagent Price | 47,040,000 | 2.65 |  |
Annual Consumables | 29,020,000 | 1.63 |  |
Total | 76,060,000 | 4.29 | 184.47 | 45.7%
EQUIPMENT HIRE |  |  |  |
Total | 0 | 0.00 | 0.00 | 0.0%
PRODUCT TRANSPORT |  |  |  |
Total | 0 | 0.00 | 0.00 | 0.0%
CONTRACT – GENERAL EXPENSES |  |  |  |
GENERAL CONSUMABLES | 260,000 | 0.01 |  |
CONTRACT EXPENSES | 910,000 | 0.05 |  |
GENERAL EXPENSES | 2,500,000 | 0.14 |  |
MINING CONTRACT | 0 | 0.00 |  |
Total | 3,670,000 | 0.21 | 8.90 | 2.2%
TOTAL | AUD | 166,490,000 | 9.38 | 403.79 | 100%

21.2.3.2Labor

Estimated labor costs were developed by a build-up of base labor rates, on-costs and required work force numbers.

Workforce numbers were developed using a bottom-up approach by assessing requirements in each area, and in consultation with Vista personnel, adjusting for areas specific to Mt Todd requirements.

Labor rates were initially taken as the TTP standard rates (actual operating mine data from 2010), but were subsequently adjusted up by 7% in consultation with Vista. A review was conducted by recruitment consultant Michael Page, which indicated labor rates for 4 out of the 154 categories presented required an upwards adjustment.

Labor rates have since been revised, based on recently completed projects (2017) and industry consultation.

The whole site labor force was presented in the TTP operating cost analysis to ensure that there was some consistency in labor rates across the board, however mining and mining related labor costs were not included

in the TTP operating cost estimate as these costs were ultimately in the domain of the mining consultant MDA’s operating cost schedule.

Final process plant and general and administrative (G&A) labor cost estimates issued for the PFS were AUD28.21 million per year.

21.2.3.3Transport and Accommodation

ACCOMMODATION COST DEVELOPMENT

Taking on board the Vista model for labor force accommodation of a workforce self-funded housing scheme based in Katherine and Pine Creek, the requirements for on-going use of any camp post the construction period was estimated as follows:

Accommodation allowance to cover personnel recruitment, assuming a 20% turnover of the entire workforce annually, and assuming these personnel would consist of a four unit family requiring accommodation in the camp for an average of 2 months before sourcing their own accommodation. This provided an estimated requirement for 54 rooms in the camp per annum.

Accommodation for contractors flying to site, largest of which would predominantly consist of the mill reline crew. Assuming a nominal sum of 10 other contractors throughout the year, and assuming these could be staggered to require accommodation for periods other than during mill relines, gave an estimated requirement for an additional 18 rooms.

Accommodation for miscellaneous visitors, etc. where accommodation for whatever reason could not be mutually exclusive with mill relines provided a nominal requirement for 7 rooms.

For the total on-going accommodation estimate of 69 rooms per annum, a requirement for 70 rooms was anticipated.

An allowance of AUD61.45 per man per day was made for a continuation of the partial construction camp.

TRANSPORT COST DEVELOPMENT

Using the numbers developed for the accommodation requirement, flights to Darwin from Perth were estimated at 225 return flights per annum. Allowing a 42% / 17% / 42% split between Low, Shoulder and High seasons respectively, and assuming all flights were at fully flexible fares provided the basis for annual flight expenditures.

TRANSPORT AND ACCOMMODATION COSTS

Final transport and accommodation cost estimates issued for the PFS were AUD1.780 million per year.

21.2.3.4Power Requirements

Power usage was developed by a combination of methods, namely:

Significant power consuming items had power consumptions calculated from base formulae and models, and included the following items:

Primary crusher;

Secondary crushers;

Ball mills;

Secondary Mills; and

HPGR Units.

For smaller or steady state power consumers the power consumed was calculated as a factor of installed power, with the factor varying on known vendor motor oversizing propensities.

Nominal allowances were made for some areas where actual installed power was estimated based on usual loads for such duties, and included items such as the air conditioners, lighting and small power, etc.

The total estimated power consumption is 618 GWh/year.

21.2.3.5Fuel

Fuel consumption estimates were developed for each item of process plant mobile equipment, by estimating an annual operating hours and using vendor documented or estimated fuel consumptions for each equipment item.

Other plant items usually consuming fuel, namely power generation, product drying, borefield, etc. were all zero for the Mt Todd proposed operating plant.

21.2.3.6Maintenance

Maintenance costs were developed by applying factors to FIS equipment costs for each of the two OPEX cases. The TTP maintenance cost estimating methodology is consistent with that of the Australasian Institute of Mining and Metallurgy (Cost Estimation Handbook for the Australian Mining Industry, AUSIMM, 1993). TTP factors have been developed over a period of time and fall within the AUSIMM guidelines.

Large wear items (crusher wear liners, ball mill lifters / liners) were identified and listed separately in the consumables section.

An additional allowance of 2% was applied across the site equipment to allow for sustaining capital expenditure. Maintenance cost estimates issued for the PFS were AUD10.750 million per year.

21.2.3.7Reagents

Reagent costs were estimated by applying the ALS determined consumption rates with a quoted cost of delivered reagents to site.

Instances where consumption rates were altered from the original ALS testwork or previous assumptions included:

Consumption of carbon was increased from 15 g/t to 20g/t based on TTP industry experience.

Flocculant consumption was increased to 65 g/t due to the addition of a Tails Thickener and an increased Pre-Leach Thickener flocculant requirement.

Sodium Cyanide decreased to 450 g/t (leach feed) and Quick Lime increased to 1,283 g/t (leach feed) as advised by Vista and RDi Minerals.

Reagent prices were obtained from quotes from relevant suppliers. For the PFS only one vendor quote for the majority of reagents was available, with multiple additional quotes still pending.

Multiple (11) suppliers were engaged for the highest expenditure reagent (sodium cyanide), with a Chinese supplier chosen as the most cost effective supplier. Further price sourcing from overseas suppliers was on-going at the time of writing.

Transport costs of reagents to site were sourced from reagent suppliers, in addition to an independent quote from a transport agency. As the majority of the reagents supplied by Orica (with the exception of cyanide) were delivered to Darwin, transport costs from Darwin to site were required. The cheapest of the quotes for delivery from Darwin to Katherine was chosen as the cost to be used in the PFS, in this case it was from Seatram.

Reagent cost estimates issued for the PFS were AUD47.040 million per year.

21.2.3.8Consumables

Consumable costs were estimated by calculating or estimate consumable consumption rates coupled with quotes or estimates for unit prices.

Consumption of mill balls was estimated by the selected mill vendor and based on the ore abrasion index, and since this item was the largest expenditure in the consumable category three quotes were received, with the most cost effective being Shandong Humain (China)

Where possible, transport costs were sourced from suppliers, however if they were not provided costs were sourced from other quotes. The quote from Shandong Huamin only included shipping to Darwin. Transport costs from Darwin to Katherine were sourced from the Molycorp quote.

In some instances where vendor advice was not received in a timely fashion, consumable quotes were scaled from previous studies. Consumable cost estimates issued for the PFS were AUD29.020 million per year.

21.2.3.9Equipment Hire

The Vista requirement to minimize upfront capital costs was used as the basis to initially assume all process plant mobile equipment, all process plant light vehicles and general site vehicles (ambulance, bus, coaches, etc.) would be hired or leased rather than purchased outright.

The overall cost effectiveness of the lease decision was further analyzed with the ultimate decision to purchase the vehicles outright. Consequently the equipment hire operating costs reverted to zero, with the purchase costs then added to the capital costs. With all plant vehicles then treated as fully owned, an allowance was added for vehicle maintenance.

21.2.3.10Contract/General Expenses

TTP standard factors were used for general expenses and general consumables, some items of which are a standard allowance and others which are linked to site personnel numbers (clothing, medical supplies, etc.).

General expenses and consumables allowed for included:

General Consumables; Office and General Supplies, Tools and Equipment, Communications Maintenance Materials, Sampling and Analysis Consumables

Contract Expenses; Environmental Monitoring Costs, Contracting Electrical Expenses

General Expenses; Emergency Supply, Personnel Recruitment, Legal/Compliance, Office Communications, Safety Supplies

TTP’s standard allowances were included for contract expenses, with the adjustments specific for Mt Todd including:

Additional allowances for environmental monitoring costs as advised by Vista

Additional contract electrical costs to allow for the complexity of interaction and maintaining dual source High Voltage power supplies

General / Contract Expenses in addition to General Consumables cost estimates issued for the PFS were $3.670 million per year.

21.2.4Power Plant (POWER Engineers)

Please note that this Section describes costs in Australian Dollars (AUD).

21.2.4.1Fuel Costs

Fuel cost analysis is based on baseload operation of the power station calculated using generating equipment Higher Heating Value (HHV) heat rates operating at the manufacturer standard conditions (37°C and 38% relative humidity).

Table 21‑18 lists the fuel gas requirements for the Jenbacher J920 reciprocating engines.

Table 21‑18: Estimated Fuel Cost Summary (AUD)

Description | Base Case GE Jenbacher J920 (Seven) (AUD)
Gross Output (kW) | 72,709
Net Output (kW) | 71,255
Plant Required Average Output (kW) | 69,664
Auxiliary Loads (kW) | 1,454
Availability | 99.9%
HHV Net Heat Rate (kJ/kWh) | 8,961
Thermal Efficiency | 44.47%
Annual Fuel Gas Consumption (GJ/yr) | 5,253,349
Annual Fuel Cost for Base Load (AUD/yr) | $ 36,773,445
Pipeline toll fees (AUD/yr) | $ 3,152,010

More than 80% of the costs associated with the power station operation can be attributed to fuel costs with the remainder being scheduled maintenance and replacement parts. Properly maintained reciprocating gas engines can operate up to 80,000 hours (approximately 9 years) between major engine block overhauls with annual inspections and general maintenance. Annual maintenance and inspections consist of functional checks and parts replacement of the lubricating oil and filters, inlet air filters, combustion system (spark plugs), and cooling system inspection. Minor maintenance occurs at the 20,000 and 40,000 hour intervals. Minor maintenance consists of inspection and possible replacement of bearings, cylinder heads/liners, pistons, couplings, and turbocharger components. It is anticipated that engine maintenance will be sequenced such that only a single engine is down at any time for maintenance to avoid plant disruption.

Operating costs include a minimal staff dedicated to the power station and will be based in the Administration Building. Personnel staff will consist of three swing shifts, at two operators per 12 hour shift, with a mechanic and instrumentation/electrical technician on one shift per day. One of the day shift operators can also serve as the control room manager. Labor rates are provided by Vista and include 27% salary on-costs. Plant personnel costs are shown in Table 21‑19.

Table 21‑19: Estimated Personnel Costs, Power Plant (AUD)

| Annual Salary (AUD) | Number on Staff
Power & Water Superintendent | $ 190,500 | 1
Power Station Operator | $ 114,300 | 5
Instrumentation/Electrical Technician | $ 152,400 | 1
Mechanic | $ 152,400 | 1
Total | - | 8

Table 21‑20 lists the annual operating costs for the GE Jenbacher J920 gas engines to generate a nominal net output of 69.6MW, as required by the Base Case for the 20 year life of the mining project. This estimate includes operating overhead personnel to meet the mine’s base power demand. Engine maintenance costs are estimated to be AUD0.0115 AUD/kWh, which are typical costs for reciprocating engine maintenance with a full maintenance contract (NREL, GRI, 2013). In addition, the engine maintenance costs include the cost associated with purchasing power from the utility grid during engine maintenance downtime. The power plant was estimated on a 20-year life to account for pre-production, production, and closure operations, which is approximately 19 years, and then rounded to 20 years of life to estimate the value for determination of sale at the planned end of mine life.

Table 21‑20: Estimated Gas Turbine Maintenance Cost Schedule – 70MW (AUD)

Year | Engine Fuel Costs (AUD) | Engine Maintenance Costs (AUD) | On-Site Personnel (AUD)
1 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
2 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
3 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
4 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
5 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
6 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
7 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
8 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
9 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
10 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
11 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
12 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
13 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
14 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800

15 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
16 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
17 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
18 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
19 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
20 | $ 36,773,445 | $ 12,780,408 | $ 1,066,800
Subtotal | $ 735,468,905 | $ 255,608,168 | $ 21,336,000
|  | TOTAL | $ 1,012,413,073

21.2.5Water Treatment Plant (Tetra Tech)

Water treatment plant operating costs averaging US$0.08/t-milled.

21.2.6Tailings Storage Facilities (Tetra Tech)

Tailings operating costs are estimated to average US$0.08/t-milled over the LoM. Tailings operating costs include shaping and compaction of the mine waste in the tailings embankments that hauled as a mining cost. Pumping and power costs for tailings facility operation are included in the Process Plant costing.

21.2.7General & Administrative

G&A is estimated to be an average of US$1.22/t-milled over the LoM.

22.0ECONOMIC ANALYSIS

The Base Case is identified as a 50,000 tpd operation, as presented in this section. Another option, defined as the Alternate Case (33,000 tpd) is presented in Section 24.7 – Alternate Case.

Project economics for the Base Case are based on inputs developed by MDA, TTP, POWER Engineers and Tetra Tech. Economic results presented in the report suggest the following conclusions, assuming a 100% equity project:

.........................................Mine Life 13 years;

.........................Pre-tax NPV5%US$1,178 million, IRR: 27.3%;

.....................After-tax NPV5%US$679 million, IRR: 20.5%;

.............Payback (After-tax)3.2 years;

.....NT Royalty Taxes Paid US$357 million;

Australian Income Taxes Paid US$466 million; and

Cash costs (including Royalty)US$645.33/oz-Au.

After-tax net present value (NPV) is US$697 million, discounted at 5%. The after-tax internal rate of return (IRR) for the Project is 20.5%.

Costs and economic results are presented in Q4 2017 U.S. dollars unless otherwise stated. No escalation has been applied to capital or operating costs. The 5% discount rate used is a gold industry norm.

Technical economic tables and figures presented in this volume require subsequent calculations to derive subtotals, totals, and weighted averages. Such calculations inherently involve a degree of rounding. Where these occur they are not considered to be material.

22.1Principal Assumptions

Parameters used in the analysis are shown in Table 22‑1. These parameters are based upon current market conditions, vendor quotes, design criteria developed by Vista and their consultants, and benchmarks against similar existing projects.

The Project will commence at a production rate of 50,000 tpd. Fresh ore production will originate from the open pit mine and will be treated using conventional CIP technology. Once ore is exhausted from the pit, the reserves in the existing heap leach pad will then be processed.

Projected revenues from the sale of gold doré are based upon a market price of US$1,300/oz-Au. Vista has used indicative pricing from the Perth Mint for the sale of its product. It is too early to enter into definitive agreement with refiners as of the date of this PFS. However, refinery assumptions used in the technical economic model (TEM) are indicative of current refiner rates.

Table 22‑1: TEM Principal Assumptions

Principal Assumptions | Unit | Parameter
Pre-Production Period | Years | 2
Mine Life | Years | 13
Closure Period | Years | 4
Operating Days | Days / Year | 355
Gold Price | US$ | $ 1,300
JAAC Royalty | % | 1%
Exchange Rate | AUD:US$ | 0.8:1
Diesel Fuel | AUD/L | $ 0.760
Natural Gas | AUD/GJ | $ 7.00
Electric Power – From Grid | AUD/kWh | $ 0.300
Electric Power – From Plant | AUD/kWh | $ 0.074

Refining costs are summarized in Table 22‑2 resulting in an all in refining cost of US$3.18/Au-oz over the LoM.

Table 22‑2: Estimated Refining Costs (US$)

Cost Component | Units | Cost (US$)
Refining Fee | $/oz | 0.75
Gold Retention | % of gold sales | 0.10%
Purchase Discount-Gold | $/oz | 0.50
Assay Fee | $/oz | 95.00
Environmental Fee | $/oz | 50.00
Freight & Insurance | $/oz | 0.20

The Project will be subject to a 30% Commonwealth corporate income tax, and will also pay a profits-based royalty to the Northern Territory Government of 20%.

22.2LoM Production

Ore will be mined using open pit mining methods. Production over the LoM is summarized in Table 22‑3.

Table 22‑3: LoM Ore Production

Production | kt | g/t | Contained Au (koz)
Waste | 522,990 | - | -
Ore | 207,687 | 0.84 | 5,616
Heap Leach | 13,354 | 0.54 | 232
Total Production* | 221,041 | 0.82 | 5,848

*Total production excludes waste tonnes.

The Project has been planned as an open-pit truck and shovel operation. Open pit ore totals 208 Mt grading 0.84 g/t and contains 5.6 Moz of gold. Open pit production will have a 2.5:1 strip ratio over the 13-year LoM. Upon completion of conventional mining, the existing heap leach pad will be processed.

Ore is planned to be processed in a large comminution circuit consisting of a gyratory crusher, two cone crushers, two HPGR crushers, and two primary ball mills followed by six ISAmills for secondary grinding as discussed in Section 17.0 – Recovery Methods. Vista plans to recover gold in a conventional carbon-in-pulp (“CIP”) recovery circuit. Process recovery was determined based on ore types. Three ore types, sulfide, mixed, and oxide were identified for the open pit and will have recoveries of 86.5%, 79.5%, and 79.5%, respectively. The heap leach pad will have a recovery of 71.5% (net of solution loss).

22.3Capital Costs

LoM capital cost requirements are estimated at US$1,191 million as summarized in Table 22‑4. Initial capital of US$839 million is estimated to be required to commence operations. Sustaining capital of US$352 million is required over the LoM and accounts for capitalized stripping in the open pit, mine equipment additions and replacements, and tailings dam raises.

Table 22‑4: Estimated LoM Capital Costs (US$000s)

Area | Description | Cont. (%) | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
Estimate | Contingency | Total | Estimate | Contingency | Total | Estimate | Contingency | Total
SUMMARY
2000 | Mining | 0.3% | $144,278 | $1,204 | $145,482 | $385,207 | $237 | $385,445 | $529,485 | $1,441 | $530,927
3000 | Process Plant | 20.7% | $340,768 | $72,386 | $413,154 | $22,992 | $2,951 | $25,943 | $363,760 | $75,337 | $439,097
4000 | Project Services | 3.5% | $117,676 | $7,085 | $124,760 | $82,857 | $0 | $82,857 | $200,533 | $7,085 | $207,617
5000 | Project Infrastructure | 12.9% | $23,263 | $3,001 | $26,264 | $0 | $0 | $0 | $23,263 | $3,001 | $26,264
6000 | Permanent Accommodation | 10.0% | $68 | $7 | $75 | $0 | $0 | $0 | $68 | $7 | $75
7000 | Site Establishment & Early Works | 11.4% | $19,723 | $2,248 | $21,972 | $0 | $0 | $0 | $19,723 | $2,248 | $21,972
8000 | Management, Engineering, EPCM Svcs | 9.5% | $85,941 | $8,139 | $94,080 | $0 | $0 | $0 | $85,941 | $8,139 | $94,080
9000 | Pre-Production Costs | 17.2% | $11,127 | $1,918 | $13,045 | $0 | $0 | $0 | $11,127 | $1,918 | $13,045
10000 | Asset Sale | 0.0% | $0 | $0 | $0 | ($141,788) | $0 | ($141,788) | ($141,788) | $0 | ($141,788)
| Capital Cost | 9.1% | $742,844 | $95,988 | $838,832 | $349,269 | $3,189 | $352,458 | $1,092,113 | $99,177 | $1,191,290

22.3.12000 Mining

LoM capital cost requirements are estimated to value US$531 million with an initial cost of US$145 million as seen in Table 22‑5.

Table 22‑5: Estimated Mining Costs (US$000s)

Area | Description | Cont. (%) | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
Estimate | Contingency | Total | Estimate | Contingency | Total | Estimate | Contingency | Total
2000 | MINING
2100 | Capitalized Mine Operating | 0.0% | $46,425 | $0 | $46,425 | $86,755 | $0 | $86,755 | $133,179 | $0 | $133,179
2200 | Mine Production Equipment | 0.0% | $85,976 | $0 | $85,976 | $152,374 | $0 | $152,374 | $238,350 | $0 | $238,350
2300 | Mine Support Facilities | 8.7% | $9,849 | $856 | $10,705 | $1,350 | $117 | $1,467 | $11,199 | $973 | $12,173
2400 | Mine Support Services | 20.0% | $1,366 | $273 | $1,639 | $0 | $0 | $0 | $1,366 | $273 | $1,639
2500 | Mine Dewatering/Drainage | 14.1% | $533 | $75 | $608 | $851 | $120 | $972 | $1,384 | $195 | $1,580
2900 | Mine Closure | 0.0% | $129 | $0 | $129 | $143,877 | $0 | $143,877 | $144,007 | $0 | $144,007
| Mining | 0.3% | $144,278 | $1,204 | $145,482 | $385,207 | $237 | $385,445 | $529,485 | $1,441 | $530,927

22.3.23000 Process Plant

Estimated CIP process plant capital costs are shown in Table 22‑6. Initial capital totaling US$413 million is estimated to be required for the CIP process plant; a total capital of US$439 million is required.

Table 22‑6: Estimated CIP Process Plant Capital Costs (US$000s)

Area | Description | Cont. (%) | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
Estimate | Contingency | Total | Estimate | Contingency | Total | Estimate | Contingency | Total
3000 | PROCESS PLANT
3100 | Crushing & Screening | 13.3% | $43,295 | $5,770 | $49,065 | $3,019 | $402 | $3,421 | $46,314 | $6,172 | $52,486
3200 | Coarse Ore Stockpile, Reclaim, HPGR | 13.5% | $83,638 | $11,299 | $94,938 | $6,459 | $873 | $7,332 | $90,098 | $12,172 | $102,270
3300 | Classification & Grinding | 11.4% | $88,607 | $10,087 | $98,694 | $9,058 | $1,031 | $10,089 | $97,665 | $11,118 | $108,783
3400 | Pre-leach Thickening, Leach & CIP | 12.7% | $63,304 | $8,034 | $71,338 | $2,358 | $299 | $2,657 | $65,662 | $8,333 | $73,995
3500 | Desorption & Goldroom | 12.9% | $7,273 | $941 | $8,214 | $1,036 | $134 | $1,170 | $8,309 | $1,075 | $9,384
3600 | Detoxification & Tailings | 14.6% | $11,015 | $1,605 | $12,621 | $702 | $102 | $805 | $11,718 | $1,707 | $13,425
3700 | Reagents | 11.4% | $10,136 | $1,151 | $11,286 | $282 | $32 | $315 | $10,418 | $1,183 | $11,601
3800 | Process Plant Services | 100.0% | $33,499 | $33,499 | $66,998 | $77 | $77 | $155 | $33,576 | $33,576 | $67,153
| Process Plant | 20.7% | $340,768 | $72,386 | $413,154 | $22,992 | $2,951 | $25,943 | $363,760 | $75,337 | $439,097

22.3.34000 Project Services

Project services are estimated to have a LoM capital value US$208 million, with an initial capital value of US$125 million.

Table 22‑7: Estimated Project Services Capital Costs (US$000s)

Area | Description | Cont. (%) | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
Estimate | Contingency | Total | Estimate | Contingency | Total | Estimate | Contingency | Total
4000 | PROJECT SERVICES
4100 | Water Distribution & Water Treatment Plant | 3.8% | $19,327 | $733 | $20,060 | $0 | $0 | $0 | $19,327 | $733 | $20,060
4200 | Power Supply | 7.0% | $86,221 | $6,038 | $92,260 | $0 | $0 | $0 | $86,221 | $6,038 | $92,260
4300 | Communications | 15.9% | $736 | $117 | $853 | $0 | $0 | $0 | $736 | $117 | $853
4400 | Tailings Dams 1 & 2 | 0.0% | $8,028 | $0 | $8,028 | $82,857 | $0 | $82,857 | $90,886 | $0 | $90,886
4500 | Waste Disposal | 15.0% | $285 | $43 | $328 | $0 | $0 | $0 | $285 | $43 | $328
4600 | Plant Mobile Equipment | 5.0% | $3,078 | $154 | $3,231 | $0 | $0 | $0 | $3,078 | $154 | $3,231
4800 | Fuel Storage & Distribution (Plant) | 0.0% | $0 | $0 | $0 | $0 | $0 | $0 | $0 | $0 | $0
4900 | Project Services - Closure | 0.0% | $0 | $0 | $0 | $0 | $0 | $0 | $0 | $0 | $0
| Project Services | 3.5% | $117,676 | $7,085 | $124,760 | $82,857 | $0 | $82,857 | $200,533 | $7,085 | $207,617

22.3.45000 Project Infrastructure

The total project infrastructure is estimated to value US$26 million, which consists only of initial costs, no sustaining capital is expected. A detailed outline of costs is shown in Table 22‑8.

Table 22‑8: Estimated Project Infrastructure Capital Costs (US$000s)

Area | Description | Cont. (%) | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
Estimate | Contingency | Total | Estimate | Contingency | Total | Estimate | Contingency | Total
5000 | PROJECT INFRASTRUCTURE
5100 | Site Preparation | 13.4% | $14,159 | $1,903 | $16,062 | $0 | $0 | $0 | $14,159 | $1,903 | $16,062
5200 | Support Buildings | 10.1% | $4,639 | $470 | $5,109 | $0 | $0 | $0 | $4,639 | $470 | $5,109
5300 | Access Roads, Parking & Laydown | 10.0% | $846 | $85 | $930 | $0 | $0 | $0 | $846 | $85 | $930
5400 | Heavy Lift Cranage | 15.0% | $2,357 | $353 | $2,710 | $0 | $0 | $0 | $2,357 | $353 | $2,710
5500 | TBA | 0.0% | $0 | $0 | $0 | $0 | $0 | $0 | $0 | $0 | $0
5600 | Bulk Transport | 15.0% | $453 | $68 | $520 | $0 | $0 | $0 | $453 | $68 | $520
5700 | Power Transmission | 0.0% | $0 | $0 | $0 | $0 | $0 | $0 | $0 | $0 | $0
5800 | Communications | 15.0% | $811 | $122 | $933 | $0 | $0 | $0 | $811 | $122 | $933
| Project Infrastructure | 12.9% | $23,263 | $3,001 | $26,264 | $0 | $0 | $0 | $23,263 | $3,001 | $26,264

22.3.56000 Permanent Accommodation

Total capital for Permanent Accommodations values at US$75 million as shown in Table 22‑9.

Table 22‑9: Estimated Permanent Accommodation Costs (US$000s)

Area | Description | Cont. (%) | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
Estimate | Contingency | Total | Estimate | Contingency | Total | Estimate | Contingency | Total
6000 | PERMANENT ACCOMMODATION
6100 | Permanent Accommodation | 10.0% | $68 | $7 | $75 | $0 | $0 | $0 | $68 | $7 | $75
| Permanent Accommodation | 10.0% | $68 | $7 | $75 | $0 | $0 | $0 | $68 | $7 | $75

22.3.67000 Site Establishment & Early Works

Site Establishment and early works capital costs are estimated to total US$22 million over the LoM as shown in Table 22‑10. These costs occur in pre-production.

Table 22‑10: Estimated Site Establishment & Early Works (US$000s)

Area | Description | Cont. (%) | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
Estimate | Contingency | Total | Estimate | Contingency | Total | Estimate | Contingency | Total
7000 | SITE ESTABLISHMENT & EARLY WORKS
7300 | Construction Camp | 11.4% | $19,723 | $2,248 | $21,972 | $0 | $0 | $0 | $19,723 | $2,248 | $21,972
7400 | Dewatering | 0.0% | $0 | $0 | $0 | $0 | $0 | $0 | $0 | $0 | $0
7500 | Demolition & Removal | 0.0% | $0 | $0 | $0 | $0 | $0 | $0 | $0 | $0 | $0
| Site Establishment & Early Works | 11.4% | $19,723 | $2,248 | $21,972 | $0 | $0 | $0 | $19,723 | $2,248 | $21,972

22.3.78000 Management, Engineering, EPCM Services

Management, engineering, and EPCM services are estimated to value US$94 million. These costs are shown in Table 22‑11.

Table 22‑11: Estimated Management, Engineering, EPCM Services (US$000s)

Area | Description | Cont. (%) | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
Estimate | Contingency | Total | Estimate | Contingency | Total | Estimate | Contingency | Total
8000 | MANAGEMENT, ENGINEERING, EPCM SVCS
8100 | EPCM Services | 10.0% | $46,580 | $4,658 | $51,238 | $0 | $0 | $0 | $46,580 | $4,658 | $51,238
8200 | External Consulting & Testing | 20.0% | $960 | $192 | $1,152 | $0 | $0 | $0 | $960 | $192 | $1,152
8300 | Commissioning | 16.6% | $4,456 | $740 | $5,196 | $0 | $0 | $0 | $4,456 | $740 | $5,196
8400 | Owner's Engineering & Management | 5.8% | $29,962 | $1,752 | $31,715 | $0 | $0 | $0 | $29,962 | $1,752 | $31,715
8800 | License, fees & Legal Services | 20.0% | $1,991 | $398 | $2,390 | $0 | $0 | $0 | $1,991 | $398 | $2,390
8900 | Project Insurance | 20.0% | $1,991 | $398 | $2,390 | $0 | $0 | $0 | $1,991 | $398 | $2,390
| Management, Engineering, EPCM Svcs | 9.5% | $85,941 | $8,139 | $94,080 | $0 | $0 | $0 | $85,941 | $8,139 | $94,080

22.3.89000 Pre-Production Costs

Pre-production capital values US$13 million as shown in Table 22‑12. This cost will occur pre-preproduction.

Table 22‑12: Estimated Pre-Production Costs (US$000s)

Area | Description | Cont. (%) | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
Estimate | Contingency | Total | Estimate | Contingency | Total | Estimate | Contingency | Total
9000 | PRE-PRODUCTION COSTS
9100 | PPD Labor | 20.0% | $996 | $199 | $1,195 | $0 | $0 | $0 | $996 | $199 | $1,195
9200 | Commissioning Expenses | 25.0% | $1,991 | $498 | $2,489 | $0 | $0 | $0 | $1,991 | $498 | $2,489
9300 | Capital Spares | 15.0% | $6,783 | $1,017 | $7,801 | $0 | $0 | $0 | $6,783 | $1,017 | $7,801
9400 | Stores & Inventory | 15.0% | $1,357 | $203 | $1,560 | $0 | $0 | $0 | $1,357 | $203 | $1,560
9500 | Working Capital & Finance | 0.0% | $0 | $0 | $0 | $0 | $0 | $0 | $0 | $0 | $0
9600 | Escalation & Foreign Currency Exchange | 0.0% | $0 | $0 | $0 | $0 | $0 | $0 | $0 | $0 | $0
| Pre-Production Costs | 17.2% | $11,127 | $1,918 | $13,045 | $0 | $0 | $0 | $11,127 | $1,918 | $13,045

22.3.91000 Asset Sale

Table 22‑13 depicts a total capital value of US$142 million.

Table 22‑13: Estimated Asset Sale (US$000s)

Area | Description | Cont. (%) | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
Estimate | Contingency | Total | Estimate | Contingency | Total | Estimate | Contingency | Total
10000 | ASSET SALE
10100 | Mine | 0.0% | $0 | $0 | $0 | ($46,592) | $0 | ($46,592) | ($46,592) | $0 | ($46,592)
10200 | Process Plant | 0.0% | $0 | $0 | $0 | ($26,001) | $0 | ($26,001) | ($26,001) | $0 | ($26,001)
10300 | Power Plant (Sold to 3rd Party) | 0.0% | $0 | $0 | $0 | ($69,195) | $0 | ($69,195) | ($69,195) | $0 | ($69,195)
| Asset Sale | 0.0% | $0 | $0 | $0 | ($141,788) | $0 | ($141,788) | ($141,788) | $0 | ($141,788)

22.4Operating Costs

Estimated LoM operating costs are summarized in Table 22‑14. The operating costs will average US$14.18/t-milled over the LoM.

