EX-99.1 2 ego_ex991.htm TECHNICAL REPORT
Exhibit 99.1
Technical Report
Olympias Mine
Greece
Centered on Latitude E 40° 36' and Longitude N 23° 50' approximately
Effective Date: December 31, 2019
Prepared by:
Eldorado Gold Corporation
1188 Bentall 5 - 550 Burrard Street
Vancouver, BC V6C 2B5

Qualified Persons | Company
Mr. David Sutherland, P.Eng. | Eldorado Gold Corporation
Mr. Paul Skayman, FAusIMM. | Eldorado Gold Corporation
Mr. Colm Keogh, P.Eng. | Eldorado Gold Corporation
Mr. Sean McKinley, P.Geo. | Eldorado Gold Corporation
Mr. Ertan Uludag, P.Geo. | Eldorado Gold Corporation

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Table of Contents

SECTION ● 1 SUMMARY | 1-1
1.1 Property Description and Location | 1-1
1.2 Geology | 1-2
1.3 Drilling, Sampling and Verification | 1-2
1.4 Metallurgical Testwork | 1-3
1.5 Mineral Resources | 1-4
1.6 Mineral Reserves and Mining | 1-5
1.7 Processing | 1-8
1.8 Infrastructure | 1-9
1.9 Market studies and contracts | 1-10
1.10 Environmental | 1-10
1.11 Capital and Operating costs | 1-11
1.12 Economics | 1-12
1.13 Conclusions and recommendations | 1-14
SECTION ● 2 INTRODUCTION | 2-1
SECTION ● 3 RELIANCE ON OTHER EXPERTS | 3-1
SECTION ● 4 PROPERTY DESCRIPTION AND LOCATION | 4-1
4.1 Property Location | 4-1
4.2 Land Tenure | 4-2
4.3 Permitting | 4-2
4.4 Royalties and Taxes | 4-3
4.5 Environmental Liabilities | 4-3
4.6 Other | 4-3
SECTION ● 5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY | 5-1
5.1 Location and Accessibility | 5-1
5.2 Infrastructure and Local Resources | 5-2
5.3 Climate and Physiography | 5-3
5.4 Surface Rights | 5-3
SECTION ● 6 HISTORY | 6-1
6.1 Introduction | 6-1
6.2 Ownership and Work Carried Out | 6-1
SECTION ● 7 GEOLOGICAL SETTING AND MINERALIZATION | 7-1
7.1 Regional Geology | 7-1
7.2 Local Geology | 7-3
SECTION ● 8 DEPOSIT TYPES | 8-1
8.1 Deposit Geology | 8-1
8.2 Deposit Model | 8-4
SECTION ● 9 EXPLORATION | 9-1

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SECTION ● 10 DRILLING | 10-1
10.1 Introduction | 10-1
10.2 Core Handling and Sampling Procedures | 10-2
10.3 Surveys | 10-2
10.4 Core Recovery | 10-2
10.5 Bulk Density Determinations | 10-2
SECTION ● 11 SAMPLE PREPARATION, ANALYSES AND SECURITY | 11-1
11.1 Introduction | 11-1
11.2 Sample Preparation and Assaying | 11-1
11.3 QA/QC Program | 11-1
11.4 Concluding Statement | 11-10
SECTION ● 12 DATA VERIFICATION | 12-1
12.1 Introduction | 12-1
12.2 Database checks | 12-1
12.3 Model to Mill Reconciliation | 12-1
12.4 Concluding Statement | 12-2
SECTION ● 13 MINERAL PROCESSING AND METALLURGICAL TESTING | 13-1
13.1 Historical Metallurgical Testing | 13-1
13.2 SGS Metallurgical Test Report No. 14607-001 – Final, 2015 | 13-1
13.3 Conclusions and Recommendations | 13-22
SECTION ● 14 MINERAL RESOURCE ESTIMATES | 14-1
14.1 Introduction | 14-1
14.2 Mineralized Domains | 14-1
14.3 Data Analysis | 14-3
14.4 Block Model | 14-5
14.5 Grade Estimation | 14-5
14.6 Block Model Validation | 14-8
14.7 Mineral Resource Classification | 14-14
14.8 Mineral Resource Estimate | 14-18
SECTION ● 15 MINERAL RESERVE ESTIMATES | 15-1
15.1 Cut-Off Grade | 15-1
15.2 Mining Shapes | 15-3
15.3 Dilution and Recovery Factors | 15-3
15.4 Reserve Resource Reconciliation | 15-4
SECTION ● 16 MINING METHODS | 16-1
16.1 Introduction | 16-1
16.2 Geotechnical Considerations | 16-2
16.3 Mining Methods | 16-5
16.4 Backfill | 16-9
16.5 Mine Design | 16-10
16.6 Ventilation | 16-11