Table 22‑14: Estimated LoM Operating Costs (US$)

Description | US$/t-milled
OPEN PIT MINE-OWNER |
Mine General Service | $0.07
Mine Maintenance | $0.12
Engineering | $0.05
Geology | $0.03
Drilling | $0.73
Blasting | $1.20
Loading | $0.62
Hauling | $2.89
Mine Support | $0.37
Mine Dewatering | $0.01
Open Pit Mine-Owner | $6.08
CIP PROCESS PLANT |
Labor | $0.89
3100-Crush/Screen/Stockpile | $0.19
3200-Reclaim & HPGR | $0.49
3300-Classification & Grinding | $2.66
3400-Pre-Leach,Thick/Aeration/CIP | $0.15
3500-Desorption, Gold Room | $0.02
3600-Detox & Tailings Pumping | $0.08
3700-Reagents | $2.11
3800-Plant Services | $0.04
Mining, Infrastructure & Misc | $0.05
General Consumables | $0.01
Plant Mobile Equipment | $0.01
Plant Gas Consumption | $0.02
CIP Process Plant | $6.71
Project Services | $0.16
G&A | $1.22
Operating Costs | $14.18

22.4.1Open Pit Mining

Open pit mining costs are shown in Table 22‑15. Costs will average US$1.84/t-mined (US$6.08/t-milled) over the LoM. Hauling is the highest cost item, US$0.87/t-mined (US$2.89/t-milled). Hauling costs include transport of select mine waste to the TSF for embankment construction. Note also that unit costs per tonne milled include 13.4 Mt of heap leach ore which is not mined.

Table 22‑15: Estimated Open Pit Operating Costs (US$)

Description | US$/t-mined | US$/t-milled | Total (US$000s)
Mine General Service | 0.02 | 0.07 | 15,771
Mine Maintenance | 0.03 | 0.12 | 25,534
Engineering | 0.01 | 0.05 | 10,826
Geology | 0.01 | 0.03 | 7,515
Drilling | 0.22 | 0.73 | 160,624
Blasting | 0.36 | 1.20 | 265,824
Loading | 0.19 | 0.62 | 136,213
Hauling | 0.87 | 2.89 | 638,298
Mine Support | 0.11 | 0.37 | 80,894
Subtotal | 1.84 | 6.07 | 1,341,499
Mine Dewatering | 0.003 | 0.011 | 2,418
Total Open Pit Mining | 1.84 | 6.08 | 1,343,917

22.4.2CIP Process Plant

CIP process plant operating costs averaging US$6.71/t-milled are shown in Table 22‑16.

Table 22‑16: Estimated CIP Process Plant Operating Costs (US$)

Description | US$/t-milled | Total (US$000s)
Labor | 0.89 | 196,733
3100 – Crush/Screen/Stockpile | 0.19 | 42,518
3200 – Reclaim & HPGR | 0.49 | 107,989
3300 – Classification & Grinding | 2.66 | 586,979
3400 – Pre-Leach, Thick/Aeration/CIP | 0.15 | 33,582
3500 – Desorption, Gold Room | 0.02 | 3,711
3600 – Detox & Tailings Pumping | 0.08 | 17,140
3700 – Reagents | 2.11 | 465,389
3800 – Plant Services | 0.04 | 9,508
Mining, Infrastructure, & Misc | 0.05 | 10,875
Generable Consumables | 0.01 | 2,694
Plant Mobile Equipment | 0.01 | 1,920
Plant Gas consumption | 0.02 | 5,222
Total CIP Process Plant | 6.71 | 1,484,260

22.4.3Water Treatment Plant

Water treatment plant operating costs averaging US$0.08/t-milled are shown in Table 22‑17.

Table 22‑17: Estimated Water Treatment Plant Operating Costs (US$)

Description | US$/t-milled | Total (US$000s)
CHEMICALS |  |
Caustic | 0.00 | 0
Chlorine | 0.00 | 0
Citric Acid | 0.00 | 0
Ferric Chloride | 0.015 | 3,210
Ferrous Sulfate | 0.000 | 0
Lime | 0.030 | 6,674
Sodium Hydrosulfate | 0.001 | 287
Sulfuric Acid | 0.005 | 1,032
POWER |  |
Electricity | 0.009 | 2,068

LABOR |  |
Operator | 0.006 | 1,229
Maintenance | 0.015 | 3,261
Total Water Treatment Plant | 0.080 | 17,761

22.4.4Tailings

Tailings will average US$0.08/t-milled over the LoM as shown in Table 22‑18. Tailings operating costs include shaping and compaction of the mine waste in the tailings embankments that hauled as a mining cost. Pumping and power costs for tailings facility operation are included in the Process Plant costing.

Table 22‑18: Estimated Tailings Operating Costs (US$)

| US$/t-milled | Total (US$000s)
Labor | 0.039 | 8,647
Equipment | 0.044 | 9,762
Total Tailings | 0.083 | 18,409

22.4.5General & Administrative

G&A will average US$1.22/t-milled over the LoM as shown in Table 22‑19.

Table 22‑19: Estimated G&A Operating Costs (US$)

| US$/t-milled | Total (US$000s)
Labor, G&A | 0.536 | 118,435
Expenses | 0.160 | 35,466
Transport & Accommodation | 0.084 | 18,471
Fleet Vehicles | 0.016 | 3,576
Corporate Overhead | 0.427 | 94,375
Total G&A | 1.223 | 270,324

22.4.6JAAC Royalty

JAAC Royalty costs averaging US$0.291/t-milled are shown in Table 22‑20.

Table 22‑20: Estimated JAAC Royalty Costs (US$)

| US$/t-milled | Total (US$000s)
JAAC Royalty | 0.291 | 64,425
Total JAAC Royalty | 0.291 | 64,425

22.4.7Refining Costs

Refining costs averaging US$0.071/t-milled are shown in Table 22‑21.

Table 22‑21: Estimated Refining Costs (US$)

| US$/t-milled | Total (US$000s)
Refining Fee | 0.017 | 3,717
Golden Retention | 0.029 | 6,442
Purchase Discount-Gold | 0.011 | 2,478
Assay Fee | 0.003 | 771
Environmental Fee | 0.006 | 1,360
Freight & Insurance | 0.005 | 995
Total Refinery Costs | 0.071 | 15,763

22.4.8Operating Cost Inputs

Inputs used to estimate operating costs are summarized in this section.

22.4.8.1Labor

The labor breakdown shown in Table 22‑22 represents the personnel contingent at steady state operations. Labor rates are fully burdened, are presented in Australian Dollars, and are based upon recent Australian labor rate surveys provided by Vista Gold. Additionally, matrix showing salaries at levels by position is provided in Table 22‑23.

Table 22‑22: Estimated Labor Rates & Costs (AUD)

| Salary AUD | Salary On-Costs % | AUD | Number of Employees per Shift | Shift Codes | Total Employees Required | Annual Labor Costs Total AUD/Annum | Total Employees Required
MANAGEMENT
Resident Manager | $ 330,000 | 27.0% | $ 89,100 | 1 | DP | 1 | $ 419,100 | 1
Mining Manager | $ 240,000 | 27.0% | $ 64,800 | 1 | DP | 1 | $ 304,800 | 1
Processing Manager | $ 240,000 | 27.0% | $ 64,800 | 1 | DP | 1 | $ 304,800 | 1
Admin Manager | $ 200,000 | 27.0% | $ 54,000 | 1 | DP | 1 | $ 254,000 | 1
OHS Manager | $ 175,000 | 27.0% | $ 47,250 | 1 | DP | 1 | $ 222,250 | 1
NPI Manager | $ 240,000 | 27.0% | $ 64,800 | 1 | DP | 1 | $ 304,800 | 1
Subtotal |  |  |  | 6 |  | 6 | $ 1,809,750 | 6
ADMINISTRATION
HR Director | $ 180,000 | 27.0% | $ 48,600 | 1 | DP | 1 | $ 228,600 | 1
Recruiting Officer | $ 105,000 | 27.0% | $ 28,350 | 2 | DP | 2 | $ 266,700 | 2
Administration Secretary | $ 85,000 | 27.0% | $ 22,950 | 1 | DP | 1 | $ 107,950 | 1
Administrative Assistant | $ 80,000 | 27.0% | $ 21,600 | 1 | SW | 3 | $ 304,800 | 3
Receptionist | $ 65,000 | 27.0% | $ 17,550 | 1 | DP | 1 | $ 82,550 | 1
Indigenous Liaison Officer | $ 95,000 | 27.0% | $ 25,650 | 1 | DP | 1 | $ 120,650 | 1
Security Officer | $ 80,000 | 27.0% | $ 21,600 | 2 | SW | 6 | $ 609,600 | 6
Community Liaison Officer | $ 95,000 | 27.0% | $ 25,650 | 1 | DP | 1 | $ 120,650 | 1
Head of Security | $ 110,000 | 27.0% | $ 29,700 | 1 | DP | 1 | $ 139,700 | 1
External Affairs Director | $ 175,000 | 27.0% | $ 47,250 | 1 | DP | 1 | $ 222,250 | 1
Support Services Director | $ 175,000 | 27.0% | $ 47,250 | 1 | DP | 1 | $ 222,250 | 1
Subtotal |  |  |  | 13 |  | 19 | $2,425,700 | 19

FINANCE
Financial Controller | $ 180,000 | 27.0% | $ 48,600 | 1 | DP | 1 | $ 228,600 | 1
Senior Accountant | $ 135,000 | 27.0% | $ 36,450 | 1 | DP | 1 | $ 171,450 | 1
Accountant | $ 110,000 | 27.0% | $ 29,700 | 1 | DP | 1 | $ 139,700 | 1
Accounting Clerk | $ 75,000 | 27.0% | $ 20,250 | 1 | DW | 2 | $ 190,500 | 2
Payroll Clerk | $ 75,000 | 27.0% | $ 20,250 | 1 | DP | 1 | $ 95,250 | 1
Subtotal |  |  |  | 5 |  | 6 | $825,500 | 6
INFORMATION TECHNOLOGY
IT Supervisor | $ 110,000 | 27.0% | $ 29,700 | 1 | DP | 1 | $ 139,700 | 1
IT Technician | $ 90,000 | 27.0% | $ 24,300 | 1 | DP | 1 | $ 114,300 | 1
Database Administrator | $ 90,000 | 27.0% | $ 24,300 | 1 | DP | 1 | $ 114,300 | 1
Subtotal |  |  |  | 3 |  | 3 | $ 368,300 | 3
PROCESSING TECHNICAL
Metallurgical Superintendent | $ 175,000 | 27.0% | $ 47,250 | 1 | DP | 1 | $ 222,250 | 1
Chief Metallurgist | $ 175,000 | 27.0% | $ 47,250 | 1 | DP | 1 | $ 222,250 | 1
Plant / Production Metallurgist | $ 155,000 | 27.0% | $ 41,850 | 2 | DP | 2 | $ 393,700 | 2
Process Control Engineer | $ 135,000 | 27.0% | $ 36,450 | 1 | DP | 1 | $ 171,450 | 1
Metallurgical Clerk | $ 90,000 | 27.0% | $ 24,300 | 1 | SW | 3 | $ 342,900 | 3
Gold Room Supervisor | $ 110,000 | 27.0% | $ 29,700 | 1 | DP | 1 | $ 139,700 | 1
Refiner | $ 95,000 | 27.0% | $ 25,650 | 1 | DW | 2 | $ 241,300 | 2
Gold Room Technician | $ 90,000 | 27.0% | $ 24,300 | 1 | SW | 3 | $ 342,900 | 3
Subtotal |  |  |  | 9 |  | 14 | $2,076,450 | 14
PROCESSING PRODUCTION
Production Superintendent | $ 180,000 | 27.0% | $ 48,600 | 1 | DP | 1 | $ 228,600 | 1
General Foreman | $ 150,000 | 27.0% | $ 40,500 | 1 | DP | 1 | $ 190,500 | 1
Shift Foreman | $ 125,000 | 27.0% | $ 33,750 | 1 | SW | 3 | $ 476,250 | 3

Plant Lead Operator | $ 110,000 | 27.0% | $ 29,700 | 1 | SW | 3 | $ 419,100 | 3
Shift Operator - Crushing | $ 100,000 | 27.0% | $ 27,000 | 1 | SW | 3 | $ 381,000 | 3
Shift Operator - HPGR | $ 100,000 | 27.0% | $ 27,000 | 1 | SW | 3 | $ 381,000 | 3
Shift Operator - Mills | $ 100,000 | 27.0% | $ 27,000 | 1 | SW | 3 | $ 381,000 | 3
Shift Operator - Leach | $ 100,000 | 27.0% | $ 27,000 | 1 | SW | 3 | $ 381,000 | 3
Shift Operator - Elution | $ 100,000 | 27.0% | $ 27,000 | 1 | SW | 3 | $ 381,000 | 3
Shift Operator - Detox / Tailings | $ 100,000 | 27.0% | $ 27,000 | 1 | SW | 3 | $ 381,000 | 3
Shift Operator - Reagents | $ 100,000 | 27.0% | $ 27,000 | 2 | SW | 6 | $ 762,000 | 6
Shift Operator - CCR | $ 100,000 | 27.0% | $ 27,000 | 1 | SW | 3 | $ 381,000 | 3
Shift Operator - Tailings Dam | $ 90,000 | 27.0% | $ 24,300 | 2 | SW | 6 | $ 685,800 | 6
Shift Operator - Day Gang | $ 90,000 | 27.0% | $ 24,300 | 4 | SW | 12 | $ 1,371,600 | 12
Subtotal |  |  |  | 19 |  | 53 | $ 6,800,850 | 53
FIXED PLANT MAINTENANCE
Maintenance Superintendent | $ 175,000 | 27.0% | $ 47,250 | 1 | DP | 1 | $ 222,250 | 1
Maintenance General Foreman | $ 150,000 | 27.0% | $ 40,500 | 1 | DP | 1 | $ 190,500 | 1
Maintenance Planner | $ 155,000 | 27.0% | $ 41,850 | 1 | DP | 1 | $ 196,850 | 1
Mechanical Fitter | $ 120,000 | 27.0% | $ 32,400 | 3 | SW | 9 | $ 1,371,600 | 9
Crane Operator | $ 100,000 | 27.0% | $ 27,000 | 1 | DW | 2 | $ 254,000 | 2
Boilermaker / Welder | $ 125,000 | 27.0% | $ 33,750 | 2 | DW | 4 | $ 635,000 | 4
Pipe Fitters | $ 125,000 | 27.0% | $ 33,750 | 1 | DW | 2 | $ 317,500 | 2
Greasers | $ 90,000 | 27.0% | $ 24,300 | 1 | SW | 3 | $ 342,900 | 3
Trades Assistants | $ 85,000 | 27.0% | $ 22,950 | 1 | SW | 3 | $ 323,850 | 3
Electrical General Foreman | $ 145,000 | 27.0% | $ 39,150 | 1 | DP | 1 | $ 184,150 | 1
HV Electrical Supervisor | $ 120,000 | 27.0% | $ 32,400 | 1 | DW | 2 | $ 304,800 | 2
Electrician | $ 120,000 | 27.0% | $ 32,400 | 3 | SW | 9 | $ 1,371,600 | 9
Instrument Technician | $ 120,000 | 27.0% | $ 32,400 | 1 | SW | 3 | $ 457,200 | 3

Apprentices | $ 60,000 | 27.0% | $ 16,200 | 2 | SW | 6 | $ 457,200 | 6
Subtotal |  |  |  | 20 |  | 47 | $ 6,629,400 | 47
LABORATORY
Laboratory Supervisor | $ 125,000 | 27.0% | $ 33,750 | 1 | SW | 3 | $ 476,250 | 3
Chemist | $ 120,000 | 27.0% | $ 32,400 | 1 | SW | 3 | $ 457,200 | 3
Lab Technician | $ 100,000 | 27.0% | $ 27,000 | 2 | SW | 6 | $ 762,000 | 6
Sample Prep Technician | $ 75,000 | 27.0% | $ 20,250 | 3 | SW | 9 | $ 857,250 | 9
Subtotal |  |  |  | 7 |  | 21 | $ 2,552,700 | 21
MINING TECHNICAL
Engineering Superintendent | $ 180,000 | 27.0% | $ 48,600 | 1 | DP | 1 | $ 228,600 |
Chief Mining Engineer | $ 160,000 | 27.0% | $ 43,200 | 1 | DP | 1 | $ 203,200 | 1
Senior Mining Engineer | $ 140,000 | 27.0% | $ 37,800 | 0 | DP | 0 | $ 0 |
Mining Engineer | $ 130,000 | 27.0% | $ 35,100 | 1 | DW | 2 | $ 330,200 | 3
Senior Mine Planning Engineer | $ 130,000 | 27.0% | $ 35,100 | 1 | DP | 1 | $ 165,100 |
Mine Clerk | $ 100,000 | 27.0% | $ 27,000 | 1 | DP | 1 | $ 127,000 | 1
Subtotal |  |  |  | 4 |  | 5 | $ 825,500 | 5
MINING PRODUCTION
Operations Superintendent | $ 180,000 | 27.0% | $ 48,600 | 1 | DP | 1 | $ 228,600 |
Mine General Foreman | $ 150,000 | 27.0% | $ 40,500 | 1 | DP | 1 | $ 190,500 |
Drill and Blast Foreman | $ 125,000 | 27.0% | $ 33,750 | 1 | DW | 2 | $ 317,500 |
Drill and Blast Technician | $ 90,000 | 27.0% | $ 24,300 | 1 | DW | 2 | $ 228,600 | 2
Blasting Assistant | $ 85,000 | 27.0% | $ 22,950 | 1 | DW | 2 | $ 215,900 | 2
Loading Operator | $ 115,000 | 27.0% | $ 31,050 | 3 | SW | 9 | $ 1,314,450 | 11
Haul Truck Operator | $ 100,000 | 27.0% | $ 27,000 | 16 | SW | 48 | $ 6,096,000 | 81
Drill Operators | $ 115,000 | 27.0% | $ 31,050 | 8 | SW | 24 | $ 3,505,200 | 30
Mechanics | $ 120,000 | 27.0% | $ 32,400 | 0 | SW | 0 | $ 0 |

Welders | $ 125,000 | 27.0% | $ 33,750 | 0 | SW | 0 | $ 0 |
Servicemen | $ 80,000 | 27.0% | $ 21,600 | 0 | SW | 0 | $ 0 |
Aux Equipment Operators | $ 105,000 | 27.0% | $ 28,350 | 12 | SW | 36 | $ 4,800,600 | 17
Mine Shift Foreman | $ 125,000 | 27.0% | $ 33,750 | 1 | SW | 3 | $ 476,250 | 6
Subtotal |  |  |  | 45 |  | 128 | $ 17,373,600 | 149
MINE MAINTENANCE
Maintenance Superintendent | $ 175,000 | 27.0% | $ 47,250 | 1 | DP | 1 | $ 222,250 | 1
Maintenance General Foreman | $ 150,000 | 27.0% | $ 40,500 | 1 | SW | 3 | $ 571,500 | 3
Light Vehicle Mechanic | $ 120,000 | 27.0% | $ 32,400 | 2 | DW | 4 | $ 609,600 | 2
Tireman | $ 90,000 | 27.0% | $ 24,300 | 1 | DW | 2 | $ 228,600 | 2
Shop Laborer | $ 85,000 | 27.0% | $ 22,950 | 2 | SW | 6 | $ 647,700 | 2
Service, Fuel & Lube | $ 80,000 | 27.0% | $ 21,600 | 3 | SW | 9 | $ 914,400 | 6
Maintenance Planner | $ 125,000 | 27.0% | $ 33,750 | 1 | DP | 1 | $ 158,750 | 2
Subtotal |  |  |  | 11 |  | 26 | $ 3,352,800 | 17
SURVEYING
Chief Surveyor | $ 150,000 | 27.0% | $ 40,500 | 1 | DP | 1 | $ 190,500 |
Mine Surveyor | $ 115,000 | 27.0% | $ 31,050 | 1 | DW | 2 | $ 292,100 | 2
Surveying Helper | $ 80,000 | 27.0% | $ 21,600 | 1 | DW | 2 | $ 203,200 | 2
Subtotal |  |  |  | 3 |  | 5 | $ 685,800 | 4
GEOLOGY |  |  |  |  |  |  |  |
Geology Superintendent | $ 175,000 | 27.0% | $ 47,250 | 1 | DP | 1 | $ 222,250 | 1
Grade Control Geologist | $ 145,000 | 27.0% | $ 39,150 | 1 | DW | 2 | $ 368,300 | 2
Exploration Geologist | $ 110,000 | 27.0% | $ 29,700 | 1 | DP | 1 | $ 139,700 | 1
Resource Geologist | $ 155,000 | 27.0% | $ 41,850 | 1 | DP | 1 | $ 196,850 | 1
Pit Geology Technician | $ 90,000 | 27.0% | $ 24,300 | 3 | DP | 3 | $ 342,900 | 3
Geology Field Technician | $ 80,000 | 27.0% | $ 21,600 | 2 | DP | 2 | $ 203,200 | 2

Subtotal |  |  |  | 9 |  | 10 | $ 1,473,200 | 10
SUPPLY
Purchasing Director | $ 150,000 | 27.0% | $ 40,500 | 1 | DP | 1 | $ 190,500 | 1
Business Development Officer | $ 95,000 | 27.0% | $ 25,650 | 1 | DP | 1 | $ 120,650 | 1
Logistics Officer | $ 130,000 | 27.0% | $ 35,100 | 1 | DP | 1 | $ 165,100 | 1
Purchasing Officer | $ 95,000 | 27.0% | $ 25,650 | 1 | DP | 1 | $ 120,650 | 1
Contracts Officer | $ 110,000 | 27.0% | $ 29,700 | 1 | DP | 1 | $ 139,700 | 1
Store Person | $ 85,000 | 27.0% | $ 22,950 | 1 | SW | 3 | $ 323,850 | 3
Subtotal |  |  |  | 6 |  | 8 | $ 1,060,450 | 8
OHS & ENVIRONMENT
OHS Superintendent | $ 175,000 | 27.0% | $ 47,250 | 1 | DP | 1 | $ 222,250 | 1
Safety Officer | $ 115,000 | 27.0% | $ 31,050 | 1 | SW | 3 | $ 438,150 | 3
Paramedic / Nurse | $ 115,000 | 27.0% | $ 31,050 | 1 | SW | 3 | $ 438,150 | 3
Environmental Superintendent | $ 150,000 | 27.0% | $ 40,500 | 1 | DP | 1 | $ 190,500 | 1
Environmental Officer - Monitoring | $ 105,000 | 27.0% | $ 28,350 | 1 | DP | 1 | $ 133,350 | 1
Environmental Officer - Compliance | $ 105,000 | 27.0% | $ 28,350 | 1 | DP | 1 | $ 133,350 | 1
Subtotal |  |  |  | 6 |  | 10 | $ 1,555,750 | 10
TRAINING
Training Coordinator | $ 150,000 | 27.0% | $ 40,500 | 1 | DP | 1 | $ 190,500 | 1
Training Officer - Plant | $ 120,000 | 27.0% | $ 32,400 | 2 | DW | 4 | $ 609,600 | 4
Training Officer - Mining | $ 120,000 | 27.0% | $ 32,400 | 2 | DW | 4 | $ 609,600 | 1
Subtotal |  |  |  | 5 |  | 9 | $ 1,409,700 | 6
CAMP / MESSING
Camp Manager | $ 95,000 | 27.0% | $ 25,650 | 0 | DP | 0 | $ 0 | 0
Camp Admin | $ 70,000 | 27.0% | $ 18,900 | 0 | DW | 0 | $ 0 | 0
Cook Staff | $ 85,000 | 27.0% | $ 22,950 | 0 | SW | 0 | $ 0 | 0

Cleaning Staff | $ 85,000 | 27.0% | $ 22,950 | 0 | DW | 0 | $ 0 | 0
Camp Maintenance | $ 100,000 | 27.0% | $ 27,000 | 0 | DP | 0 | $ 0 | 0
Bus Drivers | $ 80,000 | 27.0% | $ 21,600 | 4 | SW | 12 | $ 1,219,200 | 12
Subtotal |  |  |  | 4 |  | 12 | $ 1,219,200 | 12
POWER PLANT
Power Station Operator | $ 90,000 | 27.0% | $ 24,300 | 1 | SW | 3 | $ 342,900 | 5
Electrician | $ 120,000 | 27.0% | $ 32,400 | 1 | SW | 3 | $ 457,200 | 1
Mechanic | $ 120,000 | 27.0% | $ 32,400 | 1 | SW | 3 | $ 457,200 | 1
Subtotal |  |  |  | 3 |  | 9 | $ 1,257,300 | 7
WATER TREATMENT PLANT
Power & Water Superintendent | $ 150,000 | 27.0% | $ 40,500 | 1 | DP | 1 | $ 190,500 | 1
Water Plant Operator | $ 90,000 | 27.0% | $ 24,300 | 1 | SW | 3 | $ 342,900 | 1
Water Plant Mechanic | $ 120,000 | 27.0% | $ 32,400 | 1 | SW | 3 | $ 457,200 | 2
Subtotal |  |  |  | 3 |  | 7 | $ 990,600 | 4
RECLAMATION
Dozer Operator | $ 105,000 | 27.0% | $ 28,350 | 2 | DW | 4 | $ 533,400 |
Loader Operator | $ 105,000 | 27.0% | $ 28,350 | 0.8 | DP | 0.8 | $ 106,680 |
Haul Truck Operator | $ 100,000 | 27.0% | $ 27,000 | 0.3 | DP | 0.3 | $ 38,100 |
Subtotal |  |  |  | 3 |  | 5 | $ 678,180 | 0
TAILINGS MANAGEMENT
Dozer Operator | $ 105,000 | 27.0% | $ 28,350 | 1.3 | DP | 1.3 | $ 173,355 | 6
Loader Operator | $ 105,000 | 27.0% | $ 28,350 | 0.17 | DP | 0.17 | $ 22,670 |
Haul Truck Operator | $ 100,000 | 27.0% | $ 27,000 | 0.51 | DP | 0.51 | $ 64,770 |
Crane Operator | $ 110,000 | 27.0% | $ 29,700 | 0.14 | DP | 0.14 | $ 19,558 |
Subtotal |  |  |  | 2 |  | 2 | $ 280,353 | 6

PROJECTS
Project Superintendent | $ 200,000 | 27.0% | $ 54,000 | 1 | DP | 1 | $ 254,000 | 1
Project Engineer | $ 165,000 | 27.0% | $ 44,550 | 1 | DW | 2 | $ 419,100 | 2
Civil Engineer | $ 130,000 | 27.0% | $ 35,100 | 0 | DW | 0 | $ 0 | 0
Geotechnical Engineer | $ 150,000 | 27.0% | $ 40,500 | 0 | DW | 0 | $ 0 | 0
CAD Draftsman | $ 88,000 | 27.0% | $ 23,760 | 0 | DW | 0 | $ 0 | 0
Piping Engineer | $ 105,000 | 27.0% | $ 28,350 | 0 | DW | 0 | $ 0 | 0
Document Controller | $ 75,000 | 27.0% | $ 20,250 | 0 | DW | 0 | $ 0 | 0
Construction Supervisor | $ 130,000 | 27.0% | $ 35,100 | 0 | DW | 0 | $ 0 | 0
Subtotal |  |  |  | 2 |  | 3 | $ 673,100 | 3
TOTAL ONSITE PERSONNEL |  |  |  |  | $ 56,324,183 | 408

Table 22‑23: Position & Salary Matrix (AUD)

Salary (AUD) | Mine Technical | Mine Operations | Mine Maintenance | Plant Technical | Plant Operations | Plant Maintenance | NPI | Administration
$330,000 |  |  |  |  |  |  |  | Resident Manager
$240,000 |  | Mining Manager |  |  | Processing Manager |  | NPI Manager |
$220,000 |  |  |  |  |  |  |  | Admin Manager
$200,000 |  |  |  |  |  |  | Projects Supt. |
$180,000 | Engineering Supt. | Operations Supt. |  |  | Production Supt |  |  | Financial Controller
|  |  |  |  |  |  | HR Director
$175,000 | Geology Supt. |  | Maintenance Supt | Metallurgy Supt |  | Maintenance Supt. |  | OHS Supt.
|  |  |  |  |  |  | External Affairs Director
|  |  |  |  |  |  | Support Services Director
$165,000 |  |  |  |  |  |  | Project Engineer |
$160,000 | Chief Mining Engineer |  |  |  |  |  |  |
$155,000 | Resource Geologist |  |  | Chief Metallurgist |  | Maintenance Planner |  |
$150,000 | Chief Surveyor | Mine General Foreman |  | Metallurgist | General Foreman | Mechanical General Foreman | Power & Water Supt | Environmental Supt
|  |  |  |  |  |  | Purchasing Director
|  |  |  |  |  |  | Training Coordinator
$145,000 | Ore Control Geologist |  |  |  |  |  |  |
$140,000 |  |  | Maintenance General Foreman |  |  | Electrical General Foreman |  |

$135,000 |  |  |  |  |  |  |  | Sr. Accountant
$130,000 | Sr. Mine Planning Engineer |  |  |  |  |  |  | Logistics Officer
Mining Engineer |  |  |  |  |  |  |
$125,000 |  | Mine Shift Foreman | Maintenance Planner | Laboratory Supervisor | Plant Shift Foreman | Welder/Pipefitter |  |
$120,000 |  |  | Light Vehicle Mechanic | Chemist |  | High Voltage Electrician | Power Station Electrician | Training Officer - Mine Equip
| Drill & Blast Foreman |  |  |  | Electrician | Power Station Mechanic | Training Officer - Fixed Plant
|  |  |  |  | Instrumentation Tech | Water Plant Mechanic |
|  |  |  |  | Mechanical Fitter |  |
$115,000 | Mine Surveyor | Drill Operator |  |  |  |  |  | Safety Officer
| Shovel Operator |  |  |  |  |  | Paramedics
$110,000 | Exploration Geologist |  |  |  | Plant Lead Operator |  |  | Contracts Officer
|  |  |  |  |  |  | Accountant
|  |  |  | Gold Room Supervisor |  |  | IT Supervisor
$105,000 |  | Aux Equipment Operator |  |  |  |  |  | Env Officer - Monitoring
|  |  |  |  |  |  | Env. Officer - Compliance
$100,000 | Mine Clerk | Haul Truck Operator |  |  | Crushing/ Sorting Operator | Crane Operator |  | Recruiting Officer
|  |  |  | Grinding/Leach Operator |  |  | Head of Security

$95,000 |  |  |  |  | Refiner |  |  | Purchasing Officer
|  |  |  |  |  |  | Business Dev. Officer
|  |  |  |  |  |  | Community Liaison Officer
|  |  |  |  |  |  | Indigenous Liaison Officer
$90,000 | Pit Geology Technician | Drill & Blast Technician | Tireman | Metallurgy Clerk | Grinding/Leach Technician | Greaser | Power Station Operator | IT Technician
|  |  | Lab Technician | Tailings Technician |  | Water Plant Operator | Database Administrator
|  |  |  | Gold Room Technician |  |  |
|  |  |  | Plant Day Gang |  |  |
$85,000 |  | Blasting Assistant | Maintenance Shop Labor |  |  | Trades Assistant |  | Store Person
|  |  |  |  |  |  | Administrative Secretary
$80,000 | Surveyor Helper |  | Fuel & Lube Technician |  |  |  |  | Administrative Assistant
Geology Field Technician |  |  |  |  |  |  | Security Officer
|  |  |  |  |  |  | Bus Driver
$75,000 |  |  |  | Sample Prep Technician |  |  |  | Accounting Clerk
|  |  |  |  |  |  | Payroll Clerk
$65,000 |  |  |  |  |  |  |  | Receptionist

22.4.8.2Reagents

Reagent consumption rates and costs are shown in Table 22‑24. Consumption rates are based upon metallurgical testwork and prices are based on vendor quotes, including a delivery to site. Unit costs of reagents are provided in AUD.

Table 22‑24: Process Reagents (AUD)

Reagent | Consumable Rate | Unit | Unit Cost (AUD) | Unit
Quick Lime | 1,283 | g/t leach feed | $ 371 | per tonne
Sodium Cyanide | 450 | g/t leach feed | $ 2,715 | per tonne
Sodium Hydroxide | 40 | g/t ore | $ 1,271 | per tonne
Flocculant | 65 | g/t leach feed | $ 3,520 | per tonne
Sodium Metabisulphite (SMBS) | 730 | g/t leach feed | $ 628 | per tonne
Hydrochloric Acid | 81 | g/t ore | $ 646 | per tonne
Lead Nitrate | 100 | g/t ore | $ 3,471 | per tonne
Activated Carbon | 20 | g/t ore | $ 2,609 | per tonne
Borax | 150 | kg/t conc. | $ 3,286 | per tonne
Silica | 150 | kg/t conc. | $ 1,566 | per tonne
Soda Ash | 100 | kg/t conc. | $ 1,966 | per tonne
Potassium Nitrate | 30 | kg/t conc. | $ 4,066 | per tonne

22.4.8.3Consumables

Consumable consumption rates are based upon benchmark data and vendor information given the ores processed at the site. Costs for consumables are based upon vendor quotes including delivery to site. These costs are shown in Table 22‑25. Unit costs of consumables are provided in AUD.

Table 22‑25: Process Consumables (AUD)

Consumables | Consumable Rate | Unit | Unit Cost (AUD) | Unit
CRUSHING
Primary Crusher mantle | 131 | days per set | $242,587 | per mantle
Primary Crusher concaves | 272 | days per set | $232,628 | per set
Secondary Crushers Main frame Liners | 481 | days per unit | $38,600 | per unit
Secondary Crushers Bowl Liners | 61 | days per unit | $32,544 | per unit
Secondary Crusher Mantle | 61 | days per unit | $26,247 | per unit
MILLING
Mill Balls 65mm | 0.06 | kg / kWh | $1,203 | per tonne
Mill Liners | 1 | sets per annum / mill | $793,375 | per set
Secondary Grinding Media | 0.01 | kg / kWh | $5.45 | per kg

HPGR
Roller Units | 26,000 | h/set | $1,215,500 | per set
Tires | 8,000 | h/set | $1,281,500 | per set
LIME SLAKER
Mill Balls | 50 mm Lime Slaking Mill | 0.5 | kg/t lime | $1,209 | per tonne

22.4.8.4Diesel Consumption

The primary consumer of diesel is mining, which totals 487 million liters of diesel. The total project consumption of diesel is 497 million liters.

22.4.8.5Plant Power Consumption

The primary consumer of power is the process facility, which totals 7,716,280 MWh power. The total project consumption of power is 7,759,246 MWh.

22.5Economic Results

Project cost estimates and economics are prepared on an annual basis. Based upon design criteria presented in this report, the level of accuracy of the estimate is considered ±25%.

Economic results are summarized in Table 22‑26. The analysis suggests the following conclusions, assuming a 100% equity project at a gold price of US$1,300:

.........................................Mine Life 13 years;

.........................Pre-Tax NPV5%US$1,178 million, IRR: 27.3%;

.....................After-tax NPV5%US$679 million, IRR: 20.5%;

.............Payback (After-tax)3.2 years;

.....NT Royalty Taxes Paid US$356 million;

Australian Income Taxes Paid US$466 million; and

Cash costs (including Royalty)US$645.33/oz-Au.

Table 22‑26: Technical-Economic Results (US$000s)

Cash Flow Summary | LoM Cost (US$000s) | Unit Cost US$/t-milled | Unit Cost US$/oz-Au
Gold Produced (koz) | 4,956 | - | -
Gold Price (US$/oz) | 1,300 | - | -
Gold Sales | 6,442,462 | 29.15 | 1,300
Refinery Costs | (15,763) | (0.071) | (3.18)
Net Smelting Return | 6,426,699 | 29.075 | 1,297
JAAC Royalty | (64,425) | (0.291) | (13.00)
Gross Income from Mining | 6,362,274 | 28.783 | 1,284
Open Pit Mine | (1,343,917) | (6.08) | (271)

CIP Process Plant | (1,484,260) | (6.71) | (300)
Project Services | (35,169) | (0.16) | (7.10)
G&A | (270,324) | (1.22) | (54.55)
Operating Costs | (3,133,670) | (14.18) | (632.33)
Power Sales Credit | 13,715 | 0.062 | 2.77
Cash Cost of Goods Sold (COGS) | (3,135,719) | (14.48) | (645.33)
Operating Margin | 3,242,319 | 14.67 | 654.26
Mining | $530,927 |  |
Process Plant | $439,097 |  |
Project Services | $207,617 |  |
Project Infrastructure | $26,264 |  |
Permanent Accommodation | $75 |  |
Site Establishment & Early Works | $21,972 |  |
Management, Engineering, EPCM Services | $94,080 |  |
Pre-Production Costs | $13,045 |  |
Asset Sale | ($141,788) |  |
Capital Costs | $1,191,290 |  |
Pre-Tax Cash Flow | 2,051,029 |  |
NPV5% | $ 1,178,110 |  |
IRR (%) | 27.3% |  |
After-tax Cash Flow | 1,245,596 |  |
NPV5% | 679,021 |  |
IRR (%) | 20.5% |  |
After-tax Payback (years) | 3.2 |  |

Table 22‑27: Annual Cash Flow

22.5.1Taxes, Royalties

22.5.1.1Royalties

NORTHERN TERRITORY ROYALTY

Under the NT Mineral Royalty Act (November 2014) (the “MRA”), royalties are based on the net value of production from a mine, irrespective of the nature of the land holding. A flat rate of 20% is applied. No royalty is payable on the first AUD50,000 of net value.

The Northern Territory Government imposes a profit-based royalty, much like an income tax, on mine production (the “NT Royalty”). The MRA codifies the calculation of the Northern Territory Royalty; however, the payment and collection of NT Royalties is not transparent. Some mines are subject to legacy customized arrangements with the NT Government, which seem to offer relief with respect to the amount and/or timing of royalty payments; some mines pay no royalty. Such agreements are strictly confidential, not a matter of public record; and each is subject to a formal application process. The NT Royalty calculated for the Project cashflows is based on the MRA rules, together with reasonable assumptions about the nature of relief that appears to be available to new mines.

Net value of production for the purposes of calculating the royalty is based the formula:

GR – (OC + CRD + EEE + AD)

where:

GR is the gross realization from the production unit in the royalty year;

OC is the operating costs of the production unit for the royalty year;

CRD is the capital recognition deduction;

EEE is the eligible exploration expenditure, if any; and

AD is any additional deduction.

22.5.1.2Other Royalties

If the Project proves feasible and subject to several conditions, Vista has agreed to offer the JAAC the opportunity to establish a Joint Venture Company with Vista holding 90% and the JAAC holding a 10% participating interest, to finance and develop the Project.

For rent of the surface rights from the current mining licenses, including the mining license on which the Batman deposit is located, the JAAC is entitled to an annual amount equal to 1% of the value of production with a minimum annual payment of AUD50,000.

There is also a royalty of 5% of the gross value of gold or other metals commercially extracted from certain mineral concessions which are located outside the zone of mineralization currently defined as the Batman deposit (the “Denehurst royalty”).

22.5.1.3Taxes

AUSTRALIAN FEDERAL INCOME TAX

The rate of corporate income tax in Australia is currently 30%; however, The Australian Parliament is considering reductions to this rate. There is no alternative minimum tax.

Taxable income is based on assessable income less allowable deductions. Assessable income generally includes gross income from the sale of goods, the provision of services, dividends, interest, royalties and rent. Assessable income may also include capital gains after offsetting capital losses. Normal business expenses are deductible.

Tax losses may be utilized and carried forward indefinitely to offset against future assessable income provided a “continuity of ownership” (more than 50% of voting, dividend and capital rights) or a “same business” test is satisfied.

Thin capitalization provisions can limit the deductibility of interest and other “debt deductions” in certain cases. In general, a deduction will be partly disallowed if the company’s debt exceeds three times its equity.

Transfer pricing rules apply to international transactions/dealings between separate legal entities. Covered cross-border transactions include those involving tangible or intangible property, the provision of services and financing. There are several generally accepted transfer pricing methods available in Australia.

Consolidation allows wholly owned corporate groups to operate as a single entity for income tax purposes.

Australia operates a full imputation system for the avoidance of double taxation of dividends. Under this system, the payment of company tax is imputed to shareholders so that shareholders are relieved of their tax liability to the extent profits have been taxed at the corporate level. Dividends paid out of profits on which corporate tax has been paid are said to be “franked” and generally entitle shareholders to an offset for the corporate tax paid.

22.5.2Sensitivity

Project sensitivities are summarized in Table 22‑28, Table 22‑29, and Table 22‑30; sensitivities are shown graphically in Figure 22‑1. As seen, the Project is most sensitive to gold production and gold price. Sensitivity on operating and capital cost is closely matched, with the Project being only slightly more sensitive to operating costs.

Table 22‑28: Project Sensitivity

Parameter | -15% | -10% | -5% | Base | 5% | 10% | 15%
Gold Price | $301,889 | $427,750 | $553,168 | $679,021 | $804,215 | $929,133 | $1,053,636
Opex | $861,235 | $800,705 | $739,938 | $679,021 | $617,587 | $556,116 | $494,830
Capex | $841,295 | $787,177 | $733,099 | $679,021 | $624,943 | $571,244 | $517,592

Figure 22‑1: Project Sensitivity

Table 22‑29: Base Case Sensitivity to Gold Price versus Foreign Exchange Rate (US$:AUD)

| GOLD PRICE (US$/oz-Au)
$1,000 | $1,100 | $1,200 | $1,300 | $1,400 | $1,500 | $1,600
FEX Rate | IRR | NPV5 | IRR | NPV5 | IRR | NPV5 | IRR | NPV5 | IRR | NPV5 | IRR | NPV5 | IRR | NPV5
0.65 | 13.3% | $ 319 | 18.1% | $ 512 | 22.8% | $ 705 | 27.2% | $ 897 | 31.5% | $ 1,089 | 35.7% | $ 1,281 | 39.8% | $ 1,472
0.70 | 11.2% | $ 247 | 15.9% | $ 439 | 20.5% | $ 632 | 24.9% | $ 825 | 29.1% | $ 1,016 | 33.2% | $ 1,208 | 37.2% | $ 1,400
0.75 | 9.3% | $ 176 | 13.8% | $ 366 | 18.3% | $ 559 | 22.6% | $ 752 | 26.8% | $ 944 | 30.8% | $ 1,136 | 34.7% | $ 1,328
0.80 | 7.4% | $ 103 | 11.9% | $ 292 | 16.3% | $ 486 | 20.5% | $ 679 | 24.6% | $ 872 | 28.6% | $ 1,063 | 32.4% | $ 1,255
0.85 | 5.5% | $ 23 | 10.1% | $ 222 | 14.3% | $ 412 | 18.5% | $ 606 | 22.5% | $ 799 | 26.4% | $ 991 | 30.2% | $ 1,183
0.90 | 3.0% | $ (94) | 8.3% | $ 150 | 12.5% | $ 339 | 16.6% | $ 532 | 20.6% | $ 726 | 24.4% | $ 918 | 28.1% | $ 1,110
0.95 | 0.4% | $ (218) | 6.6% | $ 76 | 10.7% | $ 268 | 14.7% | $ 459 | 18.7% | $ 652 | 22.5% | $ 845 | 26.1% | $ 1,038
1.00 | -2.2% | $ (341) | 4.8% | $ (8) | 9.1% | $ 197 | 13.0% | $ 385 | 16.9% | $ 579 | 20.6% | $ 772 | 24.2% | $ 965

Base Case uses a foreign exchange rate of US$0.80:AUD1.00 and a gold price of $1,300/oz-Au.

Table 22‑30 Base Case Sensitivity to Gold Prices versus NPV Discount Rate

Discount Rate | GOLD PRICE (US$/oz-Au)
$ 1,000 | $ 1,100 | $ 1,200 | $ 1,300 | $ 1,400 | $ 1,500 | $ 1,600
5% | $ 103 | $ 292 | $ 486 | $ 679 | $ 872 | $ 1,063 | $ 1,255
8% | $ (23) | $ 138 | $ 299 | $ 460 | $ 619 | $ 778 | $ 937
10% | $ (85) | $ 59 | $ 203 | $ 346 | $ 488 | $ 629 | $ 770

Foreign Exchange is held constant at US$0.80:AUD1.00

23.0ADJACENT PROPERTIES

There are no adjacent properties that are considered relevant to this Technical Report.

24.0OTHER RELEVANT DATA AND INFORMATION

The Base Case is identified as a 50,000 tpd operation, as presented in this section. Another option, defined as the Alternate Case (33,000 tpd) is presented in Section 24.7 – Alternate Case.

24.1Process Plant Geotechnical

Bulk earthworks for the process plant are designed to minimize the import of fill material and excavation of rock. Where fill material is required to be imported, material from the existing RoM Pad ramp and from the existing stockpile located adjacent to the Tollis and Golf Pits will be utilized. The civil basis of design took into consideration the following geotechnical information:

Geotechnical Desktop Study Mt Todd Process Plant DFS undertaken by Coffey Geotechnics in December 2012. The study reviewed previous Soil and Rock Engineering (SRE) geotechnical data from December 1992 and April 1993 for the original Mt Todd development. The study also reviewed SRE earthworks monitoring data for construction of the original Mt Todd development. Geotechnical test pit data was also reviewed (Tetra Tech, 2013). The study focused on foundations for heavy vibrating loads including the crusher and mill as well as screening structures and ancillary plant buildings. A review of potential borrow material in close proximity to the proposed plant site suitable for structural fill and pavement construction was also included.

Foundation Recommendations report produced by Tetra Tech in April 2012. The geotechnical test pit investigation was conducted in October 2011 at the then proposed location of the process plant. A summary of the test pit investigation and preliminary recommendations for foundation design at the site were provided.

Technical Memorandum regarding “Results of Test Pit Excavation Program and Borrow Source Investigation, Mt Todd Project, Vista Gold Corporation, Northern Territory, Australia” from Tetra Tech dated 20 December 2012. A summary of the test pit results and potential borrow sources were provided.

Foundation Recommendations report produced by Tetra Tech in February 2013. The report reviewed previous SRE geotechnical data from December 1992 and April 1993 for the original Mt Todd development. The report also reviewed previous Foundation Recommendations report produced by Tetra Tech in April 2012 and the previous geotechnical test pit investigation conducted in December 2012 at the proposed location of the process plant. A summary of the test pit investigation and recommendations for foundation design at the site were provided.

Further geotechnical investigation is recommended during the design phase of the project to obtain geotechnical data in the final location of foundations for heavy vibrating loads including the crusher and mill as well as screening structures and ancillary plant buildings and also to confirm fill material and rock excavation requirements. As well as locating borrow sources that are closer to the planned operation that may reduce these costs.

24.2Water Management

This section describes the overall Project water management and infrastructure considerations.

24.2.1Site-wide Water Balance

A site-wide water balance (SWWB) was developed within the GoldSim® software platform (Version 11.1) to simulate 13 years of mine production (12 active mining years and 1 additional year processing stockpiles) at the Vista Project Site for the Base Case.

The site-wide water balance was developed to simulate site conditions in order to:

Identify water treatment plant capacity;

Determine equalization pond sizing; and

Quantify make up water requirements from the RWD for process make up water, dust control, and potable/elution needs.

24.2.1.1Site-wide Water Balance Model

WATER BALANCE MODELING

The site-wide water balance model was constructed using deterministic (known with certainty) inputs, such as pond stage-storage relationships, as well as stochastic (known, but with some uncertainty) inputs, such as rainfall. Water storage within retention ponds (RPs) was modeled using the basic formula:

Change in Storage = Inputs – Outputs

Information provided to the model and the rules by which the site features interacted are summarized below.

MODEL ELEMENTS

The site features (pits, facilities and associated RPs) represented within the model are:

Waste Rock Dump (WRD, RP1);

Low Grade Ore Stockpile (LGOS)

Low Grade Ore Stockpile Retention Pond (LGRP);

Batman Pit (BP, RP3);

Process Plant Retention Pond (PRP);

Heap Leach Pad (HLP);

Raw Water Dam (RWD);

Process Water Pond (PWP);

Water Treatment Plant (WTP);

Process Plant (PP);

Dust Control;

Tailings Storage Facility 1 (TSF 1, RP7); and

Tailings Storage Facility 2 (TSF 2).

GENERAL ASSUMPTIONS

Interaction between site features was modeled based on the following set of guidelines:

RP1, LGRP, RP3, PRP and the HLP report to the PWP which feeds the WTP;

The PWP receives water only if it is not at risk for overtopping. Given this logic, overtopping events are allowed to occur at the RPs;

Inputs to ponds included precipitation, catchment runoff (where applicable), seepage (where applicable), and groundwater inflow (where applicable);

Outputs from ponds included evaporative loss, pumping and overtopping events (uncontrolled releases);

All RPs report to the PWP which feeds directly to WTP. The PWP was sized to contain six days of WTP capacity (72,000m 3);

A process plant bleed stream of 229 m 3/hr is sent to the PWP to maintain chemistry of the process circuit;

Discharges to the Edith River were not allowed from any of the RPs.

WTP effluent was allowed to discharge to the Edith River only during the wet season;

The HLP was run through process at the end of the Life of Mine (LoM);

WRD water that reported to RP1 was not allowed to be used for dust control;

Seepage losses from the ponds are not modeled and are assumed to be zero; and

RWD is modeled as an infinite source for site water needs due to lack of information about its catchment area and stage-storage relationship.

INITIAL CONDITIONS

RP1, LGRP, RP3, PRP and the PWP were assigned water surface elevations based on outputs from the pre-production model, which concluded at the initiation of production with initial conditions based on real site water surface elevation observations from December 2016.

The dust suppression tank was assumed to be full at the initiation of production.

FLOW RATES

Process makeup water requirements throughout the LoM were 1,764 m 3/hr. This value accounts for recycle from the thickener overflow within the Process circuit.

TSF decant flows were 1,460 m 3/hr.

RWD process makeup water flows were 304 m 3/hr.

Dust suppression requirements varied between 220 and 1,153 m 3/day.

WTP capacity is 500 m 3/hr.

CLIMATOLOGICAL INPUTS

The Vista Project Site-wide water balance model was designed to reflect weather conditions as accurately as possible, given the arid tropical climate (i.e., wet, monsoon conditions with intense, short-lived events and extended hot, dry periods). Features within the climatological section of the model included:

Stochastic precipitation inputs, such that a range of likely scenarios may be determined, thus allowing the user an understanding of dry (5-percentile), typical (mean) and wet (95‑percentile) climatological effects on the site. Precipitation was entered as a statistical probability, rather than assuming rainfall on a given day and assigning a rainfall depth. The mean monthly total precipitation values (total mm per month) provided to the model are shown in Table 24‑1.

Table 24‑1: Mean Monthly Precipitation

Month | Precipitation (mm)
January | 292
February | 259
March | 193
April | 38
May | 6
June | 2
July | 1
August | 1
September | 7
October | 35
November | 111
December | 239

Synthetic data were used to extend the precipitation period of record and identify extreme rainfall events. Project Site precipitation data were collected from 1993 to present. A correlation of the site data to a nearby Katherine gage allowed the rainfall time series to be extended to that of the Katherine gage, 138 years, thus allowing determination of extreme rainfall events that may not otherwise present within a data set spanning less than 25 years.

Linking incidental rainfall and runoff within the Edith River using the Australian Water Balance Method. Catchment parameters within the model were adjusted to ensure optimal agreement between modeled runoff and observed Edith River flows. Measured precipitation, evaporation and Edith River flow data were used to calibrate this portion of the model.

The model used stochastic evaporation, based on a monthly time step and calculated by the Blaney-Criddle method. The Blaney-Criddle approach recommends against use of a time step smaller than one month.

MODEL RUN

A time step of one day was selected for the site-wide water balance model. Use of stochastic inputs allowed a “Monte Carlo” analysis to be run wherein the 13 year LoM was simulated across 1,000 realizations (or equally likely weather scenarios), each incorporating the uncertainty associated with meteorological conditions and collectively providing an envelope of expected outcomes at the site. All RPs were subjected to the stochastic weather events as described in the previous section and reported to the WTP.

24.2.1.2Results

The site-wide water balance model results show that the Batman Pit will see water storage during the wet season and later in the life of mine. This is a result of incorporating groundwater inflows into this iteration of the SWWB model. Further investigation of potential groundwater inflows will be required to validate these groundwater inflows. Optimization of process makeup water usage and onsite reuse of stormwater could reduce the amount of water reporting to the PWP, thus, allowing for faster dewatering of the Batman Pit.

The greatest amount of make-up water required from the RWD is quantified as 11,955 m 3/day. RWD requirements were found to be the most dependent upon TSF decant volumes.

RP1, LGRP, PRP, and the HLP show less than a 1% percent probability of having an overtopping over the LoM 1. LGRP storage may be optimized.

24.2.2Wet Infrastructure

Section 18.2 – Facility 4000 Project Services discusses water supply inclusive of the water treatment plan (WTP), raw water, and potable water supply. Additional information regarding regulations, design criteria and receiving water is provided herein.

24.2.2.1Water Treatment Plant

Flow to the Process Water Pond, a combination of decant return, runoff pond water, and pit dewatering discharge, is stored and pumped to the Water Treatment Plant (WTP). The maximum design capacity of the WTP 500 m³/hr. The WTP has been designed by Tetra Tech and its discharge will be returned to the Edith River for disposal, pursuant to the conditions defined by Water Discharge License 178-05. During the dry seasons, when discharges are not allowed by License, the WTP effluent will be used in the process plant for process water and around the site as dust suppression.

24.2.2.2Water Quality Standards for Waste Water Discharge

Discharges from the site are currently regulated by Waste Discharge Licence 178-05 (WDL), issued by the Northern Territory Government on January 16, 2017. The WDL is formal approval under section 74 of the Northern Territory Water Act that authorizes and regulates the release of potential contaminants to water in the Northern Territory to ensure environmental protection objectives are met. The Mt Todd Environmental Impact Statement (GHD, 2013) indicates that after the WTP is operational, the WDL will be revised to implement 95% species protection trigger values, as defined in the Australian and New Zealand Environment and Conservation Council (ANZECC) Guidelines for Fresh and Marine Water Quality (ANZECC 2000 Guidelines). This change will be reflected in a revision to WDL 178.

1 A typical value is given. Separate model runs provide a range of overtopping events, due to the stochastic nature of the model.

The purpose of the 95% species protection trigger value (TV) is to protect water quality in the Edith River downstream of the discharge from the WTP. The WTP will discharge effluent into Batman Creek, a tributary to the Edith. Concentrations of contaminants measure in the Edith River shall not exceed the TV during discharge events. For the contaminants of concern at the Project, the TVs are presented in Table 24‑2.

Table 24‑2: Site-specific Trigger Values, Edith River Downstream of WTP Discharge

Analyte | Unit | Trigger Value | Source
pH | Su | 6-8 | ANZECC 2000 Guidelines, Table 3.3.4, Lowland River value
Dissolved Oxygen | % Saturation | 85-120 | ANZECC 2000 Guidelines, Table 3.3.4, Lowland River Value
Conductivity | µS/cm | 20-250 | ANZECC 2000 Guidelines, Table 3.3.5, Upland and Lowland River Value
Magnesium | Mg/L | 2.5 | Van Dam, et al. 2010 Environ Toxicol Chem 29(2):410-421
Sulphate | Mg/L | 129 | Elphick et al. 2011 Environ Toxicol Chem 30(1):247-253
Aluminum | µg/L | 55 | ANZECC 2000 Guidelines Table 3.4.1
Cadmium | µg/L | 0.2 | ANZECC 2000 Guidelines Table 3.4.1
Cobalt | µg/L | 90 | ANZECC 2000 Guidelines p. 8.3-118
Chromium (III) | µg/L | 3.3 | ANZECC 2000 Guidelines p. 8.3-116
Chromium (VI) | µg/L | 1.0 | ANZECC 2000 Guidelines Table 3.4.1
Copper | µg/L | 1.4 | ANZECC 2000 Guidelines Table 3.4.1
Manganese | µg/L | 1900 | ANZECC 2000 Guidelines Table 3.4.1
Nickel | µg/L | 11 | ANZECC 2000 Guidelines Table 3.4.1
Lead | µg/L | 3.4 | ANZECC 2000 Guidelines Table 3.4.1
Iron | µg/L | 300 | ANZECC 2000 Guidelines p. 8.3-123
Mercury | µg/L | 0.6 | ANZECC 2000 Guidelines Table 3.4.1
Zinc | µg/L | 8.0 | ANZECC 2000 Guidelines Table 3.4.1

The TVs for magnesium and sulphate have been held over from previous work, and are not referenced in the ANZECC 2000 Guidelines.

To determine the allowable level of water quality constituents in the discharge of the WTP, a mass balance was performed on the Edith River system. Upstream water quality values at sampling location SW2 on the Edith River, flow in the river at sampling location SW4 downstream of the WTP discharge, and the maximum WTP effluent flow were used to calculate effluent limits at the WTP that would maintain the site specific trigger value at site SW4. The equation used to determine the effluent limits is:

Where:

QWTP is the WTP maximum flow rate

CWTP is the allowable concentration of a given analyte in the WTP effluent

QSW2 is the flow in the Edith River upstream of the WTP

CSW2 is the background concentration of a given analyte in the Edith River upstream of the WTP

QSW4 is the flow in the Edith River downstream of the WTP

CSW4 is the background concentration of a given analyte in the Edith River downstream of the WTP

Field data was used to determine values for QSW4, and CSW2. Table 24‑3 presents Edith River flows at field monitoring location SW4 for the wet season, when the water treatment plant will discharge to the environment.

Table 24‑3: Edith River Flow at SW4 (m3/h), February 2013 – September 2017

Month | Mean | Median | 5 th Percentile | 95 th Percentile | Maximum Day | Minimum Day
January | 91,374 | 46,807 | 8,990 | 307,729 | 740,452 | 4,942
February | 95,600 | 43,478 | 5,787 | 319,207 | 865,929 | 0
March | 43,820 | 20,662 | 5,575 | 154,660 | 523,141 | 3,064
April | 15,686 | 7,286 | 851 | 33,399 | 419,285 | 480
May | 3,121 | 1,892 | 0 | 9,912 | 12,655 | 0
June | 1,595 | 300 | 0 | 6,635 | 7,346 | 0
July | 1,024 | 0 | 0 | 4,485 | 4,915 | 0
August | 3,990 | 0 | 0 | 9,677 | 110,348 | 0
September | 996 | 0 | 0 | 5,489 | 19,784 | 0
October | 6,575 | 0 | 0 | 73,395 | 101,972 | 0
November | 4,199 | 0 | 0 | 35,611 | 83,242 | 0
December | 32,600 | 11,835 | 3,831 | 114,225 | 297,810 | 3,543

The wet season reliably extends between December and March, and the flow in the Edith River will provide a significant amount of dilution for the WTP effluent. Discharges to the environment will only occur between December and March.

Table 24‑4 provides a summary of field data showing background water quality concentrations of constituents of concern at sampling site SW2, upstream of the WTP on the Edith River. For this assessment, it was assumed that non-detectable sampling events were equal to the minimum detection limit of the analytical method.

Table 24‑4: Water Quality Data at Sampling Site SW2, Edith River Upstream of WTP Discharge,

January 2015 – April 2017

Analyte | Unit | No. of Samples | Minimum Value | Maximum Value | 5 th Percentile Value | 95 th Percentile Value | Average Value
Magnesium | Mg/L | 81 | 0.5 | 1.0 | 0.5 | 1.0 | 0.7
Sulphate | Mg/L | 139 | 1 | 19 | 1 | 1 | 1.16
Aluminum | µg/L | 139 | 30 | 940 | 69 | 622 | 247
Cadmium | µg/L | 139 | 0.1 | 0.1 | 0.1 | 0.1 | 0.1
Cobalt | µg/L | 75 | 0.5 | 1.4 | 0.5 | 1.1 | 0.70
Chromium | µg/L | 88 | 1 | 2 | 1 | 1 | 1.01
Copper | µg/L | 134 | 1 | 20 | 1 | 2 | 1.34
Manganese | µg/L | 88 | 7 | 51 | 8 | 24.3 | 14.4
Nickel | µg/L | 88 | 1 | 2 | 1 | 1 | 1.03
Lead | µg/L | 88 | 1 | 1 | 1 | 1 | 1
Iron | µg/L | 139 | 430 | 1600 | 450 | 1100 | 751
Mercury | µg/L | 88 | 0.05 | 0.05 | 0.05 | 0.05 | 0.05
Zinc | µg/L | 88 | 1 | 16 | 1 | 8.65 | 3.56

Using the TVs presented in Table 24‑2, the background flow rate presented in Table 24‑3, the background water quality in Table 24‑4, and a WTP discharge flow rate of 500 m 3/h, the mass balance was solved for the allowable discharge concentrations at the WTP. Table 24‑5 summarizes the allowable effluent concentrations and the WTP effluent goals, which are set at 80% of the allowable concentration to allow for a factor of safety.

Table 24‑5: Mt Todd WTP Effluent Goals

Analyte | Unit | CSW2 | TV | CWTP | Effluent Goal
Magnesium | Mg/L | 1 | 2.5 | 12.5 | 10
Sulphate | Mg/L | 1 | 129 | 982 | N/A
Aluminum | µg/L | 622 | 55 | 55 | 44
Cadmium | µg/L | 0.1 | 0.2 | 0.87 | 0.69
Cobalt | µg/L | 0.1 | 90 | 680 | 544
Chromium | µg/L | 1 | 1 | 1 | 0.8
Copper | µg/L | 2 | 1.4 | 1.4 | 1.12
Manganese | Mg/L | 0.024 | 1.9 | 14.4 | 11.5
Nickel | µg/L | 1 | 11 | 78 | 62.4
Lead | µg/L | 1 | 3.4 | 19 | 15.2
Iron | Mg/L | 1.1 | 0.3 | 0.3 | 0.24

Mercury | µg/L | 0.05 | 0.6 | 4 | 3.2
Zinc | µg/L | 8.65 | 8 | 8 | 6.4

The background water quality concentration at SW2 for aluminum, chromium, copper, iron, and zinc may exceed the site specific TV. In these cases, the WTP will remove the constituent to the TV prior to discharge. WTP effluent will also be used in the process plant for process water and around the site as dust suppression. It is assumed that the water quality requirements for environmental discharge will be satisfactory for these other uses as well.

INFLUENT WATER QUALITY AND TREATMENT

The geochemistry report presents expected water quality at the equalization pond upstream of the water treatment plant in the wet season and dry season for each of the 13 operating years of the mine. The geochemistry model includes inputs from various sources on the mine site, and considers any potential chemical reactions between the various inputs prior to entering the WTP. At the WTP, we are interested in treating the worst case scenario. Table 24‑6 presents the maximum value for each chemical constituent of concern, and compares it to the WTP effluent goal.