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16.7 Drill and Blast | 16-13
16.8 Underground infrastructure | 16-18
16.9 Mining Schedule | 16-20
SECTION ● 17 RECOVERY METHODS | 17-1
17.1 Introduction | 17-1
17.2 Process Plant Design Criteria | 17-1
17.3 Process Flowsheet | 17-4
17.4 Process Plant Description | 17-6
17.5 Reagents Mixing, Storage, and Distribution | 17-15
17.6 Services, Water, and Air Distribution | 17-19
SECTION ● 18 PROJECT INFRASTRUCTURE | 18-1
18.1 Site Access | 18-1
18.2 Local Services | 18-1
18.3 Port Facilities | 18-2
18.4 Communication System | 18-3
18.5 Water Supply | 18-4
18.6 Mine dewatering System | 18-4
18.7 Mine Waste Management | 18-4
18.8 New Facilities | 18-5
18.9 Surface Workshop | 18-5
18.10 Shaft | 18-6
18.11 Electrical Power | 18-6
SECTION ● 19 MARKET STUDIES AND CONTRACTS | 19-1
19.1 Concentrate Marketing | 19-1
19.2 Concentrate Sales | 19-1
19.3 Gold Concentrate | 19-1
19.4 Zinc Concentrate | 19-1
19.5 Lead Concentrate | 19-1
19.6 Other Contracts | 19-2
SECTION ● 20 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT | 20-1
20.1 Environmental Impact Study | 20-1
20.2 Legislative Framework for the preparation of the EIS | 20-1
20.3 Landscape | 20-1
20.4 Geological Environment | 20-1
20.5 Ground – Land Use | 20-2
20.6 Water Resources | 20-2
20.7 Atmospheric Environment | 20-2
20.8 Sound Environment | 20-2
20.9 Ecosystems – Flora – Fauna | 20-2
20.10 Social and Financial Environment | 20-3

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20.11 Closure Plan | 20-3
20.12 Environmental Costs and Guarantees | 20-3
SECTION ● 21 CAPITAL AND OPERATING COSTS | 21-1
21.1 Capital Costs | 21-1
21.2 Operating Costs | 21-2
SECTION ● 22 ECONOMIC ANALYSIS | 22-1
22.1 Methods, Assumptions and Basis | 22-1
22.2 Cash Flows | 22-1
22.3 Royalties and Other Fees | 22-5
22.4 Closure and Salvage Value | 22-6
22.5 Taxation | 22-6
22.6 Financing Costs | 22-7
22.7 Third Party Interests | 22-7
22.8 Sensitivity Analysis | 22-7
SECTION ● 23 ADJACENT PROPERTIES | 23-1
SECTION ● 24 OTHER RELEVANT DATA AND INFORMATION | 24-1
SECTION ● 25 INTERPRETATION AND CONCLUSIONS | 25-1
25.1 Introduction | 25-1
25.2 Geology and Mineral Resources | 25-1
25.3 Mining and Mineral Reserves | 25-1
25.4 Metallurgy and Processing | 25-2
25.5 Infrastructure | 25-3
25.6 Financial Performance | 25-3
25.7 Closure | 25-4
25.8 Concluding Statement | 25-4
SECTION ● 26 RECOMMENDATIONS | 26-1
26.1 Mining and Mineral Reserves | 26-1
26.2 Metallurgy and processing | 26-1
26.3 Infrastructure and other | 26-2
SECTION ● 27 REFERENCES | 27-1
SECTION ● 28 Date and Signature Page | 28-1

LIST OF FIGURES

Figure 1 1: Isometric View of Olympias Underground Zones | 1-6
Figure 4 1: Property Concessions | 4-1
Figure 5 1: Location of the Olympias Property | 5-1
Figure 5 2: Olympias Site Layout | 5-2
Figure 7 1: Regional Geology of the Kassandra Mining District (modified from Siron et al, 2018) | 7-2
Figure 8 1: Geology of the Olympias mine area (modified from Siron et al., 2018) | 8-2