Table 24‑6: Anticipated Influent Water Quality at the WTP

Analyte | Unit | WTP Influent | Effluent Goal | % Reduction Required
Magnesium | Mg/L | 195.3 | 10 | 94.9%
Sulphate | Mg/L | 2022 | N/A | -
Aluminum | µg/L | 27,940 | 44 | 99.8%
Cadmium | µg/L | 60 | 0.69 | 98.9%
Cobalt | µg/L | 684 | 544 | 20.4%
Chromium | µg/L | 1.1 | 0.8 | 27.2%
Copper | µg/L | 4,600 | 1.12 | 99.9%
Manganese | Mg/L | 9.63 | 11.5 | 0%
Nickel | µg/L | 646 | 62.4 | 90.3%
Lead | µg/L | 26.8 | 15.2 | 43.3%
Iron | Mg/L | 0.28 | 0.24 | 14.3%
Mercury | µg/L |  | 3.2 | -
Zinc | µg/L | 12,732 | 6.4 | 99.9%

The water treatment process is designed to meet the reductions as shown in Table 24‑6.

Water to be treated at the site will be collected in the PWP. Collected wastewater will flow by gravity from the PWP to the Feed Pump Station. The pump station is adjacent to the PWP and uses concrete wet well construction. Two wet wells (for redundancy) will each house two submersible feed pumps. The Feed Pump Station pumps the collected water to the WTP building for treatment. The WTP process will consist of hydrated lime and chemical precipitation and high rate sedimentation, followed by filtration to remove remaining solids to meet effluent goals. Two identical treatment trains will provide full redundancy at the

WTP at 250 m 3/hr, with a maximum available treatment capacity at 500 m 3/hr. Expected capital costs are presented in Table 24‑7.

All prices are given in US$ unless otherwise noted. Costs in the table include the equipment cost and an installation cost of approximately 30% of the capital cost of the equipment.

Table 24‑7: Opinion of Probable Capital Costs

Parameter | Cost (US$)
Feed Pumps | 59,000
High pH Sludge Pumps | 33,000
Neutral pH Sludge Pumps | 13,000
Treated Water Pumps | 42,000
Dust Suppression Pumps | 4,000
High Density Lime Clarifier Package System | 2,334,000
NaSH Reaction Tanks 1B & 2B | 137,000
NaSH Reaction Tanks 1B & 2B Mixers | 260,000
NaSH Clarifiers 1B & 2B | 588,000
Neutral pH Reaction Tank 1C & 2C | 137,000
Neutral pH Reaction Tank 1C & 2C Mixers | 195,000
Pressure Filters | 702,000
Waste Tank Clarifiers 1 & 2 | 520,000
Treated Water Holding Tank | 169,000
Ferric Chloride Feed System | 73,000
Lime Feed System | 988,000
Lime Metering Pumps | 29,000
Polymer Feed System | 29,000
Sodium Hydrosulfide Feed System | 29,000
Sulfuric Acid Feed System | 49,000
Earthwork | 490,000
Concrete | 225,000
Pre-engineered Building | 1,400,000
Electrical and Instrumentation | 2,000,000
Piping, Pipe Supports, and Valves | 1,460,000
Equipment Shipping | 650,000
Contingency | 1,411,000
Cyanide Probes | 6,500
HCN Gas Alarms | 13,000
Total | 14,045,500

The opinion of probable operating costs consist of electricity, labor and chemical consumption. The estimated electrical use at the site is 2,254,000 kWh annually. The estimated labor use at the site includes one (1) supervisor/certified operator and two and a half (2.5) maintenance personnel.

Table 24‑8 presents the probable annual chemical consumption for the Mt Todd WTP during average flow conditions. All prices are given in US$ unless otherwise noted.

Table 24‑8: Opinion of Probable Annual Chemical Consumption

Date, Month and Season | Jan | Feb | Mar | Apr | May | Jun | Jul | Aug | Sept | Oct | Nov | Dec
Chemical | Wet | Wet | Wet | Wet | Dry | Dry | Dry | Dry | Dry | Wet | Wet | Wet
Ferric Chloride | tonne | 40 | 42 | 42 | 41 | 36 | 29 | 24 | 23 | 22 | 23 | 31 | 35
Lime | tonne | 172 | 180 | 181 | 176 | 154 | 124 | 104 | 100 | 97 | 100 | 132 | 151
Sodium Hydrosulfide | tonne | 1.7 | 1.8 | 1.8 | 1.8 | 1.5 | 1.2 | 1.0 | 1.0 | 1.0 | 1.0 | 1.3 | 1.5
Sulfuric Acid | m 3 | 12.1 | 12.6 | 12.7 | 12.3 | 10.8 | 8.7 | 7.3 | 7.0 | 6.8 | 7.0 | 9.2 | 10.5

24.2.2.3Raw Water Reservoir and Pipeline

The Raw Water Dam (RWD) is the sole source for potable and elution water, as it is the only freshwater source on site. The RWD reservoir provides storage of fresh water for use at the mine and processing facility. The reservoir is on a tributary of Horseshoe Creek, located north and east of TSF 1, and retains a reservoir storage volume of approximately 4.5 million cubic m.

The RWD reservoir provides a ready supply of fresh water for several uses. The water balance indicates that process water obtained from recycled process water and TSF decant water will need to be supplemented, particularly in the dry season. The RWD reservoir can also provide water for dust control and onsite potable water supply. Dust control will be needed during the dry season on roads and exposed soil surfaces around the project site. The reservoir generally fills in the wet season (November through April) and will be used during the dry season (May through October). It can also supply wet season fresh water, if needed.

The existing dam is a 13-m high, 114-m long, zoned-embankment dam with a low level outlet and a spillway. The outlet works are connected to the fresh water pipeline that extends to the process plant. The spillway is designed for the flood- event discharges to Horseshoe Creek.

The existing line from the RWD will need to be replaced with a 600 mm line approximately 4 km in length in order to handle the increased raw water requirements of the higher throughput.

The Raw Water Pipeline is described in Section 18.2.1.2 – Sub-Area 4120 – Raw Water.

24.2.2.4Potable Water

Potable water will be produced by a potable water treatment plant within the processing facility, and will be distributed to the process plant, mining, administration offices and laboratory facilities.

Drinking water quality guidelines that may be relevant to the Project include the Australian Drinking Water Guidelines (ADWG). These guidelines are intended to provide a framework for good management of drinking water supplies that will assure safety at point of use (NHMRC and NRMMC, 2004).

24.2.2.5Sanitary Sewer System

The sanitary sewer system will consist of gravity lines conveying the sewerage to a single sewer lift station. The lift station will then pump the sewer to the septic system for treatment. The effluent will flow by gravity to a leach field.

24.3Geochemistry

Tetra Tech was commissioned by Vista to conduct geochemical characterization studies and predictive modeling in support of the Project PFS.

Waste rock samples were selected from the three distinct rock units identified from the 18 mappable rock codes, specifically:

Greywacke;

Shale; and

Mixed greywacke/shale (interbedded).

Eighty-seven waste rock samples were subjected to acid-base accounting (ABA). Nine samples, including three samples from each of the three distinct units were selected for kinetic testing using humidity cell tests. Mineralogy by quantitative x-ray diffraction (XRD) was conducted on the nine humidity cell test samples.

The greywacke waste rock sample average nitric acid (HNO3) extractable (sulfide) sulfur content of 0.19 wt. % was comparatively low with interbedded and shale samples containing 0.51 and 0.31 wt. %, respectively. Hydrochloric acid (HCl) extractable (sulfate) sulfur was largely absent suggesting that minimal sulfide oxidation occurred prior to geochemical characterization. On average, insoluble sulfur made up approximately 30% of the sulfur distribution in the 87 samples that underwent ABA testing. The average sulfur content of the waste rock samples was  0.51 wt. % HNO3 extractable sulfide sulfur; however, the potential for acid formation remains a concern due to the limited amount of neutralization potential (NP). On average, samples contained NP  11 kg CaCO3/tonne rock. A neutralizing potential ratio (NPR) acid base accounting (ABA) screening criteria < 2 suggests that a majority of the waste rock samples are either potentially acid generating or highly likely to generate acid whereas approximately 30% of the samples were highly unlikely to generate acid. The samples contained high insoluble sulfur (> 30 wt. %) which may be from sulfidic species that are resistant to HNO3 digestion such as sphalerite (ZnS) and/or galena (PbS).

Preliminary sulfur cutoff criteria were developed based on ABA and Net Acid Generation (NAG) pH results, to assist with waste rock management and closure planning. The specific sulfur cutoff values are:

Non-PAG waste rock is defined by total sulfur content from 0.005 wt. % through 0.25 wt. %;

Waste rock with uncertain acid generation potential ranges from 0.25 wt. % through 0.4 wt. % total sulfur;

The total sulfur content of PAG waste rock is > 0.4 wt. %; and

Waste rock with > 1.5 wt. % sulfur was considered to be likely acid generating.

The cutoffs were used for geochemical modeling of the WRD seepage and pit lake wall rock runoff and can be used in combination with the total sulfur block model based on the exploration database to assist with proper routing of waste rock.

The nine waste rock samples selected for kinetic testing were subjected to humidity cell testing. Weekly leachate quality results were obtained for pH, acidity, alkalinity, electrical conductivity, and sulfate over the

entire test duration. Monthly leachate composites for dissolved constituent concentrations were also obtained over the testing period. Of the nine samples subjected to kinetic testing, a shale sample with 0.43 wt. % HNO3 extractable sulfide sulfur and low NP = 3.7 kg CaCO3/tonne rock produced acidic leachate (pH < 6) from the initiation of testing. Elevated copper, lead, nickel, and zinc levels were observed in leachate from the acid generating cell. The remaining humidity cells produced circumneutral pH values, with relatively low concentrations of metals. However, it is anticipated that given ample time these cells will likely produce acidic leachate and concomitant increased metal concentrations.

Geochemical characterization of two tailings samples was also conducted including ABA, mineralogy, water leaching, and supernatant analysis. Humidity cell testing has been initiated on one of the samples. The samples contain 1.25 wt. % and 1.13 wt. % total sulfur with net acid production potential (NAPP) and NPR values that show the tailings have potential to eventually generate acid. However, the tailings supernatant and water leach testing produced alkaline pH values. Concentrations of some metals/metalloids, major ions, and cyanide in the tailings supernatant were above ANZECC water quality guidelines, whereas levels were lower in the water leachate but some metals and metalloids and cyanide remained elevated above the guidelines. After 32 weeks, kinetic testing of one of the samples shows a neutral pH with low concentration of metals. Calculations suggest that abundant sulfide sulfur still remains, suggesting the sample may produce acidic leachate given ample time.

Predictive geochemical modeling was conducted to determine the production phase water quality of the WTP Process Water Pond. The water quality estimates were used as a basis for the WTP design and further assist with LoM site water management planning.

Inputs to the Process Water Pond included precipitation and inputs from ponds/facilities from across the site including:

RP 1 – WRD Retention Pond;

RP 2 – Low Grade Ore Stockpile Retention Pond (LGRP);

RP 3 – Batman Pit;

PWP – Process Water Ponds;

HLP – Heap Leach Pad Pond; and

RP 7 or RP 8 – the TSF 1 or TSF 2 Ponds; and

Precipitation.

Biannual water quality estimates suggest the Process Water Pond may potentially be acidic, with a majority of metal concentrations above the ANZECC water quality guidelines. Metal concentrations fluctuate depending on the relative input source proportions reporting to the Process Water Pond.

In order for Vista to re-commence mining activities, the water in RP3 must be lowered to a level below where mining is scheduled to occur. Treatment of RP3 by micronized lime has been conducted with success, with pH levels becoming circumneutral with a general decrease in metal concentrations that are sufficient for discharge under WDL 178-05 during the wet season.

24.4Surface Water Hydrology

The Project Site is drained by the perennial Edith River, located approximately 1 km south of RP 1, and also drained by several ephemeral streams. Batman Creek, which bisects the center of the site and Horseshoe Creek, is located east of the site. Both Batman and Horseshoe feed Stow Creek, which enters the Edith River at a location south of the discharge point from the Waste Rock Dump Retention Basin (RP 1).

Horseshoe Creek and Batman Creek catchments are approximately 45 and 11 km 2, respectively. The RWD was built across Horseshoe Creek immediately above the mine, forming a sub-catchment covering about 55% of the Horseshoe Creek catchment. The remainder of the Stow Creek catchment is approximately 144 km 2 and is not impacted by mining activity. Stow Creek flows for a short distance after its confluences with both Batman Creek and Horseshoe Creek, prior to joining the Edith River. The catchment area of the Edith River upstream of Stow Creek confluence is approximately 540 km 2.

Surface water at the site is well-documented and its management has been the object of study by both Vista and the NT Government in recent years. Historically, flows from the mine have exceeded the capacity of the water management system, thus allowing uncontrolled discharges to the Edith River. The effectiveness of the water management system has improved as a result of revisions to the pumping systems, installation of a stage height and telemetry station at SW4 and a flow meter on the siphon and pumping outlets from RP 1. Additionally, the NT Government completed a raise of the spillway crest and dam at RP 1 by 1.5 m.

Drainage from the Project Site enters the Edith River at two locations: discharge point for RP 1 and West Creek. The RP 1 discharge point is located 0.8 km below the Stow Creek and the Edith River confluence. West Creek joins the Edith River approximately 1.5 km below the Stow Creek and the Edith River confluence. West Creek delivers water diverted from the western side of the WRD via the Western Diversion Drain, and overflow from the RP 1 spillway. The West Creek catchment is small and it is reported that the creek only delivers mine water to the Edith River after substantial rainfall events exceed capacity at RP 1. During the wet season (approximately November to April) uncontrolled discharges to the Edith River occur from the WRD Retention Pond (RP 1), the Low Grade Ore Stockpile Retention Pond (LGRP) and the Process Plant Retention Pond (PRP) during high-rainfall events. However, for a large part of the year (approximately May to October), no runoff from the mine area enters the Edith River.

24.5Regional Groundwater Model and Mine Dewatering

The Project will enlarge the existing Batman pit significantly below the water table. After the existing pit has been emptied, the pit is expected to require additional dewatering as mining progresses. Historical data indicate that the primary driver for dewatering design will likely be runoff entering the pit from precipitation during the wet season, rather than groundwater inflow.

The following sections provide a brief summary of the applicable hydrogeologic information, historical observations, and conceptual pit inflow model. This information and surface water hydrology information provide the basis for the dewatering cost estimate. Geologic information related to the geological setting, mineralization and exploration of the project site was presented in Sections 7.0 – Geological Setting and Mineralization, 8.0 – Deposit Types, 9.0 – Exploration, and 10.0 – Drilling; the geologic information in this section is presented from a hydrogeologic perspective as it relates to groundwater flow and pit dewatering.

24.5.1Regional and Site Hydrogeology

In the Mt Todd area, bedrock occurs either at the surface or, in some valleys and streambeds, beneath a thin layer of alluvial sediment. The 1:250,000 regional geologic map of Katherine, NT 2 indicates that the formations in the vicinity of the BP are the Finniss River Group (Burrell Creek and Tollis Formations) and the Cullen Batholith (specifically the Yenberrie and Tennysons Leucogranites). The Finniss River Group consists of greywacke, siltstone, and shale, interspersed with minor volcanics. Bedding normally strikes at 325° and dips 40° to 60° to the southwest. The Finniss River Group strata have been folded about north-trending F1 fold axes. The folds have moderately west-dipping axial planes, with some sections overturned. The rocks exhibit varying degrees of contact metamorphism which increases with proximity to the intrusive units of the Cullen Batholith. In the vicinity of the Project, metamorphism is typically noted as silicified or hornfelsed material.

The existing Batman pit is located in the Burrell Creek Formation, approximately 2 km from the surface expression of the Cullen Batholith units. However, at the proposed final depth of the pit, the contact has been shown to be only a few hundred meters west of the pit. Thus, the materials encountered during drilling in the immediate vicinity of the pit are typically hornfelsed or silicified greywackes and siltstones with almost no primary porosity. East-west trending faults and joint sets and north-south trending quartz sulfide veining crosscut the bedding. The faults exhibit only minor movement.

While there is little primary porosity in the bedrock of the Mt Todd area, the weathering profile is extensive. In the late 1980s and early 1990s, when the existing Batman pit was under development, a number of production and monitoring bores were installed 3. These bores are located both near the pit and up to 4 km north and south of the pit. In addition, Vista has advanced a number of boreholes both for exploration and geotechnical evaluation. The borehole logs generally indicate that the upper 3 m are unconsolidated. Below that, weathering typically extends to approximately 30 m below ground surface (m bgs), with the degree of weathering decreasing with depth.

The Mt Todd area experiences heavy rainfall during the wet season. The average rainfall is 1,129 mm/year, but more than 80% of it falls from December through March. Thus, anecdotally, sheet flow of precipitation runoff occurs as the thin crust of soil and alluvial material reaches saturation. During heavy rain events and for some time afterward numerous ephemeral streams develop in the valleys. These subsequently stop flowing during the dry season.

The conceptual model of groundwater flow is that nearly all of the precipitation becomes runoff. Of the precipitation that does infiltrate, most flows within the upper 3 meters of unconsolidated material toward the nearest valley, where it feeds the stream system. Within the valleys, flow occurs as surface water in the streams and also within the thin layer of alluvium beneath and adjacent to the streams. Within bedrock, most water is believed to flow in the weathered profile, through fractures. The regional flow of groundwater is generally from higher to lower elevations.

2 National Geoscience Mapping Accord, Katherine (NT), Sheet SD 53-9, Second Edition, 1994.

3 Rockwater, 1994. Mt Todd Gold Mine, Bore Water Supply Expansion Programme Bore Completion Report.

24.5.2Regional Numerical Groundwater Flow Model

Tetra Tech constructed a regional numerical groundwater flow model to estimate groundwater inflows to the open pit at Mt Todd and potential impacts to regional and local water resources. The model uses the finite-difference model code MODFLOW-SURFACT, which is widely accepted and commonly used for such applications. The model is regional in scale and incorporates hydraulic properties for regional and local geologic units as derived from on-site testing, precipitation-derived recharge, natural and man-made surface hydrologic features such as ephemeral and perennial streams, the RWD, TSF, WRD, and the existing Batman pit. The proposed enlargement of the Batman pit is incorporated into predictive simulations of groundwater inflows to the pit and post-mining recovery of the groundwater system. Although preliminary calibration of the regional groundwater model has been completed, additional calibration is required and the model has not yet been finalized. Thus, only preliminary estimates of groundwater inflow to the expanded Batman pit and post-mining groundwater system recovery are currently available. The model will be finalized and used to generate updated estimates of dewatering flows and dewatering effects on the groundwater system for the feasibility study.

For this PFS, Tetra Tech developed preliminary estimates of groundwater discharge into the pit based on preliminary model output coupled with historical observations (discussed below). Preliminary estimates from the groundwater modeling conducted to date suggest that groundwater inflows should initially be approximately 3 m 3/hr and will gradually increase as the pit is enlarged and deepened, reaching a cumulative average rate of approximately 75 to 120 m 3/hr during the latter part of Phase IV. Under expected normal conditions, a portion of the groundwater inflow would be removed by evaporation from the pit walls and floor. Pit dewatering is expected to lower groundwater levels in the vicinity of the pit. The preliminary modeling suggests that dewatering-related water level declines of 1 m or more should not extend farther than approximately 300 m from the pit.

24.5.2.1Historical Observations

During the development of the existing Batman pit, very little dewatering was required. The following observations were made:

In 1994, one bore (BW-30P) was installed to provide dewatering capability if needed for the pit. This bore targeted a production zone between 36 and 50 m bgs and was expected to yield up to 600 cubic m per day (Rockwater, 1994).

Bore BW-30P may never have been used, since in 1997 a dewatering investigation indicated that the method in use was sumps and sump pumps (Dames & Moore, 1997). The geologic materials exposed in the pit were identified to have an extremely low primary permeability but slightly higher secondary permeability along fractures, bedding planes, and joints.

In December 1999 to January 2000, a geotechnical investigation described minor seepage on bedding planes and more consistent seepage in the southwest, northwest, and northeast corners of the pit (Pells Sullivan Meynink Pty Ltd., 2000). These seepages were related closely to rainfall and were greatly diminished in the dry season. However, these seepages did not appear to raise any concern at the time with respect to water removal.

The Batman pit operations were shut down in June 2000. Vista personnel visited the site in June 2006 and reported that only 1.5 m to 2 m of water was present in the bottom of the pit, despite the pit floor being approximately 90 m to 100 m below the water table near the pit. Considering that no dewatering had been done in the intervening six years, groundwater inflow is expected to be small and, therefore, a relatively minor component of dewatering.

While the groundwater inflow component is expected to be relatively minor, precipitation during the wet season has historically been significant, especially on a short-term basis. Monthly reports prior to June 2000 indicate that on several occasions large storm events generated sufficient storm-water inflow to interrupt mine operations. One event in particular resulted in the pit floor being inaccessible for approximately a month (General Gold Operations Pty Ltd (GGO), 2000). Thus, a dewatering plan will be required to ensure that surface water runoff and precipitation inflows do not significantly hamper consistent mine operation.

24.5.3Inflow Estimates

As noted above, groundwater inflow is expected to be a relatively minor component of dewatering. However, the large amount of precipitation and storm-water runoff has historically been a cause for concern. Therefore, for the PFS level dewatering design, the role of groundwater inflow has been assumed to be negligible, and storm-water runoff is the primary consideration for pit dewatering design. While negligible in terms of dewatering system design, groundwater inflows are more continuous than storm-water inflows and hence are significant relative to estimation of dewatering operating costs.

Thus, Tetra Tech based the conceptual dewatering plan on the 10-year recurrence interval, 72-hour and 100-year recurrence interval, 24-hour duration storm events. The precipitation values for those storm events, as obtained from the Bureau of Meteorology (BOM, 2017), are 3.47 mm/hr for 72 hours, which results in a total amount of 249.84 mm, and 10.7 mm/hr for 24 hours, which results in a total amount of 256.80 mm. The precipitation is assumed to fall uniformly over the pit and its catchment area. As the pit increases in size during mine development, the catchment area outside the pit would decrease until the pit comprises the entire drainage area in Phase II. Total volumes of storm water runoff inflow to the pit at the end of each phase of mine development are listed in Table 24‑9. The volumes were calculated using the SCS Curve Number method (USDA, 1996).

Table 24‑9: Catchment and Pit Areas, Inflow Volumes,

and Dewatering Times for Mine Dewatering Design

Mine Phase | Catchment Area (m 2) | Inflow Volume (m 3) | Days to dewater pit are 10 year storm event 1
Phase I | 569,000 | 134,676 | 14
Phase II | 738,787 | 176,479 | 19
Phase III | 973,336.3 | 232,508 | 25
Phase IV | 1,251,963 | 299,065 | 32

24.5.4Mine Dewatering

Dewatering of the proposed Mt Todd Mine Batman Pit is anticipated to be through passive collection of water in the pit floor sump. The sump would collect surface water, pit wall run-off and precipitation, as well as groundwater inflow. Table 24‑9 shows the days to dewater the pit after a 10-year storm event for each phase of pit development at a 500 m 3/hr pump rate. The design pumping rate does not increase throughout the LoM; instead, the time it takes to dewater the pit after a storm event increases. The dewatering system has been sized to 700 m 3/hr and reports to the PWP.

Sump water would be removed through pumping and discharge lines to the pit rim and ultimately to the PWP. Water pumped from the pit floor would first go through a pair of pumps mounted on pontoons and then through skid mounted booster pumps placed at 96-120m lifts. Lifts with booster pumps will be added in stages with increasing pit depth. Once at the surface, the water would be piped the PWP. Figure 24‑1 shows the pit floor pump, booster pumps, and pipeline conceptual design, and Figure 24‑2 shows the conceptual layout of the dewatering system. Costs for dewatering are provided in Section 22.0 – Economic Analysis.

The mine dewatering system may require modification and refinement as empirical data become available during advanced exploration and initial mine construction and operation. In particular, groundwater-related mine inflow estimates should be refined based on numerical model updates incorporating observed groundwater inflow rates to the pit and observed water level changes in groundwater monitoring bores at the site.

Figure 24‑1: Open Pit Dewatering System Conceptual Design

Figure 24‑2: Conceptual Layout of Dewatering System

24.6Project Implementation

24.6.1Project Implementation Strategy

This section outlines a high level Project Development Strategy, which will be further developed during the next study phase of the Project.

The PFS definitions of Scope, Cost and Schedule have been established on the presumption that Vista will implement the Project utilizing the Engineering, Procurement and Construction Management (EPCM) Execution model.

Vista will appoint an EPCM Contractor (Engineer) with the prerequisite capability and experience to undertake the work.

To complement the EPCM approach, Vista will adopt Design and Construct (D&C) implementation strategies, for select areas of the Project.

Properly executed, the EPCM Execution strategy will afford Vista the following benefits:

Lower Capital Cost Outcomes

Project Implementation flexibility

Fast-Track Execution opportunities

Flexible Project Funding Strategies

Optimal Project Quality Outcomes

24.6.2Project Organization

24.6.2.1EPCM Contracts

Vista’s Project Manager will direct all activities including EPCM and D&C Contractors.

For the EPCM Scope, two organization charts are developed:

EPCM Stage 1 – Design & Procure. Refer to Figure 24‑3.

EPCM Stage 2 – Construct & Commission. Refer to Figure 24‑4.

Figure 24‑3: EPCM Stage 1 – Design & Procure. Refer Diagram 1

Figure 24‑4: EPCM Stage 2 – Construct & Commission. Refer Diagram 2

24.6.2.2D&C Contracts

Two D&C contracts are proposed, specifically:

Non Process Infrastructure (NPI) – Transportable Buildings (Package C160); and

NPI – Site Erected Buildings (Package C170).

24.6.2.3EPCM Contract Scope of Services

Generally, the Engineer will perform the following tasks:

Detailed process, civil, structural, mechanical and electrical design;

Establish a document control system;

Preparation of specification documentation;

Calling and review of tenders for supply and installation of equipment;

Contract evaluation, negotiations, documentation and management;

Preparation of Purchase Orders and Contracts;

Quality audits of major contractors and manufacturers;

Construction management;

Equipment and site inspections;

Cost control, procurement, scheduling and planning, contract administration;

Regular reporting on progress against schedule and cost against budget;

Site testing and commissioning; and

Preparation and review of Operation and Maintenance manuals.

24.6.3EPCM Management

The Engineer will provide an experienced and suitably qualified Project Manager who will manage all aspects of the EPCM Contract. The EPCM Manager will be the single point of contact for the Vista Project Manager and will work closely with the Vista senior managers and other Project Managers associated with the project.

24.6.4Engineering

The Engineer will provide an experienced and suitably qualified Engineering Manager who will manage discipline based groups of Engineers and Draftsmen that will be responsible for coordination, direction, administration and completion of all detail design. Effort will be primarily aimed at optimizing design, uniformity and quality of design and monitoring of time spent against budget.

Where Engineering Design is undertaken, progress will be reported by the Engineering Manager through the EPCM Project Manager to the Vista Project Manager.

24.6.5EPCM Controls

Using the Feasibility Study report as the basis for project scope and the capital cost estimate as the control budget in the first instance, the project will be managed in accordance with the Project Schedule submitted in the Study report.

Initial activities will be directed to the awarding of Construction Contracts and/or Supply Contracts for long lead time items of plant and equipment, immediately upon Vista’s approval to proceed. The budget and schedule will be continually updated to reflect the current understanding of the project status.

The Engineer’s Project Controls group will report to the EPCM Project Manager and will have responsibility for the following activities:

Monitoring and reporting of contract package progress. This will be performed on a daily basis as necessary and reported weekly/fortnightly/monthly as required by means of the Procurement Status Report.

Definitive estimate maintenance and forecasting. Records of variations to budget and other forecast estimates of cost to completion will be updated as necessary and reported by means of the Cost Control Report Summary.

Cost control for procurement and contracting. Actual costs (invoiced and payments made), committed costs (orders placed) and estimated costs will be reported against budget using the Cost Control Report. The Trend Notice/Scope Change Notice system will be incorporated with these activities to ensure accurate forecasting.

Coordination of construction planning and scheduling. Weekly meetings of all TTP-controlled site contractors’ Project Managers will be held to coordinate changes, clashes and priorities between contractors.

Maintenance of an overall Schedule. The Schedule will be formatted using the WBS information received from all contractors and will be updated using information obtained from the various contractors and reviewed by the Project Manager on a weekly basis as a minimum.

Project reporting. A monthly project progress report will be issued including, but not limited to, the following information:

Highlights for the reporting period:

Safety, Health and Environment issues;

Overall project status;

Engineering progress;

Procurement and fabrication progress;

Construction activities;

Planned activities for the next reporting period;

Current project cost reports;

Outstanding issues, Variations, Technical Queries, etc.;

Project S-curves; and

Photographs depicting project progress.

24.6.6Procurement

24.6.6.1Procurement Strategy

The key procurement aims and objectives are to:

Achieve the project objectives of earliest possible completion, cost-effective execution, quality workmanship and high degree of safety from suppliers.

Adhere to the project plan, aims and schedule.

Ensure that commercial and schedule risks are at acceptable levels.

Provide a purchasing environment that minimizes claims and protracted disputes.

Provide a procurement arrangement that encourages suppliers to be innovative and efficient.

Carry out the procurement function for the project in an ethical and professional manner.

Key success factors are to:

Meet or exceed expectations for health and safety requirements.

Meet or better the project schedule.

Meet or better the project budget.

Meet project quality objectives.

24.6.6.2Procurement Overview

The Engineer’s Procurement Manager will report directly to the Engineer’s EPCM Project Manager but will also liaise directly with Vista’s Commercial Manager.

The Engineer’s Procurement Manager will be responsible for the preparation, all approvals and proper implementation of the Project Procurement Plan.

Prior to the project receiving all necessary approvals (Vista and Statutory), award of clearly identified and specified packages containing long lead time delivery items will only be initiated by written authorization from Vista.

The Engineer’s Procurement Manager will adhere to Vista’s procurement policy and procedures in place at the time with regard to authorization levels for capital expenditure and the requirements to obtain competitive quotations at discreet capital expenditure levels.

All packages for supply of all project related goods and services will be prepared, tendered, assessed and awarded by the Engineer’s Procurement group. All purchase orders and contracts will be prepared by the Engineer’s Procurement group but issued through Vista’s purchasing system.

Where goods and services are required from outside Australia, the Engineer’s Procurement Manager will ensure, through liaison with Vista, that sufficient forward cover on foreign exchange transactions is in place to mitigate any risk of currency fluctuation.

24.6.6.3Construction Packages

The Engineers Procurement Manager will be responsible for the development of a Construction Contracting Strategy.

A preliminary strategy is documented in the Contracting and Procurement Plan.

The following Construction packages are envisaged as a minimum:

Table 24‑10: Construction Packages

Package No. | Package Description
C015 | Demolition Works
C020 | Bulk Earthworks
C025 | Power Plant
C030 | Construction Camp
C032 | Camp Management
C035 | Tailings Dam
C040 | Roads, Drainage & Fencing
C045 | Concrete Supply
C050 | Concrete Works
C055 | Water Treatment Plant
C060 | Structural Mechanical and Piping Installation
C070 | Site Erection of Field Tankage
C080 | Electrical and Instrumentation Installation
C090 | Control System - Install & Commission
C100 | Fuel Farm
C110 | ANFO Facility
C120 | Fire Systems
C140 | Power Lines Reticulation
C160 | NPI Transportable Buildings
C170 | NPI Site Erected Buildings
C180 | Communications – Telstra Interface
C190 | Communications – Temporary

24.6.6.4Supply Packages

The Engineers Procurement Manager will be responsible for the development of an Equipment and Services Supply Contracting strategy.

The following supply packages are envisaged as a minimum:

Table 24‑11: Supply Packages

Package No. | Package Description
P001 | Ball Mills
P030 | Secondary Grinding Mills
P002 | Primary Crusher
P003 | Secondary Crushers
P004 | HPGRs
P005 | Dry Screens
P006 | Wet Screens
P007 | Slurry Pumps
P008 | Solution Pumps
P009 | Apron Feeders
P010 | Belt Feeders
P011 | Cyclone Clusters
P012 | Agitators
P013 | Thickener
P014 | Inter Tank Screens
P015 | Carbon Transfer Pumps
P016 | Gold Room
P017 | Vibrating Feeders
P018 | Container Tippers
P020 | Flocculant Mixing Package
P021 | Lime Slaker
P023 | Potable Water Plant
P024 | Mill Relining Machine
P025 | Overhead Travelling Cranes
P026 | Air Compressors, Driers & Receivers
P028A | Conveyor Drives
P028B | Conveyor Pulleys
P028C | Conveyor Idlers
P028D | Conveyor Belts & Splicing
P028E | Conveyor Skirts
P028F | Conveyor Scrapers & Ploughs
P029 | Ore Sorting
P031 | Wet Scrubber

P032 | Isolation Gates
P033 | Ventilation Fans
P034 | Screw Feeders
P035 | Rotary Valves
P036 | Filters
P038 | Hoists
P039 | Ball Charging Magnets
P040 | Tramp Magnets
P041 | Sump Pumps
P042 | Firewater System
P043 | Weightometers
P045 | Samplers
P046 | Analyzers
P047 | Rock Breaker
P048 | Blowers - Detox
P049 | Metal Detectors
P050 | FRP Tanks
P051 | Winches
P053 | Manual Valves
P054 | Laboratory Equipment
P055 | Bag Splitters
P056 | Safety Showers
P057 | Pressure Relief Valves
P058 | Pressure Regulators
P060 | Weighbridge
P101 | HV Switchgear
P102 | HV Cables
P103 | Transformers
P104 | Motor Control Centers (MCCs)
P105 | HV Variable Speed Drives
P106 | Neutral / Earth Resistors
P107 | Overhead Power Lines
P108 | Control System - Supply
P109 | Instruments
P110 | Switchrooms/MCCs
P111 | LV Variable Speed Drives
P112 | Power Factor Correction / Harmonic Filters
P113 | Control Valves
P114 | CCTV

P115 | 2 way Radios
P116 | Plant Fire Detection Systems
P117 | RMUs/Kiosk Substations
P118 | Spares
P119 | Telemetry
P120 | Emergency Power
P121 | Security
P122 | UPS
P123 | WAD Cyanide Analyzers
P124 | HCN Monitors
P125 | Data Room
P126 | Motors
P200 | Fabricated Structural Steel Work
P210 | Fabricated Platework

24.6.6.5Indirect Packages

The Engineer’s Procurement Manager, in collaboration with Vista, will establish and manage a series of Indirect Packages.

The Indirect Packages are envisaged as a minimum:

EPCM Services

Environmental Consultants

Human Resources (HR) & Industrial Relations (IR) Consultants

HSEC Consultants

Commissioning

Licenses, Fees and Legals

Project Insurances

Pre-Production Costs

Capital Spare

Stores and Inventories

Heavy Lift Cranage

24.6.6.6Expediting

The senior expeditor will plan and control expediting activities in consultation with procurement, establishing material status reports and ensuring suppliers comply with agreed delivery of drawings, data, materials and equipment. The post-award responsibility for the Supply Contract is vested with expediting; however, commercial responsibility stays with the purchasing officer. Expeditors will anticipate and act at the earliest possible stage to eliminate or reduce delays which may impact on the project schedule.

Manufacturing and delivery progress will be monitored and reported to the project via expediting status reports. Status reports will verify the milestones reported. Exceptions will be reported to management. These reports will detail actions being taken to resolve any issues causing concern.

The senior expeditor will utilize global support offices of a worldwide expediting third-party provider if necessary, to achieve the project schedule.

24.6.6.7Logistics and Transport

The Engineer will be responsible to manage the consignment of equipment and materials to the Project Site in the Northern Territory. A proven international project freight forwarding group with a global network will be appointed early in the project to provide logistics support services and to aid in the preparation of the transport and logistics plan. The focus will be on the most cost effective solution for delivery to site of equipment and materials to meet the construction schedule.

The logistics specialist will develop a transport plan to be used to manage the sea, road and airfreight costs to budget. Selected land transport subcontractors will be required to display the necessary capabilities and dedicated management that will ensure equipment is suitable and operators take every precaution to meet the project safety and quality requirements.

Transport plans will be prepared for all equipment based on maximum project transport envelopes. The review of the bulk steel supply will contribute to the plan.

The plan will include, but not be limited to:

Functional requirements of an inbound logistics system;

Assessment of existing transport nodes and linkages (ports, roads and rail);

Maximum load length, width, height and weight restrictions;

Specialist heavy-lift and over-dimensional transport requirements at port, for example, liaison with statutory authorities and utilities, permitting, road closures and escorts;

The requirement for “holding facilities” at port to manage the storage of equipment, materials and bulk steel pending transport to site;

The movement of over-size components to site;

Assessment of site conditions;

Identification of alternative operational model; methodologies, constraints and risks; and

The identification and management of shipping container and other demurrage costs.

The freight forwarder (or an independent consultant) will specifically review the movement of the bulk steel supply from place of manufacture to project site.

24.6.7Construction Management

The Engineer’s Construction Manager will establish a small on-site team prior to construction contractors mobilizing to site. The exact timing of the team’s establishment will be dependent on feedback from contractors regarding progress off site, but site establishment should not be less than four weeks in advance of contractor mobilization.

The Engineer’s Construction Manager will ensure that all construction contractors are responsible for:

Maintaining a safe site;

Maintaining compliance with all appropriate Statutory and Legislative requirements; and

Maintaining compliance with all Vista site requirements in regard to Environmental Health and Safety of the construction work force and the supervising team.

All site works will be undertaken utilizing qualified construction contractors, and the Construction Manager will act in the role of Superintendent to Vista when administering the construction contracts.

The construction supervision team will comprise of suitably qualified and experienced personnel and, where possible, preference will be given to more senior professionals when selecting staff, recognizing that the construction schedule and budget are of significant importance.

24.6.8Commissioning

The Engineer will develop a Commissioning Management Plan, in collaboration with the Vista Commissioning Representative. The Engineer’s Commissioning Manager will report to the EPCM Manager, but liaise closely with Vista’s Commissioning Representative. Three Commissioning Areas are contemplated:

Primary Crusher up to Mill;

Mill to Gold Room; and

Non Process Infrastructure (NPI).

Supervision of the various areas and disciplines during the discreet commissioning phases will be the responsibility of specifically appointed professional engineers assisted by key personnel from any design teams, construction teams, representatives of the various vendors and from the client’s staff.

Commissioning for the Process Plant will be generally carried out in three distinct phases:

Dry commissioning of all mechanical and electrical equipment including manual rotational checks, off load driven rotational checks, functional checks, instrument I/O checks, electrical continuity checks, etc.;

Wet commissioning of all mechanical and electrical equipment including hydraulic pressure testing using water, coupled with flow testing using water to ensure integrity of the various pumped circuits; and

Process commissioning of all mechanical and electrical equipment using production materials, commencing at minimum throughput requirement and gradually increasing to full design capacity prior to conducting any necessary performance testing.

Refer to Figure 24‑5 for Commissioning Phases bar chart.

Figure 24‑5: Commissioning Phases

All modifications required during commissioning will be documented in a Project Modification Register and subject to the same verification as detailed design with respect to design, fit for purpose, Environment, Health and Safety (EHS) and Hazard and Operability (HAZOP) Study requirements and drawing updates. All modifications will be carried out by either construction contractor’s representatives or vendor representatives.

24.6.9Temporary Construction Facilities

All Contractors will be responsible for the provision of their own site facilities to an appropriate standard that complies with local EHS guidelines for offices and amenities and to the approval of the Construction Manager.

All contractors will be responsible for the upkeep, cleaning and sanitation requirements of their respective facilities.

The EPCM Contractor will be responsible for the provision of suitable connection points for power, water and sewerage. The EPCM Contractor will be responsible for the provision of suitably located areas for the installation of the temporary facilities and for the provision of a suitably located receipt and lay-down area for delivered goods.

24.6.10Industrial Relations

All contractors will be required, under the terms of their contract, to take responsibility for their own industrial relations. They must be able to demonstrate and have in place suitable policies and procedures to ensure that the handling of matters of an Industrial Relations (IR) nature cause minimum disruption to the project schedule and budget.

All contractors must be able to demonstrate compliance with the HR/IR Policy. The Plan will be incorporated into all tender documentation. This plan will also contain details of the Site Agreement on wages and conditions that will apply universally to the project.

Contractors may be required to be affiliated to an equivalent Chamber of Commerce and Industry (CCI) for the Northern Territory. The CCI being a recognized and competent employer organization that can provide adequate IR advice and advocacy service, should the contractor fail to demonstrate the adequacy of his own internal services in this area.

All contractors should make an allowance to retain the CCI to develop suitable IR strategies, policies and procedures that will ensure that, in the event of industrial action being taken by contractors, the resolution of such matters will be timely and of such a nature as to not adversely affect the project schedule and budget.

24.6.11Health and Safety

All contractors will be required to comply with AS 4801 (Standard for Safety Management Systems) as a minimum.

All contractors must be able to demonstrate compliance with the EHS Project Management Plan. The Plan will be incorporated into all tender documentation.

The Engineer, in conjunction with Vista, will be responsible for developing a safety policy during the initial phase of the project. This policy should set out guidelines for the project safety procedures and the safety

targets for the project. Particular emphasis will be placed on site attendance of project personnel and the occupation of the site by the construction team and various contractors.

The policy will address the following issues:

The legislative responsibilities of Vista and the Engineer under the relevant Occupational Health, Safety and Welfare Act;

The legislative responsibilities of contractors under the relevant Occupational Health, Safety and Welfare Act;

The legislative responsibilities of employees under the relevant Occupational Health, Safety and Welfare Act;

The establishment of safety protocols and management systems required by the Act and how they will be practically implemented to suit the needs of the project; and

Any IR issues that need to be addressed as part of the overall safety management program.

All new employees attending site will be required to complete the necessary Vista site induction programs.

The Engineer will employ an experienced Safety Manager for the term of the project and a Site Safety Officer for the period of site occupation. The Safety Manager will be responsible for implementing the project safety policy, developing procedures in conjunction with the Engineer’s Site Safety Officer and implementing the provisions of the relevant Occupational Health, Safety and Welfare Act.

The Engineer’s Site Safety Officer will be responsible for enforcing all safety procedures and rules on the construction site and will organize regular communications with contractors to ensure adherence to policy, procedures and rules.

Contractors will be required to support the project safety protocols, provide individual safety management plans, perform Job Safety Analysis and ensure their employees are provided with Personal Protective Equipment to the standard defined by the overall site policy. Contractors must also provide a nominated individual at supervisory level, who has received adequate training in Occupational Health and Safety (OH&S), who will be responsible for safety procedures within the contract.

Contractors will be required to provide adequately equipped First Aid kits and have at least one formally qualified First Aid person on each shift to administer minor injuries not requiring medical attention from a Doctor. In the event of a more serious injury, Vista will make available the site First Aid facilities and personnel to all project related employees.

The Engineer’s Construction Manager will ensure that adequate records are kept of all safety incidents, irrespective of whether First Aid is required. TTP’s Site Safety Officer will report Lost Time Injury Frequency Rate, Disabling Injury Frequency Rate and Medically Treated Injury Frequency Rate, along with severity information, on a weekly basis as a minimum.

24.6.12Environment

The Engineer’s staff and all contractors will be made aware of the site environment conditions and constraints at the time of induction. Vista Environmental staff will be asked to audit site works on a periodic basis to identify issues of concern or non-conformance with site environmental policies and procedures.

All contractors must be able to demonstrate compliance with the EHS Project Management Plan. The Plan will be incorporated into all tender documentation.

24.6.13Schedule

24.6.13.1Schedule Objectives and Scope

The key objective of the PFS phase EPCM schedule is to provide a Class 3, Level 3 detail Schedule with an accuracy range of ±15%.

Class of Schedule defines the degree of completeness required for schedule development, Class 5 being a low degree of completeness, and Class 1 being a high degree of completeness. Level of Schedule defines the degree of detail for communication, reporting, and execution, Level 1 being a low degree of detail and Level 5 being a high degree of detail.

The scope included in the Schedule is that which is included in the EPCM contractor’s scope, as defined in the PFS. Consequently, Client Activities, Mine Development, Tailings Dam, Power plant detail, or Waste Water Treatment Plant are excluded from Schedule.

24.6.13.2Schedule Assumptions

For the Project, the specific schedule assumptions include:

The Northern Territory wet season runs from ~1st December to ~17th April when heavy rains can impact construction activities at times, particularly civil and concrete works;

No force majeure disruptions to scheduled work (IR or otherwise);

Open access to all work fronts is available;

Transportation to and from site (both air and land) is without delay; and

The schedule has assumed that project approval will be given by Vista on or about April 1, 2019. Start up, as defined by handover after completion of commissioning is scheduled to mid-2021.

24.6.13.3Critical Activities

The Critical Path of the EPCM Schedule runs through the Vista approval process and the purchase packages and contracts for Area 3300 (Classification and Grinding) as follows:

P001 Ball Mills Scope Development and Tender Period 11 weeks

P001 Ball Mills Manufacture and Delivery 63 weeks

.................P001 Ball Mills SMP Construction 24 weeks (Total)

Area 3300 Verification and Commissioning 5 weeks

The above critical activities determine a critical path of 119 weeks duration after Project approval to proceed has been given.

24.6.13.4Significant Activities

Major procurement packages with a lead time ex-works greater than 40 weeks are:

Table 24‑12: Supply Packages with Significant Lead Times

Package | Lead Time
P001 – Ball Mills | 55 weeks
P002 – Primary Crusher | 48 weeks
P004 – HPGRs | 48 weeks
P012 – Agitators | 60 weeks
P024 – Mill Relining Machine | 52 weeks
P029 – Ore Sorting | 45 weeks
P030 – Secondary Grinding Mills | 44 weeks

24.6.13.5Commissioning Schedule

The Commissioning Schedule has been broken into five specific activities in each area / sub area:

1)Construction Verification (CV)

a)Occurs immediately after construction completion for each area (e.g. 3100, 3200, etc.) with each area CV start date independent of the others

2)Pre commissioning

a)Occurs once CV is finished for each sub area

b)100,000t of ore available before pre-commissioning commences

3)Dry Commissioning (DC)

a)Requires equipment power up so each of the five sub stations (one for crushing / stockpile, one for HPGRs, one for milling, Secondary Grinding and one for leach / CIP / gold room / air / water services) need to be completed prior to commencing DC.

b)Should also occur in the order of:

Safety Systems (fire water / safety showers, etc.)

Process ancillary equipment (instrument air / gland water, etc.)

Process equipment substation groupings

Check Spares receipted into site store for equipment items in the area

4)Wet Commissioning (WC)

a)The order for WC needs to be:

Safety systems (fire water / safety showers, etc.)

Environmental systems (storm water pond pumps, sump pumps, etc.)

Process ancillary equipment (instrument air / gland water, etc.)

Process area where both the current area and downstream area dry commissioning has been completed

Workforce training completed

5)Process Commissioning

a)Will occur sequentially in the order of process flows, with the proviso that each area within a process zone terminated by a large storage buffer has been completed. Large storage buffers likely to create independent process commissioning zones include:

Crushed ore stockpile

Thickener

Leach / CIP Tanks

24.6.13.6Schedule Interfaces

The EPCM Schedule does not include detailed activities from contractors undertaking scopes of work outside the PFS scope such as the Tailings Dam, Waste Water Treatment Plant, Mine Development, and the Power Plant.

The Construction schedule is currently based on best estimate for the logical sequence of activities as developed by the Feasibility Study contractor. Upon award of contracts during the EP phase, construction contractors will be required to each develop and provide their schedules which will form a Class 3 Level 4 detailed schedule. This schedule will only be baselined with the approval of Vista, EPCM contractor, and the Construction contractor.

24.6.13.7Reporting

The Engineer’s Project Manager will ensure that the Schedule is updated within 3 working days of the end of each calendar month such that progress against project milestones and activities can be clearly identified. The project schedule will also show the critical path(s) at each update such that possible improvements in project completion forecast may be made.

Each month the EPCM contractor will provide the following to Vista and contractors:

The whole schedule

Critical Path/20 day or less Total Float view that will identify the critical path while also showing the activities with less than 20 days Total Float

Mid-month short form status report covering expenditure and schedule compliance

Figure 24‑6: EPCM Summary Schedule

24.7Alternate Case

Vista also prepared an Alternate Case which considers a smaller and higher-grade project. Key differences between the Base Case and the Alternate Case include:

A 33,000 tpd milling facility with associated lower mining rates and a smaller mining fleet;

An Alternate Case pit design based on a pit shell of US$800/oz-Au and cut-off grade of 0.40 g‑Au/t; and

A shorter LoM (11 years).

Results of the Alternate Case are presented in this Section.

24.7.1Mineral Resource

The same Mineral Resource, as used in the Base Case is used in the Alternate Case.

24.7.2Mining

24.7.2.1Pit Optimization

Separate pit designs were completed for the Base and Alternate Cases. Reserves are based on the Base Case ultimate pit design as this pit is larger and encompasses all material inside of the Alternate Case ultimate pit design. Economic Parameters are shown in Table 24‑13.

Table 24‑13: Initial Economic Parameters, Alternate Case

| Alternate Case
Gold Recovery | 85% Sulfide, 80% Transition, 80% Oxide
Payable Gold | 99.9%
Overall Mining Cost | US$2.16 per tonne
Processing Cost | US$8.65 per tonne processed
Tailings | US$0.90 per tonne processed
General & Administrative | US$0.77 per tonne processed
Water Treatment | US$0.10 per tonne processed
Royalty | 1% gross proceeds

Costs above reflect the costs used for pit optimization and do not reflect the final costs for the Alternate Case

As with the Base Case, the mining cost was varied using an additional US$0.010 per each 6 m bench below the 145 m elevation. The reference mining cost of US$1.86 was used for the Alternate Case. Processing, tailings construction, tailings reclamation, and water treatment costs were provided by Vista and based on previous studies. The total mining cost (reference plus incremental) is US$2.16 for the Alternate Case.

A minimum cutoff grade of 0.40 g-Au/t was used for the Alternate Case. This was done to maintain higher grades with respect to material allowed to be processed.

Pit optimizations were completed using prices of US$300 to US$2,000 per ounce Au in increments of US$20 per ounce in order to analyze the deposit’s sensitivity to gold prices for both scenarios. Table 24‑14 includes the US$1,250 Au price, which is highlighted in light green. The $800 pit shell is highlighted as the pit shell used as an ultimate pit guide. The pit optimizations only used measured and indicated resources. Inferred materials are considered waste.

Table 24‑14: Whittle TM Pit Optimization Results – Alternate Case Using 0.40 g-Au/t Cutoff

Pit | Gold Price | Material Processed | Waste Tonnes | Total Tonnes | Strip Ratio
K Tonnes | g-Au/t | K Ozs Au
1 | $ 300 | 966 | 1.88 | 59 | 731 | 1,697 | 0.76
6 | $ 400 | 6,467 | 1.63 | 338 | 5,647 | 12,114 | 0.87
11 | $ 500 | 11,133 | 1.46 | 523 | 10,299 | 21,432 | 0.93
16 | $ 600 | 21,103 | 1.25 | 850 | 22,828 | 43,931 | 1.08
21 | $ 700 | 60,702 | 1.06 | 2,078 | 115,883 | 176,585 | 1.91
26 | $ 800 | 95,474 | 0.98 | 3,019 | 185,830 | 281,304 | 1.95
31 | $ 900 | 121,301 | 0.92 | 3,586 | 222,805 | 344,106 | 1.84
36 | $ 1,000 | 154,092 | 0.87 | 4,307 | 295,849 | 449,941 | 1.92
41 | $ 1,100 | 185,525 | 0.85 | 5,085 | 427,833 | 613,358 | 2.31
46 | $ 1,200 | 202,606 | 0.84 | 5,504 | 512,606 | 715,212 | 2.53
49 | $ 1,250 | 212,904 | 0.84 | 5,753 | 569,384 | 782,289 | 2.67
52 | $ 1,300 | 223,045 | 0.84 | 5,996 | 626,265 | 849,310 | 2.81
57 | $ 1,400 | 234,858 | 0.83 | 6,278 | 700,519 | 935,376 | 2.98
62 | $ 1,500 | 241,417 | 0.83 | 6,454 | 756,553 | 997,970 | 3.13
67 | $ 1,600 | 247,350 | 0.83 | 6,605 | 809,868 | 1,057,218 | 3.27
71 | $ 1,700 | 250,425 | 0.83 | 6,684 | 841,140 | 1,091,565 | 3.36
75 | $ 1,800 | 254,050 | 0.83 | 6,779 | 880,583 | 1,134,633 | 3.47
80 | $ 1,900 | 254,353 | 0.83 | 6,785 | 883,275 | 1,137,628 | 3.47
84 | $ 2,000 | 259,140 | 0.83 | 6,908 | 943,012 | 1,202,152 | 3.64

Pit 26 was used for design purposes and Pit 49 illustrates the potential floating cone using a US$1,250/oz-Au price.

Graphs of the Whittle TM results are shown in Figure 24‑7.

Figure 24‑7: Measured and Indicated Graph of Whittle TM Results – Alternate Case Using 0.40 g-Au/t Cutoff

The ultimate pit limit for the Alternate Case was based on trying to reduce the capital requirements and the general overall size of the Project. This used a US$800/oz-Au pit shell for the Alternate Case.

24.7.2.2Pit Designs

Detailed pit design was completed, including an ultimate pit and two internal pits. The ultimate pit was designed to allow mining economic resources identified by Whittle TM pit optimization, while providing safe access for people and equipment. Internal pits or phases within the ultimate pit were designed to enhance the project by providing higher-value material to the processing plant earlier in the mine life.

The ultimate pit design along with the ultimate dump and stockpile designs and planned infrastructure is shown in Figure 24‑8.

Figure 24‑8: Mt Todd Ultimate Pit Design – Alternate Case (February 12, 2018)

Separate phase designs for the Alternate and Base Cases were created. For the Alternate Case, Phase 1 essentially continues the western wall down from that done by prior operators, and wraps the ramp around the pit clockwise from the south. Phase 2 expands the pit to the east, north, and south, maintaining a portion of the phase 1 west wall. The Phase 2 ramp is placed on the east wall and has a total of 5 switchbacks located in the north and south ends of the pit. For the Alternate case there is only one more phase of mining to achieve the ultimate pit. This phase mines 360 degrees around the phase 2 pit, deepening it to achieve the ultimate pit.

Figure 24‑9 to Figure 24‑10 show the Alternate Case Phase I and II pit designs. The Alternate Case Phase IV design is depicted in Figure 24‑8 as the ultimate pit. Resulting reserves for each of the phases are shown in Figure 24‑9.

Figure 24‑9: Phase I Pit Design – Alternate Case (February 12, 2018)

Figure 24‑10: Phase II Pit Design – Alternate Case (February 12, 2018)

24.7.2.3Cut-Off Grade

The breakeven and internal cutoff grades calculated using the economic parameters shown in Table 24‑13 are shown in Table 24‑15. The internal cutoff grade assumes that mining is constrained to an economic pit and does not include the mining cost.

To enhance projects economics, Vista used a higher cutoff grade for reserves and scheduling than what operating costs would have predicted. Reserves are reported using a 0.40 g-Au/t cutoff grade for the Alternate Case.

Table 24‑15: US$1,250 Calculated Gold Price Cutoff Grades (g-Au/t)

| Sulfide | Transition | Oxide
Breakeven | 0.38 | 0.40 | 0.40
Internal | 0.31 | 0.33 | 0.33
Cutoff Grade Used | 0.40 | 0.40 | 0.40

For purposes of production scheduling, low-grade, medium-grade, and high-grade material was designated. The low-grade material used a 0.40g-Au/t cutoff grade. Medium-grade and high-grade material is defined using cutoffs of 0.55 and 0.85 g-Au/t.

24.7.2.4Mineral Reserves

Mineral Reserves are shown in Table 24‑16.

Note that these proven and probable Reserves are contained within the Base Case pit designs used to state the full official reserves. Thus, the Alternate Case Reserves are a subset of the official reserves.

Table 24‑16: Alternate Case Proven and Probable Reserves by Phase

| Proven | Probable | Total P&P | Waste Total | Total Tonnes | Strip Ratio
K Tonnes | g-Au/t | K Ozs Au | Tonnes | g-Au/t | K Ozs Au | Tonnes | g-Au/t | K Ozs Au
Ph_1 | 13,551 | 1.09 | 473 | 6,245 | 1.10 | 221 | 19,796 | 1.09 | 694 | 19,312 | 39,109 | 0.98
Ph_2 | 18,980 | 0.80 | 490 | 19,008 | 0.88 | 538 | 37,988 | 0.84 | 1,028 | 59,167 | 97,155 | 1.56
Ph_3 | 21,500 | 0.91 | 626 | 35,375 | 0.86 | 977 | 56,874 | 0.88 | 1,603 | 114,836 | 171,710 | 2.02
Total | 54,031 | 0.91 | 1,589 | 60,628 | 0.89 | 1,736 | 114,658 | 0.90 | 3,325 | 193,316 | 307,974 | 1.69

The 33,000 tpd reserves are reported using a cutoff grade of 0.40 g-Au/t and are a subset of the measured and indicated resources; inferred resources are considered waste.

MINE WASTE FACILITIES

Total contained waste tonnage is 193 Mt for the Alternate Case. The Alternate Case designed height is 234 meters. The south end of the Base Case dump was designed to encroach on the existing RP1 waste water storage facility. Designs are limited to the east by the process facility and to the west by another drainage basin. Designs are intended to promote any drainage to the current RP1 waste water retention pond.

MINE PRODUCTION SCHEDULE

Table 24‑17 shows the mine-production schedule, including re-handle from stockpiles. For purpose of production scheduling, low-grade, medium-grade, and high-grade material was designated. The low-grade material used a cutoff grade of 0.40 g-Au/t. Medium-grade and high-grade cutoffs used were 0.55 and 0.85 g‑Au/t.

Table 24‑17: Annual Mine Production Schedule – Alternate Case

|  |  | Pre-Prod | Yr 1 | Yr 2 | Yr 3 | Yr 4 | Yr 5 | Yr 6 | Yr 7 | Yr 8 | Yr 9 | Yr 10 | Yr 11 | Yr 12 | Yr 13 | Total
Total Mined | *StkPl | K Tonnes | 2,105 | 4,642 | 2,104 | 5,619 | 9,187 | 2,408 | 2,605 | 3,421 | 7,262 | 1,330 | - | - | - | - | 40,684
| g-Au/t | 0.73 | 0.65 | 0.53 | 0.50 | 0.69 | 0.47 | 0.47 | 0.47 | 0.60 | 0.48 | - | - | - | - | 0.59
| K Ozs Au | 49 | 98 | 36 | 91 | 205 | 36 | 39 | 52 | 141 | 21 | - | - | - | - | 767
Crusher | K Tonnes | - | 9,405 | 6,135 | 10,095 | 11,136 | 2,251 | 3,376 | 8,802 | 11,715 | 11,059 | - | - | - | - | 73,974
| g-Au/t | - | 1.22 | 1.22 | 0.82 | 1.17 | 0.72 | 0.78 | 0.91 | 1.13 | 1.25 | - | - | - | - | 1.08
| K Ozs Au | - | 368 | 240 | 265 | 420 | 52 | 84 | 258 | 425 | 446 | - | - | - | - | 2,559
Total Ore Mined | K Tonnes | 2,105 | 14,047 | 8,239 | 15,714 | 20,323 | 4,660 | 5,981 | 12,224 | 18,977 | 12,389 | - | - | - | - | 114,658
| g-Au/t | 0.73 | 1.03 | 1.04 | 0.70 | 0.96 | 0.59 | 0.64 | 0.79 | 0.93 | 1.17 | - | - | - | - | 0.90
| K Ozs Au | 49 | 466 | 276 | 355 | 625 | 88 | 123 | 310 | 566 | 466 | - | - | - | - | 3,325
NonPag_Wst | K Tonnes | 1,786 | 311 | 15,892 | 4,732 | 9,592 | 17,281 | 8,450 | 2,041 | 63 | - | - | - | - | - | 60,148
Pag_Wst | K Tonnes | 4,944 | 9,258 | 7,782 | 15,366 | 10,278 | 11,645 | 12,634 | 13,041 | 9,578 | 1,482 | - | - | - | - | 96,008
Un_Wst | K Tonnes | 1,754 | 1,428 | 5,022 | 4,996 | 4,248 | 6,955 | 6,150 | 5,269 | 1,312 | 28 | - | - | - | - | 37,160
Total Waste Mined | K Tonnes | 8,484 | 10,997 | 28,696 | 25,094 | 24,118 | 35,880 | 27,234 | 20,350 | 10,953 | 1,510 | - | - | - | - | 193,316
Total Tonnes Mined | K Tonnes | 10,589 | 25,044 | 36,935 | 40,808 | 44,441 | 40,540 | 33,215 | 32,574 | 29,930 | 13,899 | - | - | - | - | 307,974
Strip Ratio | W:O | 4.03 | 0.78 | 3.48 | 1.60 | 1.19 | 7.70 | 4.55 | 1.66 | 0.58 | 0.12 |  |  |  |  | 1.69
Re-Handle Material | HG_StkPl | K Tonnes | - | 554 | 879 | - | 579 | 1,274 | - | - | - | 656 | 95 | - | - | - | 4,037
| g-Au/t | - | 1.08 | 1.27 | - | 1.45 | 1.18 | - | - | - | 1.05 | 0.89 | - | - | - | 1.19
| K Ozs Au | - | 19 | 36 | - | 27 | 48 | - | - | - | 22 | 3 | - | - | - | 155
MG_StkPl | K Tonnes | - | 1,060 | 1,153 | - | - | 3,668 | - | - | - | - | 2,640 | - | - | - | 8,521
| g-Au/t | - | 0.68 | 0.68 | - | - | 0.67 | - | - | - | - | 0.67 | - | - | - | 0.67
| K Ozs Au | - | 23 | 25 | - | - | 79 | - | - | - | - | 57 | - | - | - | 184
LG_StkPl | K Tonnes | - | 697 | 3,547 | 1,652 | - | 4,522 | 8,339 | 2,945 | - | - | 6,425 | - | - | - | 28,126
| g-Au/t | - | 0.51 | 0.48 | 0.49 | - | 0.51 | 0.45 | 0.48 | - | - | 0.47 | - | - | - | 0.47
| K Ozs Au | - | 12 | 55 | 26 | - | 74 | 120 | 46 | - | - | 96 | - | - | - | 427
Leach Re-handle | K Tonnes | - | - | - | - | - | - | - | - | - | - | 2,556 | 10,798 | - | - | 13,354
| g-Au/t | - | - | - | - | - | - | - | - | - | - | 0.54 | 0.54 | - | - | 0.54
| K Ozs Au | - | - | - | - | - | - | - | - | - | - | 44 | 187 | - | - | 232
Total Re-Handle | K Tonnes | - | 2,310 | 5,580 | 1,652 | 579 | 9,464 | 8,339 | 2,945 | - | 656 | 11,715 | 10,798 | - | - | 54,038
| g-Au/t | - | 0.73 | 0.64 | 0.49 | 1.45 | 0.66 | 0.45 | 0.48 | - | 1.05 | 0.53 | 0.54 | - | - | 0.57
| K Ozs Au | - | 54 | 115 | 26 | 27 | 201 | 120 | 46 | - | 22 | 200 | 187 | - | - | 998
| Waste Re-handle | K Tonnes | 53 | 366 | - | 3,370 | - | - | - | 1,354 | 462 | 5,285 | - | - | 315 | 13,675 | 24,882
| Sorter Rejects | K Tonnes | - | 1,171 | 1,171 | 1,175 | 1,171 | 1,171 | 1,171 | 1,175 | 1,171 | 1,171 | 916 | - | - | - | 11,466
| Sorter Reject Re-handle | K Tonnes | - | - | - | - | - | - | - | - | - | - | - | - | - | 11,466 | 11,466