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Figure 8 2: Simplified Geological Cross Section through the Olympias Deposit (facing NW) | 8-3
Figure 8 3: Three-dimensional oblique View of the Major Ore Bodies in the Olympias Mine looking west-northwest | 8-3
Figure 11 1: Gold CRM Charts, 2018-2019 Olympias Mine | 11-4
Figure 11 2: Ag, Pb, Zn GBM310-14 Charts, 2018-2019 Olympias Mine | 11-5
Figure 11 3: Ag, Pb, Zn GBM913-11 Charts, 2018-2019 Olympias Mine | 11-6
Figure 11 4: Ag, Pb, Zn OREAS134A Chart, 2018-2019 Olympias Mine | 11-7
Figure 11 5: Au, Ag, Pb, Zn Blank Performance by Metal | 11-8
Figure 11 6: RPD plot for coarse duplicates Au | 11-9
Figure 11 7: RPD plot for coarse duplicates Zn | 11-9
Figure 11 8: RPD plot for pulp duplicates Au | 11-10
Figure 11 9: RPD plot for pulp duplicates Zn | 11-10
Figure 13 1: Galena Liberation – Fresh Ore | 13-4
Figure 13 2: Galena Liberation – Old Ore | 13-5
Figure 13 3: Sphalerite Liberation – Fresh Ore | 13-6
Figure 13 4: Sphalerite Liberation – Old Ore | 13-7
Figure 13 5: Pyrite Liberation – Fresh Ore | 13-8
Figure 13 6: Pyrite Liberation – Old Ore | 13-9
Figure 13 7: Locked-Cycle Test Flotation Flowsheet | 13-15
Figure 14 1: Ore Zones, Olympias Mine | 14-2
Figure 14 2: Location of Areas of previous mining | 14-7
Figure 14 3: 3D Oblique View of EAST Zone comparing Block Au Grades to Composite Au Grades | 14-8
Figure 14 4: 3D Oblique View of FLATS Zone comparing Block Au Grades to Composite Au Grades | 14-8
Figure 14 5: 3D Oblique View of WEST Zone comparing Block Au Grades to Composite Au Grades | 14-9
Figure 14 6: Au SWATH Plots | 14-11
Figure 14 7: Pb SWATH Plots | 14-12
Figure 14 8: Zn SWATH Plots | 14-13
Figure 14 9: East Zone Mineral Resource Classification | 14-15
Figure 14 10: West Zone Mineral Resource Classification | 14-16
Figure 14 11: Flats Mineral Resource classification | 14-17
Figure 15 1: 2019 Reserves Dilution Estimation | 15-4
Figure 16 1: Isometric View of Olympias Mining Areas | 16-1
Figure 16 2: Empirical Stope Stability Analysis using 50th percentile Geotechnical Data | 16-4
Figure 16 3: Isometric View of Olympias Underground – Mining Methods | 16-6
Figure 16 4: View of typical primary secondary Longhole Stopes in the Flats | 16-7
Figure 16 5: Longitudinal DAF | 16-8
Figure 16 6: Multi-Pass DAF Stope | 16-9

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Figure 16 7: LOM Ventilation Strategy Planned | 16-12
Figure 16 8: TLHOS Long Section | 16-15
Figure 16 9: TLHOS Cross Section | 16-16
Figure 16 10: Typical Lateral Development Drill Pattern | 16-17
Figure 16 11: LOM Production Profile | 16-24
Figure 17 1: Simplified Process Flowsheet of Olympias Plant | 17-5
Figure 18 1: Location of Olympias Mine, Port and TMF | 18-1
Figure 18 2: Olympias Site Layout | 18-2
Figure 18 3: Olympias Concentrator | 18-3
Figure 18 4: Kokkinolakkas TMF | 18-5

LIST OF TABLES

Table 1 1: Olympias Mine Mineral Resources as of September 30, 2019 | 1-5
Table 1 2: Olympias Mineral Reserves as of September 30, 2019 | 1-6
Table 1 3: Capital Cost Summary | 1-11
Table 1 4: Operating Costs Summary | 1-12
Table 1 5: Key Economic Results | 1-13
Table 1 6: Metal Price Sensitivity Analysis | 1-13
Table 1 7: Capital and Operating Cost Sensitivity Analysis | 1-14
Table 4 1: List of Concessions | 4-2
Table 6 1: Summary of the history of the Property | 6-1
Table 10 1: Drilling by Year | 10-1
Table 11 1: Summary of Assay Methods | 11-2
Table 11 2: Blank Failure Criteria for Au, Ag, Pb and Zn | 11-2
Table 11 3: Summary of QA/QC sampling in 2018 – 2019 | 11-2
Table 11 4: Summary of CRM expected values and standard deviations | 11-3
Table 12 1: Reconciliation for April 2017 to June 2018 | 12-2
Table 12 2: Reconciliation for January to December 2019 | 12-2
Table 13 1: Old Ore and Fresh Ore Head Assays | 13-2
Table 13 2: Mineral Modals – Fresh Ore | 13-3
Table 13 3: Mineral Modals – Old Ore | 13-3
Table 13 4: Electron Microprobe Mineral Assay Summary | 13-10
Table 13 5: UCS Test Results | 13-11
Table 13 6: Bond Rod Mill Grindability Test Results | 13-11
Table 13 7: Bond Ball Mill Grindability Test Results | 13-11
Table 13 8: Lead Concentrate Size-by-Size Analysis | 13-13
Table 13 9: LCT-1 Grind Size Summary | 13-14
Table 13 10: LCT-1 Flotation Conditions | 13-15
Table 13 11: Locked-Cycle Test Metallurgical Projection – Fresh Ore | 13-16
Table 13 12: Static Settling Results Summary | 13-17