Table 24‑18: Annual Stockpile Balance – Alternate Case

|  |  | Pre-Prod | Yr 1 | Yr 2 | Yr 3 | Yr 4 | Yr 5 | Yr 6 | Yr 7 | Yr 8 | Yr 9 | Yr 10 | Yr 11
Hg_StkPl | Added | K Tonnes | 528 | 777 | 127 | 105 | 1,748 | - | - | - | 751 | - | - | -
| g-Au/t | 1.08 | 1.26 | 1.27 | 1.28 | 1.26 | - | - | - | 1.03 | - | - | -
| K Ozs Au | 18 | 31 | 5 | 4 | 71 | - | - | - | 25 | - | - | -
Removed | K Tonnes | - | 554 | 879 | - | 579 | 1,274 | - | - | - | 656 | 95 | -
| g-Au/t | - | 1.08 | 1.27 | - | 1.45 | 1.18 | - | - | - | 1.05 | 0.89 | -
| K Ozs Au | - | 19 | 36 | - | 27 | 48 | - | - | - | 22 | 3 | -
Balance | K Tonnes | 528 | 752 | - | 105 | 1,274 | - | - | - | 751 | 95 | - | -
| g-Au/t | 1.08 | 1.27 | - | 1.28 | 1.18 | - | - | - | 1.03 | 0.89 | - | -
| K Ozs Au | 18 | 31 | - | 4 | 48 | - | - | - | 25 | 3 | - | -
Mg_StkPl | Added | K Tonnes | 945 | 1,207 | 61 | 489 | 3,179 | - | - | - | 2,640 | - | - | -
| g-Au/t | 0.69 | 0.67 | 0.71 | 0.66 | 0.67 | - | - | - | 0.67 | - | - | -
| K Ozs Au | 21 | 26 | 1 | 10 | 69 | - | - | - | 57 | - | - | -
Removed | K Tonnes | - | 1,060 | 1,153 | - | - | 3,668 | - | - | - | - | 2,640 | -
| g-Au/t | - | 0.68 | 0.68 | - | - | 0.67 | - | - | - | - | 0.67 | -
| K Ozs Au | - | 23 | 25 | - | - | 79 | - | - | - | - | 57 | -
Balance | K Tonnes | 945 | 1,092 | - | 489 | 3,668 | - | - | - | 2,640 | 2,640 | - | -
| g-Au/t | 0.69 | 0.68 | - | 0.66 | 0.67 | - | - | - | 0.67 | 0.67 | - | -
| K Ozs Au | 21 | 24 | - | 10 | 79 | - | - | - | 57 | 57 | - | -
Lg_StkPl | Added | K Tonnes | 632 | 2,658 | 1,916 | 5,025 | 4,260 | 2,408 | 2,605 | 3,421 | 3,871 | 1,330 | - | -
| g-Au/t | 0.48 | 0.47 | 0.47 | 0.47 | 0.47 | 0.47 | 0.47 | 0.47 | 0.47 | 0.48 | - | -
| K Ozs Au | 10 | 40 | 29 | 76 | 65 | 36 | 39 | 52 | 59 | 21 | - | -
Removed | K Tonnes | - | 697 | 3,547 | 1,652 | - | 4,522 | 8,339 | 2,945 | - | - | 6,425 | -
| g-Au/t | - | 0.51 | 0.48 | 0.49 | - | 0.51 | 0.45 | 0.48 | - | - | 0.47 | -
| K Ozs Au | - | 12 | 55 | 26 | - | 74 | 120 | 46 | - | - | 96 | -
Balance | K Tonnes | 632 | 2,593 | 961 | 4,334 | 8,594 | 6,481 | 747 | 1,223 | 5,094 | 6,425 | - | -
| g-Au/t | 0.48 | 0.46 | 0.42 | 0.45 | 0.46 | 0.44 | 0.42 | 0.43 | 0.46 | 0.47 | - | -
| K Ozs Au | 10 | 38 | 13 | 63 | 128 | 91 | 10 | 17 | 76 | 96 | - | -
All StkPl | Balance | K Tonnes | 2,105 | 4,437 | 961 | 4,928 | 13,536 | 6,481 | 747 | 1,223 | 8,486 | 9,159 | - | -
| g-Au/t | 0.73 | 0.65 | 0.42 | 0.49 | 0.59 | 0.44 | 0.42 | 0.43 | 0.58 | 0.53 | - | -
| K Ozs Au | 49 | 93 | 13 | 78 | 255 | 91 | 10 | 17 | 157 | 156 | - | -

Table 24‑19: Annual Ore Delivery to the Mill Crusher – Alternate Case

|  | Pre-Prod | Yr 1 | Yr 2 | Yr 3 | Yr 4 | Yr 5 | Yr 6 | Yr 7 | Yr 8 | Yr 9 | Yr 10 | Yr 11 | Total
Sulfide Ore | K Tonnes | - | 11,715 | 11,559 | 11,744 | 11,531 | 11,090 | 11,371 | 11,734 | 11,715 | 11,715 | 9,137 | - | 113,311
g-Au/t | - | 1.12 | 0.95 | 0.77 | 1.19 | 0.67 | 0.55 | 0.80 | 1.13 | 1.24 | 0.53 | - | 0.91
K Ozs Au | - | 422 | 353 | 291 | 442 | 240 | 199 | 303 | 425 | 468 | 155 | - | 3,300
Recovery | 0% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 0% | 85%
K Ozs Au Rec | - | 361 | 301 | 248 | 377 | 205 | 170 | 259 | 363 | 399 | 133 | - | 2,817
Mixed Ore | K Tonnes | - | - | 87 | 2 | 137 | 550 | 273 | 11 | - | - | 18 | - | 1,077
g-Au/t | - | - | 0.53 | 0.42 | 0.82 | 0.63 | 0.44 | 0.42 | - | - | 0.42 | - | 0.59
K Ozs Au | - | - | 1 | 0 | 4 | 11 | 4 | 0 | - | - | 0 | - | 20
Recovery | 0% | 0% | 78% | 78% | 78% | 78% | 78% | 78% | 0% | 0% | 78% | 0% | 78%
K Ozs Au Rec | - | - | 1 | 0 | 3 | 9 | 3 | 0 | - | - | 0 | - | 16
Oxidized Ore | K Tonnes | - | - | 69 | 1 | 47 | 75 | 71 | 3 | - | - | 4 | - | 270
g-Au/t | - | - | 0.60 | 0.42 | 1.11 | 0.55 | 0.44 | 0.42 | - | - | 0.42 | - | 0.63
K Ozs Au | - | - | 1 | 0 | 2 | 1 | 1 | 0 | - | - | 0 | - | 5
Recovery | 0% | 0% | 78% | 78% | 78% | 78% | 78% | 78% | 0% | 0% | 78% | 0% | 78%
K Ozs Au Rec | - | - | 1 | 0 | 1 | 1 | 1 | 0 | - | - | 0 | - | 4
Total Ore | K Tonnes | - | 11,715 | 11,715 | 11,747 | 11,715 | 11,715 | 11,715 | 11,747 | 11,715 | 11,715 | 9,159 | - | 114,658
g-Au/t | - | 1.12 | 0.94 | 0.77 | 1.19 | 0.67 | 0.54 | 0.80 | 1.13 | 1.24 | 0.53 | - | 0.90
K Ozs Au | - | 422 | 356 | 291 | 447 | 253 | 204 | 304 | 425 | 468 | 156 | - | 3,325
Recovery | 0% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 0% | 85%
K Ozs Au Rec | - | 361 | 304 | 248 | 382 | 215 | 174 | 259 | 363 | 399 | 133 | - | 2,837

For the purpose of scheduling, three ore stockpiles are assumed: High-grade ore stockpile for high-grade ore; medium-grade stockpile for medium-grade ore; and a low-grade stockpile for low-grade ore. The high-grade and medium-grade stockpiles are to be built within the low-grade stockpiling areas but are exhausted during the first year of processing when mill capacity becomes available. During the LoM, the low-grade stockpile is used as needed to feed the mill to full capacity. For this reason the stockpile grows and shrinks through the LoM. The maximum stockpile balance through the LoM is estimated to be 13.5 Mt.

Re-handling of stockpiled material will be done using a loader and trucks to haul ore to the crusher. Table 24‑18 shows the Alternate Case ore stockpile balances for the end of each year. Waste re-handle is shown on the bottom of Table 24‑17 to account for capping and reclamation.

Ore sent to the mill is shown in Table 24‑19. This is a combination of ore shipped directly from the mine and ore that is reclaimed from stockpiles. These tables summarize the ore based on level of oxidation. The recovered ounces shown are based on the recoveries used for pit optimizations and are subject to change by qualified persons completing the metallurgical sections

EQUIPMENT SELECTION AND PRODUCTIVITIES

Availability, efficiencies, operating hours and load and haul equipment requirements are shown in Table 24‑17.

Table 24‑20: Annual Load and Haul Equipment Requirements – Alternate Case

| Units | Pre-Prod | Yr 1 | Yr 2 | Yr 3 | Yr 4 | Yr 5 | Yr 6 | Yr 7 | Yr 8 | Yr 9 | Yr 10 | Yr 11 | Yr 12 | Yr 13 | Total
HAULAGE REQUIREMENTS
Productive Hours | Hrs | 12,009 | 38,337 | 60,232 | 81,440 | 77,907 | 74,838 | 73,134 | 77,468 | 78,130 | 45,391 | 8,968 | 8,450 | 441 | 30,087 | 666,831
Operating Efficiency | % | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 1743%
Operating Hours | Hrs | 14,468 | 46,189 | 72,569 | 98,120 | 93,863 | 90,166 | 88,114 | 93,335 | 94,133 | 54,688 | 10,805 | 10,181 | 531 | 36,249 | 803,411
Number of Trucks | # | 5 | 8 | 12 | 15 | 15 | 15 | 15 | 15 | 15 | 7 | 4 | 4 | 2 | 8 | 140
Truck Availability | % | 90% | 89% | 89% | 88% | 87% | 86% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% |
Available Operating Hours | Hrs | 19,963 | 49,178 | 76,102 | 96,517 | 95,155 | 93,953 | 93,225 | 93,050 | 92,788 | 43,301 | 24,744 | 24,813 | 6,012 | 49,487 | 858,287
Use of Available Hours | % | 79% | 94% | 95% | 102% | 99% | 96% | 95% | 100% | 101% | 126% | 44% | 41% | 9% | 73% | 94%
Tonnes per Operating Hour | t/Hr | 689 | 600 | 586 | 467 | 480 | 555 | 472 | 395 | 323 | 363 | 1,084 | 1,061 | 594 | 377 | 482
HYDRAULIC SHOVEL USAGE
Number of Shovels | # | 1 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 1 | 1 | 1 | 1 | 1.9
Availability | % | 90.0% | 89.4% | 88.5% | 87.5% | 86.5% | 85.5% | 85.0% | 85.0% | 85.0% | 85.0% | 85.0% | 85.0% | 85.0% | 85.0% | 86.4%
Operating Efficiency | % | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83% | 83.0%
Available Operating Hrs | Op Hrs | 4,889 | 13,027 | 12,881 | 12,772 | 12,590 | 12,445 | 12,372 | 12,407 | 12,372 | 12,372 | 12,372 | 6,203 | 3,006 | 6,186 | 145,892
Tonnes Mined | K Tonnes | 10,589 | 24,213 | 36,844 | 39,176 | 42,663 | 40,540 | 33,215 | 32,574 | 28,733 | 13,899 | 9,372 | 8,639 | 252 | 18,856 | 339,564
Operating Hours | Op Hrs | 1,831 | 6,169 | 9,387 | 9,981 | 10,869 | 10,328 | 8,462 | 8,299 | 7,320 | 3,541 | 2,388 | 2,201 | 64 | 4,804 | 85,645
Use of Available Operating Hours | % | 37% | 47% | 73% | 78% | 86% | 83% | 68% | 67% | 59% | 29% | 19% | 35% | 2% | 78% | 59%
FRONT END LOADERS
Number of Loaders | # | - | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1.0
Availability | % | 0% | 90% | 89% | 88% | 87% | 86% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 85% | 86%
Operating Efficiency | % | 0.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0% | 83.0%
Available Operating Hrs | Op Hrs | - | 6,550 | 6,477 | 6,422 | 6,331 | 6,259 | 6,186 | 6,203 | 6,186 | 6,186 | 6,186 | 6,203 | 3,006 | 6,186 | 78,381
Tonnes Mined | K Tonnes | - | 3,507 | 5,671 | 6,655 | 2,357 | 9,464 | 8,339 | 4,299 | 1,659 | 5,942 | 2,343 | 2,160 | 63 | 6,285 | 58,743
Operating Hours | Op Hrs | - | 2,089 | 3,377 | 3,963 | 1,404 | 5,636 | 4,966 | 2,560 | 988 | 3,539 | 1,395 | 1,286 | 38 | 3,743 | 34,985
Use of Available Operating Hours | % | 0% | 32% | 52% | 62% | 22% | 90% | 80% | 41% | 16% | 57% | 23% | 21% | 1% | 61% | 45%

MINE PERSONNEL

Salaries for each position were estimated based on information received from Tetra Tech and Vista. Salaries include an allowance for benefits at a rate of 27% of the base salary for each position. Note that mobile equipment labor costs are allocated to production equipment in the calculation of mining costs in later sections. In addition, a portion of the cost is allocated to construction of tailings facilities.

24.7.3Process Facility

24.7.3.1Design Criteria

The nominal headline design criteria are listed are shown in Table 24‑21.

Table 24‑21: Headline Design Criteria

| Unit | 33,000 tpd
Annual Ore Feed Rate | Mt/a | 11.72
Operating Days per Year | d/y | 355
Daily Ore Feed Rate | t/d | 33,000
Crushing Rate (6,637 hours per year availability) | tph | 1,765
HPGR Rate (7,838 hours per year) | tph | 1,495
Ore Sorting Rate (7,838 hours per year) | tph | 318
Milling Rate (7,838 hours per year) | tph | 1,332
Gold Head Grade | g/t | 0.82
Copper Head Grade | % | 0.055
Cyanide Soluble Copper | % | 0.0024
Ore Specific Gravity |  | 2.76
Primary Grind P80 to Secondary Grind | µm | 250
Grind P80 to Leach | µm | 60
Gold Recovery | % | 85
Gold Production (average) | oz/d | 740
Gold Production (average) | oz/a | 262,634

FLOWSHEET

The Alternate Case flowsheet is the same as that of the Base Case, as shown in Figure 17‑1 with the following equipment differences:

Cone Crushers: 2- Raptor 900 secondary crushers;

HPGRs: 2- HPGR Polycom PM7-20/15, each equipped with 2 x 1,800 kW drives;

Thickeners: 36m diameter; and

Leach/CIP: Carbon movement requirements will be in the order of 20 tpd.

24.7.4Infrastructure

Project support infrastructure and services are similar to the Base Case, but with the following differences:

24.7.4.1Area 2300 – Mine Support Facilities

Sub-Area 2305 – Support Facilities – HV Workshop / Warehouse

The HV Workshop will reduce in size for the Alternate Case from six bays to four bays, resulting in a footprint of 59.3 m by 24.4 m.

There will be no difference in size between the Base Case and the Alternate Case for the Warehouse.

Sub-Area 2310 – Support Facilities – Bulk Fuel Storage

The bulk fuel storage for the Alternate Case will require four bowsers as opposed to six.

Sub-Area 2335 – Support Facilities – Lube Storage

The Lube Storage area for the Alternate Case will have a reduced footprint for the storage of lubricants.

24.7.4.2Area 4000 – Project Services

Sub-Area 4130 – Potable Water

Water supply to the project does not differ greatly between the two cases except that the water treatment plant is sized to handle a smaller throughput and several supply pumps will be decreased in size.

Raw water will be brought down to the raw water tank in the process plant through the existing 400 mm poly line at the raw water dam.

Sub-Area 4231 – Power Distribution

For the 33,000 tpd case, the current carrying capacity of the buried cable between the overhead power lines and the Process Plant will be decreased to suit the reduction in plant load.

Sub-Area 4232 – Overhead Power Lines

For the 33,000 tpd case, the current carrying capacity of the overhead power lines between the Power Station and the Process Plant will be decreased to suit the reduction in plant load.

AREA 4300 – COMMUNICATIONS

Sub-Area 4310 – Fiber Optic

The Alternate Case will essentially be the same as the Base Case.

Sub-Area 4311 – Phones

The Alternate Case will essentially be the same as the Base Case. The only difference may be the number of telephone handsets required due to a decrease in site personnel.

Sub-Area 4313 – Telemetry

The Alternate Case will essentially be the same as the Base Case.

AREA 4500 – PLANT MOBILE EQUIPMENT

The plant mobile equipment to be purchased for the Alternate Case will not differ from the Base Case.

24.7.4.3Area 5000 – Project Infrastructure

AREA 5100 – SITE PREPARATION

The site footprint for the 33,000 tpd case will be smaller than the 50,000 tpd case and as a result will have reduced earthworks, road works and drainage quantities.

SUB-AREA 5230 – REAGENT STORE

The Alternate Case will require a reduced inventory of reagents and therefore the footprint of the Reagents Yard will be reduced to an area of 1,350 square meters (m 2).

SUB-AREA 5260 – SAMPLE PREPARATION AND LABORATORY

The Alternate Case requires a reduced inventory of sample processing laboratory equipment to process 300 samples/day.

AREA 5400 – HEAVY LIFT CRANAGE

The heavy lift craneage durations will be reduced for the Alternate Case based on a reduced Structural, Mechanical and Piping (SMP) contract schedule.

24.7.4.4Area 6100 – Personnel Transport

The bus transit area for the Alternate Case will be identical to that proposed for the Base Case.

24.7.4.5Area 7300 – Construction Camp

The Construction Camp for the Alternate Case will be sized for 350 construction workers based on the manning histogram developed for the Project. There will be a reduction in footprint and associated infrastructure, establishing access roads and site works, contractor preliminaries, transportable buildings and services.

24.7.5Site-wide Water Balance Model

The Alternate Case site-wide water balance model was developed using scaled flows of the Base Case water balance. Differences between the Alternate Case and the Base Case models are:

Total process plant makeup water for the Alternative Case is 1,164 m 3/hr versus 1,764 m 3/hr;

Alternate Case production occurs over 11 years versus 13 years;

Process circuit bleed stream is 151 m 3/hr for the Alternate Case and 229 m 3/hr for the Base Case;

RWD process makeup water flows were 200 m 3/hr for the Alternate Case and 304 m 3/hr for the Base Case;

Maximum draw from the RWD for make-up water in the Alternate Case is 7,890 m 3/day versus 11,955 m 3/day; and

Other water requirements (TSF decant, potable, reagent mixing, gland, etc.) are consistent with lower production rates.

24.7.6Capital Costs, Alternate Case

LoM capital cost requirements are estimated at US$870 million as summarized in Table 21‑2. Initial capital of US$641 million is estimated to be required to commence operations. At the end of operations, the Project will receive a US$96 million credit for remaining asset sales and salvage.

Table 24‑22: Estimated Alternate Case Capital Cost Summary (US$000s)

Area | Description | Cont. (%) | Initial Capital (US$000s) | Sustaining Capital (US$000s) | Total Capital (US$000s)
Estimate | Contingency | Total | Estimate | Contingency | Total | Estimate | Contingency | Total
SUMMARY
2000 | Mining | 0.4% | $75,599 | $1,194 | $76,793 | $264,265 | $168 | $264,433 | $339,863 | $1,363 | $341,226
3000 | Process Plant | 21.6% | $264,474 | $57,888 | $322,363 | $11,209 | $1,537 | $12,746 | $275,683 | $59,425 | $335,109
4000 | Project Services | 4.0% | $97,835 | $5,757 | $103,592 | $47,687 | $0 | $47,687 | $145,522 | $5,757 | $151,279
5000 | Project Infrastructure | 12.9% | $22,881 | $2,958 | $25,838 | $0 | $0 | $0 | $22,881 | $2,958 | $25,838
6000 | Permanent Accommodation | 10.0% | $68 | $7 | $75 | $0 | $0 | $0 | $68 | $7 | $75
7000 | Site Establishment & Early Works | 11.4% | $18,579 | $2,127 | $20,706 | $0 | $0 | $0 | $18,579 | $2,127 | $20,706
8000 | Management, Engineering, EPCM Svcs | 9.0% | $74,977 | $6,720 | $81,697 | $0 | $0 | $0 | $74,977 | $6,720 | $81,697
9000 | Pre-Production Costs | 17.3% | $8,734 | $1,509 | $10,243 | $0 | $0 | $0 | $8,734 | $1,509 | $10,243
10000 | Asset Sale | 0.0% | $0 | $0 | $0 | ($96,170) | $0 | ($96,170) | ($96,170) | $0 | ($96,170)
| Capital Cost | 10.1% | $563,147 | $78,160 | $641,307 | $226,991 | $1,706 | $228,697 | $790,138 | $79,865 | $870,004

24.7.7Operating Costs, Alternate Case

LoM operating cost estimates are summarized in Table 24‑23. The operating costs will average US$13.61 over the LoM.

Table 24‑23: Estimated Operating Cost Summary, Alternate Case (US$)

| US$/t-milled
OPEN PIT MINE |
Mine General Service | $0.10
Mine Maintenance | $0.16
Engineering | $0.07
Geology | $0.05
Drilling | $0.52
Blasting | $0.84
Loading | $0.45
Hauling | $1.69
Mine Support | $0.52
Mine Dewatering | $0.01
Open Pit Mine | $4.40
CIP PROCESS PLANT |
Labor | $1.22
3100-Crush/Screen/Stockpile | $0.24
3200-Reclaim & HPGR | $0.55
3300-Classification & Grinding | $2.69
3400-Pre-Leach,Thick/Aeration/CIP | $0.17
3500-Desorption, Gold Room | $0.02
3600-Detox & Tailings Pumping | $0.07
3700-Reagents | $2.12
3800-Plant Services | $0.04
Mining, Infrastructure & Misc | $0.06
General Consumables | $0.01
Plant Mobile Equipment | $0.01
Plant Gas Consumption | $0.02
CIP Process Plant | $7.23
Project Services | $0.18
G&A | $1.80
Operating Costs | $13.61

24.7.8Economic Results, Alternate Case

Project cost estimates and economics are prepared on an annual basis. Based upon design criteria presented in this report, the level of accuracy of the estimate is considered ±25%.

Economic results are summarized in Table 24‑24. The analysis suggests the following conclusions, assuming a 100% equity project at a gold price of US$1,300:

Mine Life: 11 years;

Pre-tax NPV5%: US$722 million, IRR: 23.4%;

After-tax NPV5%: US$418 million, IRR: 17.8%;

Payback (After-tax): 3.6 years;

NT Royalty Taxes Paid: US$209 million;

Australian Income Taxes Paid: US$282 million; and

Cash costs (including JAAC Royalty): US$593.38/oz-Au.

Table 24‑24: Economic Results, Alternate Case (US$000s)

Cash Flow Summary | LoM Cost (US$000s) | Unit Cost US$/t-milled | Unit Cost US$/oz-Au
Gold Produced (koz) | 3,003 | - | -
Gold Price (US$/oz) | 1,300 | - | -
Gold Sales | 3,903,895 | 30.50 | 1,300
Refinery Costs | (9,737) | ($0.076) | ($3.242)
Net Smelting Return | 3,894,158 | $ 30.420 | 1,297
JAAC Royalty | (39,039) | (0.305) | (13.00)
Gross Income from Mining | 3,855,119 | ($30.115) | 1,284
Open Pit Mine | (563,491) | (4.40) | (188)
CIP Process Plant | (925,686) | ($7.23) | (308)
Project Services | (23,326) | (0.18) | (8)
G&A | (230,369) | ($1.80) | (77)
Operating Costs | (1,742,873) | ($13.61) | (580)
Power Sales Credit | 13,056 | 0.102 | 4
Cash Cost of Goods Sold (COGS) | (1,739,554) | ($13.59) | (579)
Operating Margin | 2,125,302 | $ 16.60 | 708
Mining | $341,226 |  |
Process Plant | $335,109 |  |
Project Services | $151,279 |  |
Project Infrastructure | $25,838 |  |
Permanent Accommodation | $75 |  |
Site Establishment & Early Works | $20,706 |  |
Management, Engineering, EPCM Services | $81,697 |  |
Pre-Production Costs | $10,243 |  |

Asset Sale | ($96,170) |  |
Capital Costs | $870,004 |  |
Pre-Tax Cash Flow | 1,255,298 |  |
NPV5% | $ 721,977 |  |
IRR (%) | 23.4% |  |
After-tax Cash Flow | 778,521 |  |
NPV5% | $ 417,938 |  |
IRR (%) | 17.8% |  |
Post- Tax Payback (years) | 3.6 |  |

Project sensitivities are summarized in Table 24‑5, Table 24‑6, and Table 24‑27; sensitivities are shown graphically in Figure 24‑11. As seen, the Project is most sensitive to gold production and gold price. Sensitivity on operating and capital cost is closely matched, with the Project being only slightly more sensitive to capital costs.

Table 24‑25: Project Sensitivity

Parameter | -15% | -10% | -5% | 0% | 5% | 10% | 15%
Market Prices | $184,117 | $262,198 | $340,195 | $417,938 | $495,627 | $573,270 | $650,714
Opex | $522,743 | $487,855 | $452,900 | $417,938 | $382,848 | $347,842 | $312,659
Capex | $536,411 | $496,920 | $457,429 | $417,938 | $378,560 | $339,376 | $300,353

Figure 24‑11: Project Sensitivity

Table 24‑26: Alternate Case Sensitivity to Gold Price versus Foreign Exchange Rate (US$:AUD)

| GOLD PRICE (US$/oz-Au)
$1,000 | $1,100 | $1,200 | $1,300 | $1,400 | $1,500 | $1,600
FEX Rate | IRR | NPV5 | IRR | NPV5 | IRR | NPV5 | IRR | NPV5 | IRR | NPV5 | IRR | NPV5 | IRR | NPV5
0.65 | 12.1% | $ 202 | 16.2% | $ 322 | 20.1% | $ 442 | 23.8% | $ 561 | 27.4% | $ 680 | 30.9% | $ 800 | 34.3% | $ 919
0.70 | 10.2% | $ 154 | 14.2% | $ 274 | 18.1% | $ 394 | 21.7% | $ 514 | 25.2% | $ 633 | 28.6% | $ 752 | 31.9% | $ 871
0.75 | 8.4% | $ 106 | 12.4% | $ 226 | 16.1% | $ 346 | 19.7% | $ 466 | 23.1% | $ 585 | 26.5% | $ 704 | 29.7% | $ 824
0.80 | 6.8% | $ 57 | 10.6% | $ 178 | 14.3% | $ 298 | 17.8% | $ 418 | 21.2% | $ 537 | 24.5% | $ 657 | 27.6% | $ 776
0.85 | 5.2% | $ 7 | 9.0% | $ 132 | 12.6% | $ 250 | 16.0% | $ 370 | 19.3% | $ 490 | 22.6% | $ 609 | 25.7% | $ 728
0.90 | 2.8% | $ (74) | 7.3% | $ 81 | 10.9% | $ 202 | 14.3% | $ 322 | 17.6% | $ 442 | 20.8% | $ 561 | 23.8% | $ 681
0.95 | 0.5% | $ (157) | 5.9% | $ 32 | 9.4% | $ 155 | 12.7% | $ 274 | 16.0% | $ 394 | 19.0% | $ 514 | 22.0% | $ 633
1.00 | -1.8% | $ (238) | 3.8% | $ (45) | 7.8% | $ 105 | 11.2% | $ 227 | 14.4% | $ 346 | 17.4% | $ 466 | 20.4% | $ 585

Alternate Case uses a foreign exchange rate of US$0.80:AUD1.00 and a gold price of $1,300/oz-Au.

Table 24‑27: Alternate Case Sensitivity to Gold Prices versus NPV Discount Rate

| GOLD PRICE (US$/oz-Au)
Discount Rate | $ 1,000 | $ 1,100 | $ 1,200 | $ 1,300 | $ 1,400 | $ 1,500 | $ 1,600
5% | $ 57 | $ 178 | $ 298 | $ 418 | $ 537 | $ 657 | $ 776
8% | $ (34) | $ 70 | $ 172 | $ 273 | $ 373 | $ 473 | $ 572
10% | $ (80) | $ 15 | $ 106 | $ 196 | $ 286 | $ 375 | $ 464

Foreign Exchange is held constant at US$0.80:AUD1.00

25.0INTERPRETATION AND CONCLUSIONS

25.1Project Risks

Significant risks and uncertainties that could reasonably affect the reliability or confidence in the Project outcome are provided in Table 25‑1. The Project is an advanced-staged development project that has undergone engineering and permitting for a number of years. As such, the development risks that are within the control of Vista are considered low to moderate.

Table 25‑1: Project Risks

Risk | Description | Probability | Severity
Gold Price | The Project economics are sensitive to gold price. Sustained downward gold price trends could render the project uneconomic. | Low-Medium | High
Foreign Exchange | The Project capital and operating costs are sensitive to foreign exchange changes. A strengthen Australian dollar without an offsetting positive change in the gold price could render the Project uneconomic. | Low-Medium | High
Political Setting | Australia and the Northern Territory have historically been supportive of the extractive industries. Changes in legislation could have a negative impact on the project. | Low | Medium
Jawoyn | The JAAC is supportive of the Project. Changes in Vista’s relationship with JAAC could have social impacts on the Project. | Low | Low-Medium
Permitting & Regulatory Approvals | The Project has received both EIS and EPBC authorizations. | Low | Medium
Property Holdings | Vista has secured the Mt Todd concession holdings as described in Section 4.0. Any change could have negative impacts to the Project. | Low | Low
Infrastructure | The Project relies on the use of existing infrastructure. The condition of which is well known and is functional. Significant deficiencies would result in increased capital expense. | Low | Low
Understanding of Resource | The Project viability relies upon historical drilling as well as recent drilling to develop and assess the resource model. New drill results could adversely affect the interpretation of parts of the deposit, with impacts to resources and production estimates. | Low | Low
Power Plant Estimated Capital | The proposed power plant utilizes industry standard equipment that is currently in use in Australia. Changes in cost could affect Project economics. | Low | Low-Medium
Reagents & Consumables | The process operating costs are sensitive to global changes in reagents and consumables pricing. | Medium | Medium
Fuel | The Project operating costs are sensitive to global changes in prices for diesel and natural gas. | Medium | Medium
Mobile Equipment Capital | Mobile equipment prices are an important part of the Project capital. Significant increases could impact the Project economics. | Low | Low-Medium

Process Technology | Extensive testing has been completed to identify the most suitable technology and equipment in the process are. The performance of the selected equipment could negatively impact Project economics. | Low-Medium | Medium
Climatic Events | Day to day mining operations could be significantly impacted by either high precipitation events. | Low | Low-Medium
Groundwater | Day to day mining operations could be impacted by groundwater inflow. | Low | Low
Water Treatment | Heavy and sustained rains could result in water treatment in excess of capacity for short periods. | Low-Medium | Medium
Existing TSF 1 | Restarting of TSF 1 operations is an integral part of the Project plan. This facility has been idle for many years, delays could impact the schedule. | Low | Low-Medium
Reclamation & Closure | There is potential for reclamation activities to extend beyond the active planned closure period, and therefore generate greater sustaining costs. Additional risk lies should the closure design not perform as intended. | Low | Medium

25.2Geology and Resources

The Project is situated within the southeastern portion of the Early Proterozoic Pine Creek Geosyncline which is comprised of the Burrell Creek Formation, the Tollis Formation, and the Kombolgie Formation.

Gold mineralization in this area is constrained to a single mineralization event and the deposits are classified as orogenic gold deposits in the subdivision of thermal aureole gold style. The Batman deposit has characteristics of an intrusion related gold system making it the primary resource.

The Batman deposit is defined by approximately 7.4 million ounces (Moz) of gold within 278 Mt of measured and indicated resource at an average grade of 0.82 g-Au/t and a cutoff grade of 0.4 g-Au/t as provided in Table 14‑1.

25.3Mineral Reserve and Mine Planning

Pit designs were completed based on Whittle TM pit optimizations and are appropriate for metal prices of approximately US$800 per ounce Au for the Alternate Case and US$1,000 per ounce Au for the Base Case. The Mt Todd proven and probable reserves have been defined using economics based on a gold price of US$1,250 per ounce and an elevated cutoff grade of 0.40 g-Au/t. The proven and probable reserves were used to create a production schedule for mining, and a positive cash-flow analysis has been done based on the production schedule by Tetra Tech. The reserves have reasonable economics with respect to the statement of reserves under NI 43-101 regulations.

Mine production constraints were imposed to ensure that mining wasn’t overly aggressive with respect to the equipment anticipated for use at Mt Todd. The schedule has been produced using mill targets and stockpiling strategies to enhance the project economics. The

constraints and limits are reasonable to support the project economics which are used to justify the statement of reserves.

Pit designs use six-meter benches for mining. This corresponds to the resource model block heights, and MDA believes this to be reasonable with respect to dilution and equipment anticipated to be used in mining. In areas where the material is consistently ore or waste so that dilution is not an issue, benches may be mined in 12-m heights.

25.4Mineral Processing

The substantial quantity and quality of metallurgical test work data developed from Mt Todd drill core samples has led to the development of a robust energy efficient comminution circuit followed by a standard gold recovery process. Key conclusions drawn from the metallurgy studies are:

Mt Todd (Batman) ore is among the hardest and most competent ore types processed for mineral recovery. The most energy efficient comminution circuit has been determined to be the sequence of primary crushing, closed circuit secondary crushing, and closed circuit HPGR tertiary crushing followed by ball milling.

The ore is free-milling, is not preg-robbing, and is amenable to gold extraction by conventional cyanidation processes.

The ore has relatively high specific cyanide consumption. This is largely due to the presence of sulfides and cyanide consuming copper.

The use of sorting has helped to decrease operational costs and remove portions of the harder rock mined.

The equipment selection criteria for the Base Case operation has received considerable interaction with specialist vendors to the point where there is a reasonably high degree of confidence in selected technology and process units at this preliminary feasibility study stage. The recommended flowsheet for FS consists of primary crushing, closed circuit secondary crushing, closed circuit tertiary crushing using HPGRs, ball milling, cyclone classification, pre-leach thickening, leach and adsorption, elution electrowinning and smelting, carbon regeneration, tailings detox and disposal to conventional tailings storage facility.

25.5Infrastructure

25.5.1Site Preparation

Bulk earthworks are designed to minimize the import of fill materials.

25.5.2Support Buildings

Administration offices, gatehouse/security facilities, cribs/ablutions are planned to be transportable buildings.

The process plant offices, workshop and warehouse are located inside the existing Flotation Building.

Sample preparation and laboratory will have a purpose-built steel shed.

25.5.3Access Roads Parking and Laydown

The access road is based on the repaired existing road.

25.5.4Heavy Lifts

Heavy cranage is allowed for all lifts greater than 50 t.

25.5.5Bulk Transport

All bulk transport will be weighed.

25.5.6Communications

Site-wide communication is based on a 50 m tall communication tower that will support eight (8) channels.

25.6Environmental and Social Conclusions

25.6.1Existing Body of Work

A number of environmental studies have been conducted at the Project Site in support of development of Environmental Impact Statements and as required for environmental and operational permits. Studies conducted have investigated soils, climate and meteorology, geology, geochemistry, biological resources, cultural and anthropological sites, socio-economics, hydrogeology, and water quality.

25.6.2Environmental Impact Study and Approvals

The Environmental Impact Study (“EIS”) was submitted in June 2013. The NT Environmental Protection Authority provided its final assessment of the Project in June 2014. Notification of approval of the EIS was given September 2014.

Vista has received all major environmental approvals to proceed with the Project.

25.6.3Social or Community Impacts

The Jawoyn people have strong involvement in the planning of the Project. Areas of aboriginal significance have been designated, and the mine plan has avoided development in these restricted works areas.

25.7Results of the Site-wide Water Balance Model

The WTP rate of 500 m 3/hr and PWP sizing for 6 days of storage was determined to be appropriate for the 50,000 tpd production process water requirements.

The greatest amount of make-up water required from the RWD was quantified as 11,955 m 3/day. RWD requirements were found to be the most dependent upon TSF decant volumes.

The WRD retention pond (RP1) was typically observed to overtop less than 1% of the time during the 13 year simulation.[1] LGRP storage may be optimized.

[1]A typical value is given. Separate model runs provide a range of overtopping events, due to the stochastic nature of the model.

26.0RECOMMENDATIONS

The following recommendations are presented for consideration and were developed using best engineering judgement to estimate the costs of the respective recommendations.

26.1Feasibility Study

A FS is recommended to advance the Project to a place where any additional detailed information necessary provides support of capital and operating cost estimates which lead to a potential project development decision.

The estimated budget for the FS is approximately US$2,500,000.

26.2Resource and Exploration

The Batman deposit potentially extends along strike both to the north and south. Step out drilling should be used to explore this.

Additional deep infill drilling should be used to help define potential deep mineralization not detected by earlier drillholes.

Infill drilling within and exploration drillholes along the trend of the Quigley deposit is recommended.

Additional drilling exploring the exploration licenses, following up on geophysical and geochemical anomalies.

The estimated budget for drilling within the mining licenses is US$500,000-1,000,000 and US$500,000-1,000,000 for initial drilling on the exploration licenses.

26.3Mining Risks and Opportunities

26.3.1Opportunities

Some refinement of dump designs for the chosen ultimate pit may help to reduce the overall footprint of the resulting dumps and therefore reduce closure costs.

Current blasting patterns have been tightened up to reduce oversize. With experience, the blasting patterns can be optimized to reduce both drilling and blasting costs.

26.3.2Risks

Large stockpiles of low-grade ore are used for reasonably long periods of time. These stockpiles may oxidize during storage and the resulting recoveries from these stockpiles may be overstated due to the change of chemistry.

26.4Environmental Studies

Additional studies will be needed to further assess environmental baseline conditions to support feasibility level design, permitting, and closure planning for the Project, including:

Erosion analyses;

Waste and cover material hydraulic properties characterization and analysis;

Acid-base accounting on waste rock and tailings;

Ongoing aquatic, benthic and wildlife studies;

Comprehensive vegetation survey;

Archaeological and historical assessments for all areas to be disturbed;

Hydrogeologic investigations and site-wide hydrogeologic characterization; and

Precipitation, stream flow, and watershed data.

The estimated budget for this work is US$350,000.

26.5Site-wide Water Balance

Recommended model improvements include:

The site-wide water balance model is dependent upon the TSF water balance model, which provides decant water to the process facility, and the vadose and seepage models, which characterize seepage through the various rock piles on site (WRD, LGOS, and HLP). As such, completion of these models to the greatest detail practicable affects the overall quality of the site-wide water balance model results.

Incorporate results of future Batman Pit potential groundwater inflow investigation.

Optimize management of Batman Pit dewatering effluent and other contact water. This water may be of sufficient quality to be used as make-up water to the process circuit offsetting the water coming from the RWD. This water would also reduce the amount of water ultimately requiring treatment.

In this iteration of the site-wide water balance model, the entirety of the WTP effluent is being used as dust suppression around the mine site during the dry season. Further investigation of other uses of the WTP effluent should be conducted.

Further investigation of the adequacy of RP1 storage capacity is recommended, particularly within the early stages of the LoM when a larger fraction of the catchment reports to this pond.

Incorporate RWD stage-storage relationship and catchment area into the site-wide water balance model such that it may be modeled as a reservoir, as opposed to an infinite source.

Inclusion of process, fire, potable and raw water tanks. At present, only the dust suppression tank is modeled. The tanks above are currently modeled as drawing water directly from the RWD, rather than demands on discrete tanks.

Review and update dust suppression requirements.

Incorporate evolving Batman Pit shell geometry to more accurately model that facility.

The estimated budget for this work is US$200,000.

26.6Groundwater Hydrology and Mine Dewatering

The following work is recommended with respect to groundwater hydrology and mine dewatering:

Additional hydrogeologic study should be completed in the vicinity of the Batman Pit to provide more detailed information on which to base calibration of the regional groundwater flow model and subsequent prediction of groundwater inflows to the pit and post-mining recovery of the groundwater system. The study should include permeability testing by packer test methods in three to five borings within or immediately adjacent to the pit and drilled to the proposed ultimate depth of the pit. The packer testing in those borings should be conducted over relatively short (10 m to 30 m) intervals, beginning at the water table and extending to the total depth of the borings. This testing could be combined with exploration drilling by performing the testing in exploration core holes. The study should also include measurement of depth to water in any accessible existing borings or core holes within or immediately adjacent to the pit.

Calibration of the regional groundwater flow model should be completed with the additional data, and the calibrated model should be used to refine the estimates of groundwater inflow to the pit and predictions of the hydrogeologic effects of pit dewatering.

The post-mining version of the groundwater flow model should be updated with the calibrated model used as its basis. Output from the post-mining model should be incorporated into any geochemical modeling of post-mining pit lake formation and geochemistry.

The estimated budget for this work is US$400,000.

26.7Process Plant Geotechnical Investigation Recommendations

For the DFS future geotechnical work is suggested in the following areas, and is estimated at US$150,000.

26.7.1Grinding

This is a large vibrating structure which should be founded in rock rather than on fill to reduce dynamic effects. An accurate rock level is required to confirm foundation design and accurately estimate required concrete quantities and rock excavation. Existing structure concrete slabs are located directly over the new mill location, so it is recommended that additional new test pits around all four sides be undertaken to allow interpolation.

26.7.2Leach/CIP

Large tanks to be constructed in this area, with high foundation bearing pressure. Variation in rock level will impact on differential settlement which needs to be considered. Recommend at least four additional test pits to evaluate variance in rock levels in north-south and east-west directions.

26.7.3Stockpile & Reclaim

No test pit data anywhere near this area (nearest test pit is more than 200m away). Steeply sloping ground surface exists (in excess of 5 m variation in ground level across the reclaim tunnel) so there may be considerable variation in rock levels which will effect potential settlement due to stockpile surcharge and required excavation for the concrete vault and tunnel. Recommend a new borehole on the high side (uncertain whether this high side bench is fill material, and therefore whether rock would be encountered within the limits of an excavator for a test pit) and a new test pit on low side, to determine rock levels.

26.7.4Coarse Screening

Vibrating structure which should be founded on rock rather than on fill to reduce dynamic effect. An accurate rock level is required to confirm foundation design and accurately estimate required concrete quantities and rock excavation. Existing test pits are on one side only so unable to interpolate between existing pits, the nearest existing test pit is approximately 100m to the north so recommend a new test pit at this location.

26.7.5HPGR

Vibrating structure which should be founded on rock rather than on fill to reduce dynamic effect. Therefore need to know rock level to confirm foundation design and accurately estimate required concrete quantities and rock excavation. Nearest existing test pit is approximately 70m away so recommend new test pit at this location.

26.8Tailings Facility Design

The following studies and investigations are recommended to support TSF design for future phases of the project:

26.8.1TSF Construction Schedule

The TSF staging and construction schedule must be optimized during the feasibility study phase of the project to minimize mobilization costs.

26.8.2Investigation of TSF1 Drainage Features

The condition of the existing toe drains, underdrains, and decant towers must be investigated to confirm their condition prior to re-commissioning of TSF1.

26.8.3Geotechnical Investigation and Assessment

Additional geotechnical investigations are recommended, including drilling, sampling, and laboratory testing to characterize the geotechnical properties of the surface and subsurface materials. The geotechnical drilling program should include Standard Penetration Testing (SPT) and/or Cone Penetration Testing (CPT). CPT investigation within the existing TSF1 tailings is recommended to support assessment of the raise foundation strength and liquefaction analysis. A program of laboratory geotechnical testing of a sample of representative future tailings is recommended as part of future design work, including index testing (particle size, plasticity, specific gravity), compaction testing, and advanced testing to assess consolidation, strength and permeability after compaction, and unsaturated soil characteristics.

26.8.4Waste Rock Testing

Additional laboratory testing of the waste rock is recommended, including, but not limited to, proctor compaction, hydraulic conductivity, and shear strength testing. No testing of the run-of-mine waste was conducted for this study, and a representative range of strength parameters would improve predictions of behavior related to slope stability.

26.8.5Consolidation/Seepage Modeling

The seepage and stability analyses discussed in this report were based on laboratory test conducted on the in-situ tailings. Large scale consolidation tests should be conducted on bench scale samples of the proposed process tailings to determine hydraulic conductivities as a function of effective stress. The seepage and stability analyses must be updated based on these representative material properties. Additionally, seepage and stability analyses must be conducted assuming that the underdrain system is plugged to assess its influence on the stability of future vertical expansions.

26.8.6Water Balance

A TSF water balance analysis which includes the return water ponds must be conducted to predict the volume of recycle water available for process operations. The results of the water balance analysis will also be used to size the return water system which includes the existing decant structures and an additional barge pump.

26.8.7TSF Consequence Classification

A formal dam classification assessment is recommended as part of future design work. Under ANCOLD guidelines, the Consequence Category is determined for a potential dam failure and environmental spill. The classification should include a failure modes assessment, a dam break and inundation study, and consideration of the potential for impacts to business, social and environment, and the potential for loss of life. The resulting Consequence Category for the tailings storages will be used to identify parameters for advancing the design such as the design storm event and the design seismic event.

26.9Process Operating Costs

Two major items incurring operating costs are grinding media and reagents. Together these items make up 71% of the plant consumables operating costs. The FS should investigate options for reducing the consumption rate and the unit costs for these consumables.

This work is included in the estimated budget for the FS.

26.10Geochemical Analyses

Geochemical characterization will be updated to reflect the designations of Potentially Acid Forming, Potentially Acid Forming-Low Capacity, Non Acid Forming, Acid Consuming and Uncertain in accordance with DITR (2007) guidelines. Additionally, Tetra Tech recommends performing geochemical testing on the sorter reject material.

This work is included in the estimated budget for the FS.

26.11Process Parameter Optimization

The on-going testwork is directed at optimization of process parameters which is expected to result in both capital and operating cost reductions as detailed in this Technical Report.

27.0REFERENCES

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Ausenco, August 2009. JKSimMet Circuit Simulations for the 11 mt Vista Gold Mt Todd Plant.

Australian Government, Standing Council on Environment and Water and the National Health and Medical Research Council, 2011. National Water Quality Management Strategy - Australian Drinking Water Guidelines 6 (2011) – Version 3.4 Updated October 2017.

Barkley, Ross, March 24, 2016, Review of Geotechnical Pit Slope Design for the Project, Call & Nicholas, Inc. internal memorandum.

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Gibbs, D.R., Horton, J., August 1990. Analysis of Bias between Drilling Techniques and Drill hole Orientations Used at Batman, Mt Todd, NT, Using The ‘Preferred’ Gold Assay Database, Report No. 08.5116.

Gibbs, D.R., Horton, J., Pantalone, D., June 1990. Preliminary Analysis of Bias between Drilling Techniques Used at the Batman Deposit, N.T. Using the Original Gold Assay Database, Report No. 08.4449.

Gibbs, Duncan, July, 1990. Corrections to the Batman Assay Database and the Impact of Preferred and Bias Corrections, Report No. 08.5117.

Golder Associates, 2011. Mt Todd Gold Project: Batman Pit Slope Design Guidance in Support of the Definitive Feasibility Study, September 13, 2011.

Gustavson Associates, LLC, December 2006. Preliminary Economic Assessment, Mt Todd Gold Project, NT, Australia.

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Gustavson Associates, LLC, January 2008. Preliminary Assessment Awak Mas Gold Project.

Hein, K.A.A., 2003. The Batman and Quigleys gold deposits of the Mt. Todd (Yimuyn Manjerr) Mine, Australia: Structural, petrographic and mineralogical investigations of coeval quartz sulphide vein and lode/stockwork systems. Ore Geology Reviews 23(1-2), July 2003.

Hein, K.A.A., Zaw, K. and Mernagh, T.P. (2006) Linking mineral and fluid inclusion paragenetic studies: The Batman deposit, Mt. Todd (Yimuyn Manjerr) goldfield, Australia. Ore Geology Reviews, 28, 180-200.JK Tech. Pty. Ltd, August 2009. Comminution Test Report on Five Samples from Mt Todd Mine.

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Mineral Resources Development Property Evaluators, Developers, and Consulting Geologists and Engineers, September 1995. Zapopan NL Sampling and Reconciliation Study of the Mount Todd Gold Mine.

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28.0CERTIFICATE OF QUALIFIED PERSON

28.1Qualifications of Consultants

The Consultants preparing this Technical Report are specialists in the fields of geology, exploration, mineral resource and mineral reserve estimation and classification, underground mining, geotechnical, environmental, permitting, metallurgical testing, mineral processing, processing design, capital and operating cost estimation, and mineral economics.

None of the Consultants or any associates employed in the preparation of this report has any beneficial interest in Vista. The Consultants are not insiders, associates, or affiliates of Vista. The results of this Technical Report are not dependent upon any prior agreements concerning the conclusions to be reached, nor are there any undisclosed understandings concerning any future business dealings between Vista and the Consultants. The Consultants are being paid a fee for their work in accordance with normal professional consulting practice.

The following individuals, by virtue of their education, experience and professional association, are considered qualified persons (QP) as defined in the NI 43-101 standard, for this report, and are members in good standing of appropriate professional institutions.

This Technical Report was prepared by the following QPs, Certificates and consents of which are contained herein:

Name | Title, Company | Responsible for Sections
Rex Clair Bryan, Ph.D. | Principal Geostatistician Tetra Tech, Inc. | Sections 1.4, 1.5, 1.15.2, 6, 6.1, 6.2, 6.3, 6.4, 7, 8, 9, 10, 11, 12, 14, 24.7.1, 25.2, and 26.2
Anthony Clark, P.E. | Area Lead Mechanical Engineer POWER Engineers, Inc. | Sections 18.8, 21.1.3, and 21.2.4
Thomas L. Dyer, P.E. | Mining Engineer Mine Development Associates | Sections 1.6, 1.7, 1.15.3, 15, 15.1, 15.2, 15.3, 15.4, 15.5, 15.6, 16, 21.1.1, 21.2.1, 24.7.2, 25.3, and 26.3
Amy L. Hudson, Ph.D., CPG, REM | Principal Hydrogeologist/ Geochemist Tetra Tech, Inc. | Sections 1.15.12, 24.3, and 26.10
Chris Johns, M.Sc., P.Eng | Geological Engineer Tetra Tech, Inc. | Sections 1.15.10, 18.2.4, 21.1.8, 21.2.6, and 26.8
Benjamin S. Johnson, P.E. | Water Treatment Plant Process Engineer Tetra Tech, Inc. | Sections 21.1.6, 21.2.5, 24.2.2.1, and 24.2.2.2
Deepak Malhotra, Ph.D. | Principal Metallurgist Resource Development Inc. | Sections 1.6.1 & 15.7
Zvonimir Ponos, BE, MIEAust, CPeng, NER | General Manager Engineering Coffey Services Australia Pty Ltd (trading as Tetra Tech Proteus) | Sections 1.10, 1.15.5, 18.1, 18.2, 18.2.1, 18.2.2, 18.2.3, 18.2.5, 18.2.6, 18.3, 18.4, 18.5, 18.6, 18.7, 21.2.7, 24.6, 24.7.4, and 25.5

Guy Roemer, P.E. | Environmental Engineer Tetra Tech, Inc. | Sections 1.15.7, 24.2.1, 24.7.5, 25.7 and 26.5
Vicki Scharnhorst, P.E., LEED AP | Principal Tetra Tech, Inc. | Sections 1.1, 1.2, 1.3, 1.12, 1.15.1, 1.15.6, 2, 3, 4, 5, 18, 20, 21.1.5, 21.1.7, 23, 24, 24.2, 24.2.2, 24.2.2.3, 24.2.2.4, 24.2.2.5, 24.4, 24.7, 25.1, 25.6, 26, 26.1, and 26.4
D. Erik Spiller, QP | Principal Metallurgist Tetra Tech, Inc. | Sections 1.8, 1.9, 1.15.4, 1.15.11, 6.5, 6.6, 13, 17, 21.1.2, 21.2.3, 24.7.3, 25.4, 26.9, and 26.11
Jessica I. Spriet, P.E. | Professional Engineer Tetra Tech, Inc. | Sections 1.11, 1.13, 1.14, 1.15.9, 19, 21, 21.1, 21.2, 22, 24.1, 24.7.6, 24.7.7, 24.7.8, and 26.7
Keith Thompson, CPG, PG | Professional Geologist Tetra Tech, Inc. | Sections 1.15.8, 21.1.4, 21.2.2, 24.5, 26.6

28.2Table of Responsibility

QPs are responsible for all subsections listed beneath headings unless subsections are detailed below.

Section No | Section Name | QP
1 | SUMMARY (No Intro) | N/A
1.1 | Introduction | Scharnhorst, Vicki
1.2 | Location | Scharnhorst, Vicki
1.3 | Property Description | Scharnhorst, Vicki
1.4 | Geology and Mineralization | Bryan, Rex
1.5 | Mineral Resource Estimate | Bryan, Rex
1.6 | Mineral Reserve Estimates | Dyer, Tom
1.6.1 | Heap Leach Reserve Estimate | Malhotra, Deepak
1.7 | Mining Methods | Dyer, Tom
1.8 | Metallurgy | Spiller, Erik
1.9 | Mineral Processing | Spiller, Erik
1.10 | Project Infrastructure | Ponos, Zvonimir
1.11 | Market Studies and Contracts | Spriet, Jessica
1.12 | Social and Environmental Aspects | Scharnhorst, Vicki
1.13 | Capital and Cost Estimates | Spriet, Jessica
1.14 | Financial Analysis | Spriet, Jessica
1.15 | Conclusions and Recommendations (No Intro) | N/A

1.15.1 | Feasibility Study | Scharnhorst, Vicki
1.15.2 | Geology and Resources | Bryan, Rex
1.15.3 | Mineral Reserve and Mine Planning | Dyer, Tom
1.15.4 | Mineral Processing | Spiller, Erik
1.15.5 | Infrastructure | Ponos, Zvonimir
1.15.6 | Environmental and Social Impacts | Scharnhorst, Vicki
1.15.7 | Results of the Site-wide Water Balance Model | Roemer, Guy
1.15.8 | Groundwater Hydrology and Mine Dewatering | Thompson, Keith
1.15.9 | Process Plant Geotechnical Investigation | Spriet, Jessica
1.15.10 | TSF Design | Johns, Chris
1.15.11 | Process | Spiller, Erik
1.15.12 | Geochemical Analyses | Hudson, Amy
2 | INTRODUCTION | Scharnhorst, Vicki
3 | RELIANCE ON OTHER EXPERTS | Scharnhorst, Vicki
4 | PROPERTY DESCRIPTION AND LOCATION | Scharnhorst, Vicki
5 | ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY | Scharnhorst, Vicki
6 | HISTORY | Bryan, Rex
6.1 | History of Previous Exploration | Bryan, Rex
6.2 | Historic Drilling | Bryan, Rex
6.3 | Historic Sampling Method and Approach | Bryan, Rex
6.4 | Historic Sample Preparation, Analysis, and Security | Bryan, Rex
6.5 | Historic Process Description | Spiller, Erik
6.6 | Technical Problems with Historical Process Flowsheet | Spiller, Erik
7 | GEOLOGICAL SETTING AND MINERALIZATION | Bryan, Rex
8 | DEPOSIT TYPES | Bryan, Rex
9 | EXPLORATION | Bryan, Rex
10 | DRILLING | Bryan, Rex
11 | SAMPLE PREPARATION, ANALYSES AND SECURITY | Bryan, Rex
12 | DATA VERIFICATION | Bryan, Rex
13 | MINERAL PROCESSING AND METALLURGICAL TESTING | Spiller, Erik
14 | MINERAL RESOURCE ESTIMATES | Bryan, Rex
15 | MINERAL RESERVES | Dyer, Tom
15.1 | Pit Optimization | Dyer, Tom
15.2 | Pit Designs | Dyer, Tom
15.3 | Cutoff Grade | Dyer, Tom

15.4 | Dilution | Dyer, Tom
15.5 | Reserves and Resources | Dyer, Tom
15.6 | In-Pit inferred Resources | Dyer, Tom
15.7 | Heap Leach Reserve Estimate | Malhotra, Deepak
16 | MINING METHODS | Dyer, Tom
17 | RECOVERY METHODS | Spiller, Erik
18 | PROJECT INFRASTRUCTURE | Scharnhorst, Vicki
18.1 | Facility 2000 – Mine | Ponos, Zvonimir
18.2 | Facility 4000 – Project Services | Ponos, Zvonimir
18.2.1 | Area 4100 – Water Supply | Ponos, Zvonimir
18.2.2 | Area 4200 – Power Supply | Ponos, Zvonimir
18.2.3 | Area 4300 – Communications | Ponos, Zvonimir
18.2.4 | Area 4400 – Tailings Dam | Johns, Chris
18.2.5 | Area 4500 – Waste Disposal | Ponos, Zvonimir
18.2.6 | Area 4600 – Plant Mobile Equipment | Ponos, Zvonimir
18.3 | Facility 5000 – Project Infrastructure | Ponos, Zvonimir
18.4 | Facility 6000 – Permanent Accommodation | Ponos, Zvonimir
18.5 | Facility 7000 – Site Establishment and Early Works | Ponos, Zvonimir
18.6 | Facility 8000 – Management, Engineering, EPCM Services | Ponos, Zvonimir
18.7 | Facility 9000 – Preproduction Costs | Ponos, Zvonimir
18.8 | Electric Power Plant | Clark, Anthony
19 | MARKET STUDIES AND CONTRACTS | Spriet, Jessica
20 | ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT | Scharnhorst, Vicki
21 | CAPITAL AND OPERATING COSTS | Spriet, Jessica
21.1 | Capital Cost | Spriet, Jessica
21.1.1 | Mining (MDA) | Dyer, Tom
21.1.2 | CIP Process and Infrastructure (Proteus) | Spiller, Erik
21.1.3 | Power Plant (POWER Engineers) | Clark, Anthony
21.1.4 | Mine Dewatering (Tetra Tech) | Thompson, Keith
21.1.5 | Reclamation and Closure (Tetra Tech) | Scharnhorst, Vicki
21.1.6 | Water Treatment Plant (Tetra Tech) | Johnson, Benjamin
21.1.7 | Raw Water Dam (Tetra Tech) | Scharnhorst, Vicki
21.1.8 | Tailings Storage Facilities (Tetra Tech) | Johns, Chris
21.2 | Operating Costs | Spriet, Jessica
21.2.1 | Mining (MDA) | Dyer, Tom

21.2.2 | Mine Dewatering (Tetra Tech) | Thompson, Keith
21.2.3 | CIP Process and G&A (Proteus) | Spiller, Erik
21.2.4 | Power Plant (POWER Engineers) | Clark, Anthony
21.2.5 | Water Treatment Plant (Tetra Tech) | Johnson, Benjamin
21.2.6 | Tailings Storage Facilities (Tetra Tech) | Johns, Chris
21.2.7 | General & Administrative | Ponos, Zvonimir
22 | ECONOMIC ANALYSIS | Spriet, Jessica
23 | ADJACENT PROPERTIES | Scharnhorst, Vicki
24 | OTHER RELEVANT DATA AND INFORMATION | Scharnhorst, Vicki
24.1 | Process Plant Geotechnical | Spriet, Jessica
24.2 | Water Management | Scharnhorst, Vicki
24.2.1 | Site-wide Water Balance | Roemer, Guy
24.2.2 | Wet Infrastructure | Scharnhorst, Vicki
24.2.2.1 | Water Treatment Plant | Johnson, Benjamin
24.2.2.2 | Water Quality Standards for Waste Water Discharge | Johnson, Benjamin
24.2.2.3 | Raw Water Reservoir and Pipeline | Scharnhorst, Vicki
24.2.2.4 | Potable Water | Scharnhorst, Vicki
24.2.2.5 | Sanitary Sewer System | Scharnhorst, Vicki
24.3 | Geochemistry | Hudson, Amy
24.4 | Surface Water Hydrology | Scharnhorst, Vicki
24.5 | Regional Groundwater Model and Mine Dewatering | Thompson, Keith
24.6 | Project Implementation | Ponos, Zvonimir
24.7 | Alternate Case | Scharnhorst, Vicki
24.7.1 | Mineral Resource | Bryan, Rex
24.7.2 | Mining | Dyer, Tom
24.7.3 | Process Facility | Spiller, Erik
24.7.4 | Infrastructure | Ponos, Zvonimir
24.7.5 | Site-wide Water Balance Model | Roemer, Guy
24.7.6 | Capital Costs, Alternate Case | Spriet, Jessica
24.7.7 | Operating Costs, Alternate Case | Spriet, Jessica
24.7.8 | Economic Results, Alternate Case | Spriet, Jessica
25 | INTERPRETATION AND CONCLUSIONS (No Intro) | N/A
25.1 | Project Risks | Scharnhorst, Vicki
25.2 | Geology and Resources | Bryan, Rex
25.3 | Mineral Reserve and Mine Planning | Dyer, Tom

25.4 | Mineral Processing | Spiller, Erik
25.5 | Infrastructure | Ponos, Zvonimir
25.6 | Environmental and Social Conclusions | Scharnhorst, Vicki
25.7 | Results of the Site-wide Water Balance Model | Roemer, Guy
26 | RECOMMENDATIONS | Scharnhorst, Vicki
26.1 | Feasibility Study | Scharnhorst, Vicki
26.2 | Resource and Exploration | Bryan, Rex
26.3 | Mining Risks and Opportunities | Dyer, Tom
26.4 | Environmental Studies | Scharnhorst, Vicki
26.5 | Site-wide Water Balance | Roemer, Guy
26.6 | Groundwater Hydrology and Mine Dewatering | Thompson, Keith
26.7 | Process Plant Geotechnical Investigation Recommendations | Spriet, Jessica
26.8 | Tailings Facility Design | Johns, Chris
26.9 | Process Operating Costs | Spiller, Erik
26.10 | Geochemical Analyses | Hudson, Amy
26.11 | Process Parameter Optimization | Spiller, Erik
27 | REFERENCES | N/A
28 | CERTIFICATE OF QUALIFIED PERSON | N/A
28.1 | Qualifications of Consultants | N/A
28.2 | Table of Responsibility | N/A

CERTIFICATE OF QUALIFIED PERSON

Rex Clair Bryan, Ph.D.

Principal Geostatistician

Tetra Tech, Inc.

350 Indiana Street, Suite 500

Golden, Colorado 80401

Telephone: (303) 217-5700

Facsimile: (303) 217-5705

Email: Rex.Bryan@tetratech.com

This certificate applies to the report entitled: “NI 43-101 Technical Report – Mt Todd Gold Project – 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia” (“Technical Report”), effective date January 24, 2018, issued on March 2, 2018.

I, Rex Clair Bryan, Ph.D., do hereby certify that:

1)I am a Senior Geostatistician with Tetra Tech, Inc. with a business address at 350 Indiana Street, Suite 500, Golden, Colorado USA, 80401.

2)I graduated with a Ph.D. degree in 1980 from the Colorado School of Mines, Golden Colorado, USA. In addition, I graduated with a degree MSc. In Geology in 1976 from the Brown University, Providence, Rhode Island, USA. I have worked as a Geostatistician for a total of 38 years since my graduation. My relevant experience is in the areas of resources and reserve reporting. I am a Competent/Qualified Person (QP), with the Society of Mining Engineers in Colorado, USA (SME Registered Member #411340).

3)I have read the definition of “qualified person” set out in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” within the meaning of NI 43-101.

4)I have personally visited and inspected the property which is the subject of the Technical Report on June 28 th and June 29 th, 2017 for two days. In addition, I have visited and inspected the property September 12 th, 2011 to September 14 th, 2011 and February 6 th, 2013 to February 8 th, 2013.

5)I am responsible for Sections 1.4, 1.5, 1.15.2, 6, 6.1, 6.2, 6.3, 6.4, 7, 8, 9, 10, 11, 12, 14, 24.7.1, 25.2, and 26.2 of the Technical Report.

6)I am independent of the issuer, Vista Gold Corp., according to Section 1.5 of NI 43-101.

7)I have had prior involvement with the property that is the subject of the Technical Report. My involvement has consisted of acting as an expert who was relied upon for the NI 43-101 Technical Report – Mt Todd Gold Project, 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia, Amended & Restated; July 7, 2014.

8)I have read NI 43-101, Form 43-101F1 – Technical Report, 43-101CP - Standards of Disclosure for Mineral Projects, and the Technical Report and the Technical Report has been prepared in compliance with such instrument, form, and companion policy.

9)As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the portions of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

10)I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange and other regulatory authority and any publications by them, including electronic publication in the public company files on their websites accessible by the public.

Dated this 2 nd day of March 2018

Signed, Sealed “Rex Clair Bryan, Ph.D.”

Signature of Qualified Person

Rex Clair Bryan, Ph.D.

Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

Anthony Clark, P.E.

Area Lead Mechanical Engineer

POWER Engineers, Inc.

2041 South Cobalt Point Way

Meridian, Idaho 83642

Telephone: (208) 288-6100

Facsimile: (208) 288-6199

Email: tony.clark@powereng.com

This certificate applies to the report entitled: “NI 43-101 Technical Report – Mt Todd Gold Project – 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia” (“Technical Report”), effective date January 24, 2018, issued on March 2, 2018.

I, Anthony Clark, P.E., do hereby certify that:

1)I am a Senior Mechanical Engineer with POWER Engineers, Inc. with a business address at 2041 S. Cobalt Point Way, Meridian, Idaho, 83642, USA.

2)I graduated with a degree in Mechanical Engineering, Bachelor of Science in 2001 from the Missouri University of Science and Technology, Rolla, Missouri. I have worked as a Mechanical Engineer for a total of 16.5 years since my graduation. My relevant experience is in the area of [experience]. I am a PE in Idaho (No. 14669) and P.Eng in Alberta (No. 208138).

3)I have read the definition of “qualified person” set out in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” within the meaning of NI 43-101.

4)I have personally visited and inspected the property which is the subject of the Technical Report on June 28th through June 29th, 2017 to inspect the existing power infrastructure at the site, natural gas pipeline, and power rights-of-way.

5)I am responsible for Sections 18.8, 21.1.3, and 21.2.4 of the Technical Report.

6)I am independent of the issuer, Vista Gold Corp., according to Section 1.5 of NI 43-101.

7)I have had prior involvement with the property that is the subject of the Technical Report. My involvement has consisted of acting as an expert who was relied upon for the Power Station Feasibility Study provided in Appendix F of NI 43-101 Technical Report dated June 2013.

8)I have read NI 43-101, Form 43-101F1 – Technical Report, 43-101CP - Standards of Disclosure for Mineral Projects, and the Technical Report and the Technical Report has been prepared in compliance with such instrument, form, and companion policy.

9)As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the portions of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

10)I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange and other regulatory authority and any publications by them, including electronic publication in the public company files on their websites accessible by the public.

Dated this 2 nd day of March 2018

Signed, Sealed “Anthony Clark, P.E.”

Signature of Qualified Person

Anthony Clark, P.E.

Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

Thomas L. Dyer, P.E.

Mining Engineer

Mine Development Associates

210 South Rock Boulevard

Reno, Nevada 89502

Telephone: (775) 856-5700

Email: tdyer@mda.com

This certificate applies to the report entitled: “NI 43-101 Technical Report – Mt Todd Gold Project – 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia” (“Technical Report”), effective date January 24, 2018, issued on March 2, 2018.

I, Thomas L. Dyer, P.E., do hereby certify that:

1)I am a Senior Engineer with Mine Development Associates with a business address at 210 Rock Blvd, Reno, NV, 89502, USA.

2)I graduated with a B.S. degree in Mine Engineering in 1996 from the South Dakota School of Mines and Technology. I have worked as a Mining Engineer for a total of 22 years since my graduation. My relevant experience includes 11 years of Engineering in an operating open pit mine including underground studies. This operations experience included increasing responsibilities obtaining the position of Chief Engineer. Since that time I have worked as a Consulting Mining Engineer for numerous open pit and underground projects including Preliminary Economic Assessments, Prefeasibility, and Feasibility studies. I am a P.E. in Nevada (No. 15729) and am a Registered Member of SME (# 4029995RM) in good standing.

3)I have read the definition of “qualified person” set out in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” within the meaning of NI 43-101.

4)I have personally visited and inspected the property which is the subject of the Technical Report in April of 2017 for 2 days.

5)I am responsible for Sections 1.6, 1.7, 1.15.3, 15, 15.1, 15.2, 15.3, 15.4, 15.5, 15.6, 16, 21.1.1, 21.2.1, 24.7.2, 25.3, and 26.3 of the Technical Report.

6)I am independent of the issuer, Vista Gold Corp., according to Section 1.5 of NI 43-101.

7)I have had prior involvement with the property that is the subject of the Technical Report. My involvement has consisted of acting as an expert who was relied upon for the NI 43-101 Technical Report – Mt Todd Gold Project, 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia, Amended & Restated; July 7, 2014.

8)I have read NI 43-101, Form 43-101F1 – Technical Report, 43-101CP - Standards of Disclosure for Mineral Projects, and the Technical Report and the Technical Report has been prepared to be compliant with such instrument, form, and companion policy.

9)As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the portions of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

10)I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange and other regulatory authority and any publications by them, including electronic publication in the public company files on their websites accessible by the public.

Dated this 2 nd day of March 2018

Signed, Sealed “Thomas L. Dyer, P.E.”

Signature of Qualified Person

Thomas L. Dyer, P.E.

Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

Amy L. Hudson, Ph.D., CPG, REM

Principal Hydrogeologist/ Geochemist

Tetra Tech, Inc.

1750 Kraft Drive, Suite 1503

Blacksburg, Virginia 24060

Telephone: (703) 885-5447

Email: Amy.Hudson@tetratech.com

This certificate applies to the report entitled: “NI 43-101 Technical Report – Mt Todd Gold Project – 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia” (“Technical Report”), effective date January 24, 2018, issued on March 2, 2018.

I, Amy L. Hudson, Ph.D., CPG, REM, do hereby certify that:

1)I am a Principal Hydrogeologist/Geochemist with Tetra Tech, Inc. with a business address at 1750 Kraft Drive, Suite 1503, Blacksburg, Virginia.

2)I graduated with a degree in Geology and Environmental Science, B.S. in 1998 from the Mary Washington College, Fredericksburg, Virginia and I graduated with a degree in Environmental Science and Engineering, M.S. in 2006 from the Colorado School of Mines, Golden, Colorado. In addition, I graduated with a degree in Geoscience, Ph.D. in 2016 from the University of Massachusetts Amherst, Amherst, Massachusetts. I have worked as a Hydrogeologist/Geochemist for a total of 20 years since my graduation. My relevant experience is in the area of geochemistry, hydrogeology, and environmental science. I am a Certified Professional Geologist in Virginia (No. 002122) and a Registered Environmental Manager in the USA (No. 11854).

3)I have read the definition of “qualified person” set out in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” within the meaning of NI 43-101.

4)I have not visited and inspected the property which is the subject of the Technical Report.

5)I am responsible for Sections 1.15.12, 24.3, and 26.10 of the Technical Report.

6)I am independent of the issuer, Vista Gold Corp., according to Section 1.5 of NI 43-101.

7)I have had prior involvement with the property that is the subject of the Technical Report. My involvement has consisted of acting as an expert who was relied upon for the NI 43-101 Technical Report – Mt Todd Gold Project, 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia, Amended & Restated; July 7, 2014.

8)I have read NI 43-101, Form 43-101F1 – Technical Report, 43-101CP - Standards of Disclosure for Mineral Projects, and the Technical Report and the Technical Report has been prepared in compliance with such instrument, form, and companion policy.

9)As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the portions of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

10)I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange and other regulatory authority and any publications by them, including electronic publication in the public company files on their websites accessible by the public.

Dated this 2 nd day of March 2018

Signed, Sealed “Amy L. Hudson, Ph.D., CPG, REM”

Signature of Qualified Person

Amy L. Hudson, Ph.D., CPG, REM

Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

Chris Johns, M.Sc., P.Eng.

Geological Engineer

Tetra Tech, Inc.

1715 Dickson Avenue, Suite 150

Kelowna, British Columbia V1Y 9G6

Telephone: (250) 862-4832

Email: Chris.Johns@tetratech.com

This certificate applies to the report entitled: “NI 43-101 Technical Report – Mt Todd Gold Project – 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia” (“Technical Report”), effective date January 24, 2018, issued on March 2, 2018.

I, Chris Johns, M.Sc., P.Eng., do hereby certify that:

1)I am a Senior Consultant with Tetra Tech, Inc. with a business address at 150-1715 Dickson Avenue, Kelowna, British Columbia, Canada.

2)I graduated with a degree in Geological Engineering, B.Sc., in 1994 from Queen’s University, Kingston, Ontario. In addition, I graduated with a degree in Environmental Engineering, M.Sc. in 1999 from the University of Alberta, Edmonton, AB. I have worked as a geological engineer for a total of 20 years since my graduation. My relevant experience is in the area of tailings storage facility design from scoping study through feasibility and construction stage. I am a registered Professional Engineer in the Provinces of Alberta and British Columbia, and a Chartered Professional Engineer with the Institution of Engineers Australia.

3)I have read the definition of “qualified person” set out in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” within the meaning of NI 43-101.

4)I have personally visited and inspected the property which is the subject of the Technical Report on June 28th – 29 th for 2 days.

5)I am responsible for Sections 1.15.10, 18.2.4, 21.1.8, 21.2.6, and 26.8 of the Technical Report.

6)I am independent of the issuer, Vista Gold Corp., according to Section 1.5 of NI 43-101.

7)I have not had prior involvement with the property that is the subject of the Technical Report.

8)I have read NI 43-101, Form 43-101F1 – Technical Report, 43-101CP - Standards of Disclosure for Mineral Projects, and the Technical Report has been prepared in compliance with such instrument, form, and companion policy.

9)As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the portions of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

10)I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange and other regulatory authority and any publications by them, including electronic publication in the public company files on their websites accessible by the public.

Dated this 2 nd day of March 2018

Signed, Sealed “Chris Johns, M.Sc., P.Eng.”

Signature of Qualified Person

Chris Johns, M.Sc., P.Eng.

Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

Benjamin S. Johnson, P.E.

Water Treatment Plant Process Engineer

Tetra Tech, Inc.

1560 Broadway, Suite 1400

Denver, Colorado 80202

Telephone: (303) 825-5999

Facsimile: (303) 825-0642

Email: Ben.Johnson@tetratech.com

This certificate applies to the report entitled: “NI 43-101 Technical Report – Mt Todd Gold Project – 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia” (“Technical Report”), effective date January 24, 2018, issued on March 2, 2018.

I, Benjamin S. Johnson, P.E., do hereby certify that:

1)I am a Project Manager with Tetra Tech, Inc. with a business address at 1560 Broadway, Denver, CO 80202, USA.

2)I graduated with a Bachelor of Science degree in Civil Engineering in 2000 from the University of Colorado at Boulder, Boulder, Colorado. In addition, I graduated with a Master of Science in Environmental Engineering in 2006 from the Colorado School of Mines, Golden, Colorado. I have worked as an environmental engineer for a total of 18 years since my graduation. My relevant experience is in the area of drinking water treatment, domestic wastewater treatment, industrial wastewater treatment, and hydraulics. I am a PE in Colorado (No. 38891) and a Project Management Professional with the Project Management Institute.

3)I have read the definition of “qualified person” set out in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” within the meaning of NI 43-101.

4)I have not visited and inspected the property which is the subject of the Technical Report.

5)I am responsible for Sections 21.1.6, 21.2.5, 24.2.2.1, and 24.2.2.2 of the Technical Report.

6)I am independent of the issuer, Vista Gold Corp., according to Section 1.5 of NI 43-101.

7)I have had prior involvement with the property that is the subject of the Technical Report. My involvement has consisted of acting as an expert who was relied upon for NI 43-101 Technical Report – Mt. Todd Gold Project 50,000 tpd Preliminary Feasibility Study Northern Territory, Australia, dated June 2013.

8)I have read NI 43-101, Form 43-101F1 – Technical Report, 43-101CP - Standards of Disclosure for Mineral Projects, and the Technical Report and the Technical Report has been prepared in compliance with such instrument, form, and companion policy.

9)As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the portions of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

10)I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange and other regulatory authority and any publications by them, including electronic publication in the public company files on their websites accessible by the public.

Dated this 2 nd day of March 2018

Signed, Sealed “Benjamin S. Johnson, P.E.”

Signature of Qualified Person

Benjamin S. Johnson, P.E.

Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

Deepak Malhotra, Ph.D.

Principal Metallurgist

Resource Development Inc.

11475 W 48 th Avenue

Wheat Ridge, Colorado 80033

Telephone: (303) 422-1176

Email: deepak@rdiminerals.com

This certificate applies to the report entitled: “NI 43-101 Technical Report – Mt Todd Gold Project – 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia” (“Technical Report”), effective date January 24, 2018, issued on March 2, 2018.

I, Deepak Malhotra, Ph.D., do hereby certify that:

1)I am the President of Resource Development, Inc. (RDi) with a business address at 11475 W. I-70 Frontage Road North, Wheat Ridge, Colorado 80033.

2)I graduated with a degree in Metallurgical Engineering, Masters of Science in 1973 from the Colorado School of Mines in Golden, Colorado. In addition, I graduated with a degree in Mineral Economics, Ph.D. in 1978 from the Colorado School of Mines in Golden, Colorado. I have worked as a metallurgist and mineral economist for a total of 43 years since my graduation. My relevant experience is in the area of metallurgy and mineral economics.

3)I have read the definition of “qualified person” set out in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” within the meaning of NI 43-101.

4)I have not visited and inspected the property which is the subject of the Technical Report.

5)I am responsible for Sections 1.6.1 and 15.7 of the Technical Report.

6)I am independent of the issuer, Vista Gold Corp., according to Section 1.5 of NI 43-101.

7)I have had prior involvement with the property that is the subject of the Technical Report. My involvement has consisted of acting as an expert who was relied upon for the Process Development of the project, I was relied upon for the NI 43-101 Technical Report – Mt Todd Gold Project, 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia, Amended & Restated; July 7, 2014.

8)I have read NI 43-101, Form 43-101F1 – Technical Report, 43-101CP - Standards of Disclosure for Mineral Projects, and the Technical Report and the Technical Report has been prepared in compliance with such instrument, form, and companion policy.

9)As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the portions of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

10)I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange and other regulatory authority and any publications by them, including electronic publication in the public company files on their websites accessible by the public.

Dated this 2 nd day of March 2018

Signed, Sealed “Deepak Malhotra, Ph.D.”

Signature of Qualified Person

Deepak Malhotra, Ph.D.

Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

Zvonimir Ponos, BE, MIEAust, CPeng, NER

General Manager Engineering

Coffey Services Australia Pty Ltd

(trading as Tetra Tech Proteus)

Level 1, 235 St Georges Terrace,

Perth, Western Australia 6000

Telephone: +61-8-6218-2100

Email: zvon.ponos@tetratech.com

This certificate applies to the report entitled: “NI 43-101 Technical Report – Mt Todd Gold Project – 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia” (“Technical Report”), effective date January 24, 2018, issued on March 2, 2018.

I, Zvonimir Ponos, BE, MIEAust, CPeng, NER, do hereby certify that:

1)I am a General Manager of Engineering with Coffey Services Australia Pty Ltd (trading as Tetra Tech Proteus) with a business address at Level 1, 235 St Georges Terrace, Perth, Western Australia, 6000, Australia.

2)I graduated with a degree in Structural Engineering in 1985 from the University of Belgrade in Yugoslavia. I have worked as a Design Engineer, Engineering Manager and Project Manager for more than 30 years since my graduation. My relevant experience is in the areas of structural design, engineering management and project management of chemical, mineral processing and materials handling projects in Gold, Iron Ore, Mineral Sands, Alumina and Base Metals.

3)I am a Chartered Professional Engineer and a Member of the Institution of Engineers Australia (No. 230033). I am also Member of Concrete Institute of Australia (CIA) and Australian Steel Institute (No. 6184). I am a (lapsed) member of both Australian Institute of Project Management (AIPM No 2765) and Project Management Institute, USA (PMI No 167332). All memberships in good standing.

4)I have read the definition of “qualified person” set out in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” within the meaning of NI 43-101.

5)I have personally visited and inspected the property which is the subject of the Technical Report on the 28 th June 2017 for three (3) days.

6)I am responsible for Sections 1.10, 1.15.5, 18.1, 18.2, 18.2.1, 18.2.2, 18.2.3, 18.2.5, 18.2.6, 18.3, 18.4, 18.5, 18.6, 18.7, 21.2.7, 24.6, 24.7.4, and 25.5 of the Technical Report.

7)I am independent of the issuer, Vista Gold Corp., according to Section 1.5 of NI 43-101.

8)I have had prior involvement with the property that is the subject of this Technical Report. My involvement has consisted of acting as a Project Manager for the relevant sections of the Mt. Todd Gold Project 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia issued on the 28 th June, 2013 (document number 11431128500-REP-R0001).

9)I have read NI 43-101, Form 43-101F1 – Technical Report, 43-101CP - Standards of Disclosure for Mineral Projects, and the Technical Report and the Technical Report has been prepared in compliance with such instrument, form, and companion policy.

10)As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the portions of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

11)I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange and other regulatory authority and any publications by them, including electronic publication in the public company files on their websites accessible by the public.

Dated this 2 nd day of March 2018

Signed, Sealed “Zvonimir Ponos, BE, MIEAust, CPeng, NER”

Signature of Qualified Person

Zvonimir Ponos, BE, MIEAust, CPeng, NER

Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

Guy Roemer, P.E.

Environmental Engineer

Tetra Tech, Inc.

1100 South McCaslin Boulevard, Suite 150

Superior, Colorado 80027

Telephone: (303) 664-4624

Email: Guy.Roemer@tetratech.com

This certificate applies to the report entitled: “NI 43-101 Technical Report – Mt Todd Gold Project – 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia” (“Technical Report”), effective date January 24, 2018, issued on March 2, 2018.

I, Guy Roemer, P.E., do hereby certify that:

1)I am an Environmental Engineer with Tetra Tech, Inc. with a business address at 1100 South McCaslin Boulevard, Suite 150, Superior, Colorado, 80027, United States of America.

2)I graduated with a degree in Nuclear Engineering, B.S. in 1995 from Texas A&M University, College Station, Texas. In addition, I graduated with a degree in Nuclear Engineering, M.S. in 1997 from the University of New Mexico, Albuquerque, New Mexico. I have worked as an Environmental Engineer for a total of 20 years since my graduation. My relevant experience is in the area of conducting site-wide and storage facility water balance models of mines using analytical and numerical models. I am a P.E. in Colorado (No. 36810).

3)I have read the definition of “qualified person” set out in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” within the meaning of NI 43-101.

4)I have not visited and inspected the property which is the subject of the Technical Report.

5)I am responsible for Sections 1.15.7, 24.2.1, 24.7.5, 25.7 and 26.5 of the Technical Report.

6)I am independent of the issuer, Vista Gold Corp., according to Section 1.5 of NI 43-101.

7)I have not had prior involvement with the property that is the subject of the Technical Report.

8)I have read NI 43-101, Form 43-101F1 – Technical Report, 43-101CP - Standards of Disclosure for Mineral Projects, and the Technical Report and the Technical Report has been prepared in compliance with such instrument, form, and companion policy.

9)As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the portions of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

10)I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange and other regulatory authority and any publications by them, including electronic publication in the public company files on their websites accessible by the public.

Dated this 2 nd day of March 2018

Signed, Sealed “Guy Roemer, P.E.”

Signature of Qualified Person

Guy Roemer, P.E.

Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

Vicki Scharnhorst, P.E., LEED AP

Principal

Tetra Tech, Inc.

1560 Broadway, Suite 1400

Denver, Colorado 80202

Telephone: (303) 825-5999

Facsimile: (303) 825-0642

Email: Vicki.Scharnhorst@tetratech.com

This certificate applies to the report entitled: “NI 43-101 Technical Report – Mt Todd Gold Project – 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia” (“Technical Report”), effective date January 24, 2018, issued on March 2, 2018.

I, Vicki Scharnhorst, P.E., LEED AP, do hereby certify that:

1)I am a Principal Consultant with Tetra Tech with a business address at 1560 Broadway, Denver, CO 80202, USA.

2)I graduated with a Bachelor of Science degree in Civil Engineering in 1982 from Kansas State University, Manhattan, Kansas. In addition, I graduated with a Master of Public Administration and Policy degree in 2017 from the American University, Washington D.C. I have worked as a civil engineer for a total of 35 years since my graduation. My relevant experience includes civil engineering on large infrastructure projects inclusive of civil works, water quality programs, environmental impact studies, and permitting. I am a licensed Engineer in the states of Nevada (No. 7647), Michigan (No. 43541), Missouri (No. 27930), and Colorado (No. 41466); a water right surveyor in the State of Nevada; a LEED Accredited Professional with the U.S. Green Building Council.

3)I have read the definition of “qualified person” set out in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” within the meaning of NI 43-101.

4)I have personally visited and inspected the property which is the subject of the Technical Report on June 28 th and 29 th for two days.

5)I am responsible for Sections 1.1, 1.2, 1.3, 1.12, 1.15.1, 1.15.6, 2, 3, 4, 5, 18.0 (Introduction), 20, 21.1.5, 21.1.7, 23, 24.0 (Introduction), 24.2, 24.2.2, 24.2.2.3, 24.2.2.4, 24.2.2.5, 24.4, 24.7, 25.1, 25.6, 26, 26.1, and 26.4 of the Technical Report.

6)I am independent of the issuer, Vista Gold Corp., according to Section 1.5 of NI 43-101.

7)I have not had prior involvement with the property that is the subject of the Technical Report.

8)I have read NI 43-101, Form 43-101F1 – Technical Report, 43-101CP - Standards of Disclosure for Mineral Projects, and the Technical Report and the Technical Report has been prepared in compliance with such instrument, form, and companion policy.

9)As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the portions of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

10)I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange and other regulatory authority and any publications by them, including electronic publication in the public company files on their websites accessible by the public.

Dated this 2 nd day of March 2018

Signed, Sealed “Vicki Scharnhorst, P.E., LEED AP”

Signature of Qualified Person

Vicki Scharnhorst, P.E., LEED AP

Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

D. Erik Spiller, QP

Principal Metallurgist

Tetra Tech, Inc.

350 Indiana Street, Suite 500

Golden, Colorado 80401

Telephone: (303) 217-5700

Facsimile: (303) 217-5705

Email: Erik.Spiller@tetratech.com

This certificate applies to the report entitled: “NI 43-101 Technical Report – Mt Todd Gold Project – 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia” (“Technical Report”), effective date January 24, 2018, issued on March 2, 2018.

I, D. Erik Spiller, QP, do hereby certify that:

1)I am the Principal Metallurgist with Tetra Tech, Inc. with a business address at 350 Indiana Street, Suite 500, Golden, Colorado 80401.

2)I graduated with a Bachelor of Science degree in Metallurgical Engineering in 1970 from the Colorado School of Mines in Golden, Colorado. I have worked as a metallurgical engineer for more than 40 years since my graduation. My relevant experience is in the area of metallurgy. I am a Qualified Professional (QP) member of the Mining and Metallurgical Society of America (MMSA #01021QP). In addition, I am a Registered (QP) member of Society for Mining, Metallurgy, and Exploration, Inc. (SME #3051820RM).

3)I have read the definition of “qualified person” set out in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” within the meaning of NI 43-101.

4)I have not visited and inspected the property which is the subject of the Technical Report.

5)I am responsible for Sections 1.8, 1.9, 1.15.4, 1.15.11, 6.5, 6.6, 13, 17, 21.1.2, 21.2.3, 24.7.3, 25.4, 26.9, and 26.11 of the Technical Report.

6)I am independent of the issuer, Vista Gold Corp., according to Section 1.5 of NI 43-101.

7)I have not had prior involvement with the property that is the subject of the Technical Report.

8)I have read NI 43-101, Form 43-101F1 – Technical Report, 43-101CP – Standards of Disclosure for Mineral Projects, and the Technical Report and the Technical Report has been prepared in compliance with such instrument, form, and companion policy.

9)As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the portions of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

10)I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange and other regulatory authority and any publications by them, including electronic publication in the public company files on their websites accessible by the public.

Dated this 2 nd day of March 2018

Signed, Sealed “D. Erik Spiller, QP”

Signature of Qualified Person

D. Erik Spiller, QP

Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

Jessica I. Spriet, P.E.

Professional Engineer

Tetra Tech, Inc.

350 Indiana Street, Suite 500

Golden, Colorado 80401

Telephone: (303) 217-5700

Facsimile: (303) 217-5705

Email: Jessica.Spriet@tetratech.com

This certificate applies to the report entitled: “NI 43-101 Technical Report – Mt Todd Gold Project – 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia” (“Technical Report”), effective date January 24, 2018, issued on March 2, 2018.

I, Jessica I. Spriet, P.E., do hereby certify that:

1)I am a Professional Engineer with Tetra Tech, Inc. with a business address at 350 Indiana Street, Suite 500, Golden, Colorado 80401, United States.

2)I graduated with a degree in Geological Engineering, Bachelors of Science in 2006 from the Colorado School of Mines, Golden, Colorado. In addition, I graduated with a degree in Geological Engineering, Masters of Engineering in 2007 from the Colorado School of Mines, Golden, Colorado. I have worked as a Geological Engineer for six years, and a Mineral Economist for five year, for a total of eleven years since my graduation. My relevant experience is in the area of geotechnical engineering and mineral economics. I am a Professional Engineer in the State of Colorado (P.E. 0045454).

3)I have read the definition of “qualified person” set out in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” within the meaning of NI 43-101.

4)I have not visited and inspected the property which is the subject of the Technical Report.

5)I am responsible for Sections 1.11, 1.13, 1.14, 1.15.9, 19, 21 (Introduction), 21.1, 21.2, 22, 24.1, 24.7.6, 24.7.7, 24.7.8, and 26.7 of the Technical Report.

6)I am independent of the issuer, Vista Gold Corp., according to Section 1.5 of NI 43-101.

7)I have not had prior involvement with the property that is the subject of the Technical Report.

8)I have read NI 43-101, Form 43-101F1 – Technical Report, 43-101CP - Standards of Disclosure for Mineral Projects, and the Technical Report and the Technical Report has been prepared in compliance with such instrument, form, and companion policy.

9)As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the portions of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

10)I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange and other regulatory authority and any publications by them, including electronic publication in the public company files on their websites accessible by the public.

Dated this 2 nd day of March 2018

Signed, Sealed “Jessica I. Spriet, P.E.”

Signature of Qualified Person

Jessica I. Spriet, P.E.

Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

Keith Thompson, CPG, PG

Professional Geologist

Tetra Tech, Inc.

1100 McCaslin Boulevard, Suite 150

Superior, Colorado 80027

Telephone: (303) 664-4630

Email: Keith.Thompson@tetratech.com

This certificate applies to the report entitled: “NI 43-101 Technical Report – Mt Todd Gold Project – 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia” (“Technical Report”), effective date January 24, 2018, issued on March 2, 2018.

I, Keith Thompson, CPG, PG, do hereby certify that:

1)I am a Senior Hydrogeologist with Tetra Tech with a business address at 1100 McCaslin Boulevard, Suite 150, Superior, Colorado 80027, USA.

2)I graduated with Bachelor of Science degree in Geology in 1975 from Youngstown State University, Youngstown, Ohio, USA. In addition, I graduated with a Master of Science degree in Geology in 1979 from the University of Wyoming, Laramie, Wyoming, USA. I have worked as a hydrogeologist for a total of 38 years since my graduation. My relevant experience is in the areas of mining hydrology and hydrogeology, environmental hydrology and hydrogeology, and groundwater flow and transport modeling. I am a Certified Professional Geologist (No. 6005) and member of the American Institute of Professional Geologists and a licensed Professional Geologist in the (USA) states of Alaska (No. 700), California (No. 5572), Idaho (No. 726), Utah (No. 5258797-2250) and Wyoming (No. 2454).

3)I have read the definition of “qualified person” set out in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” within the meaning of NI 43-101.

4)I have not visited and inspected the property which is the subject of the Technical Report.

5)I am responsible for Sections 1.15.8, 21.1.4, 21.2.2, 24.5, 26.6 of the Technical Report.

6)I am independent of the issuer, Vista Gold Corp., according to Section 1.5 of NI 43-101.

7)I have had prior involvement with the property that is the subject of the Technical Report. My involvement has consisted of acting as an expert who was relied upon for “NI 43-101 Technical Report – Mt Todd Gold Project 50,000 tpd Preliminary Feasibility Study, Northern Territory, Australia, Appendix K – Regional Hydrogeology” dated June 2013.

8)I have read NI 43-101, Form 43-101F1 – Technical Report, 43-101CP - Standards of Disclosure for Mineral Projects, and the Technical Report and the Technical Report has been prepared in compliance with such instrument, form, and companion policy.

9)As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the portions of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

10)I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange and other regulatory authority and any publications by them, including electronic publication in the public company files on their websites accessible by the public.

Dated this 2 nd day of March 2018

Signed, Sealed “Keith Thompson, CPG, PG”

Signature of Qualified Person

Keith Thompson, CPG, PG

Print Name of Qualified Person

Tetra Tech | March 2018 | i