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Table 13 13: Optimized Dynamic Settling Results Summary | 13-18
Table 13 14: Underflow Rheology Results Summary | 13-18
Table 13 15: Vane Rheological Data Summary – Gold/Pyrite/Arsenopyrite Concentrate | 13-19
Table 13 16: Vacuum Filtration Results Summary of the Final Tail | 13-19
Table 13 17: Pressure Filtration Results Summary for the Concentrates | 13-21
Table 13 18: Transportable Moisture Level Test Summary | 13-22
Table 14 1: Uncapped Gold Composite Statistics for each Zone | 14-3
Table 14 2: Uncapped Silver Composite Statistics for each Zone | 14-3
Table 14 3: Uncapped Lead Composite Statistics for each Zone | 14-3
Table 14 4: Uncapped Zinc Composite Statistics for each Zone | 14-3
Table 14 5: Top-Capping Grades | 14-4
Table 14 6: Capped Gold Composite Statistics for each Zone | 14-4
Table 14 7: Capped Silver Composite Statistics for each Zone | 14-4
Table 14 8: Capped Lead Composite Statistics for each Zone | 14-4
Table 14 9: Capped Zinc Composite Statistics for each Zone | 14-4
Table 14 10: Block Model Parameters | 14-5
Table 14 11: Global Model Mean Grade Gold Values (g/t) by Domain | 14-10
Table 14 12: Olympias Mineral Resources, as of September 30, 2019 | 14-18
Table 15 1: Olympias Mineral Reserves as of September 30, 2019 | 15-1
Table 15 2: Target Operating Costs | 15-2
Table 15 3: Estimated Impact of Raised Breakeven Cut-off Grade on Reserves and LOM | 15-2
Table 15 4: 2019 Reserves Dilution Estimation | 15-4
Table 15 5: Conversion of Mineral Resources to Mineral Reserves | 15-5
Table 16 1: Results from logging of geotechnical drill holes in Flats Zone | 16-3
Table 16 2: Inputs to the Stability Graph Method | 16-3
Table 16 3: Olympias Mine Equipment List | 16-11
Table 16 4: Powder Factors for Longhole Stoping and Lateral Development/DAF | 16-17
Table 16 5: Olympias Mine, Estimated Stope Productivities | 16-20
Table 16 6: Projected Development Schedule over the LOM | 16-22
Table 16 7: Olympias LOM Ore Production Schedule | 16-23
Table 17 1: Major Process Design Criteria for Phase II | 17-1
Table 21 1: Exchange Rates | 21-1
Table 21 2: Summary of Capital Costs | 21-1
Table 21 3: Summary of Operating Costs | 21-3
Table 22 1: Key Economic Results | 22-2
Table 22 2: Olympias Mine Cashflow Analysis | 22-3
Table 22 3: Gold Royalty | 22-5
Table 22 4: Silver Royalty | 22-5
Table 22 5: Lead Royalty | 22-6
Table 22 6: Zinc Royalty | 22-6
Table 22 7: Metal Price Sensitivity Analysis | 22-8
Table 22 8: Capital and Operating Cost Sensitivity Analysis | 22-8

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GLOSSARY
Units of Measure

Annum (year) | a
Billion | B
Centimeter | cm
Cubic centimeter | cm3
Cubic meter | m3
Day | d
Days per year (annum) | d/a
Degree | °
Degrees Celsius | °C
Dollar (American) | US$
Dollar (Canadian) | CAN$
Euro | €
Gram | g
Grams per litre | g/L
Grams per tonne | g/t
Greater than | >
Hectare (10,000 m2) | ha
Hour | h
Kilo (thousand) | k
Kilogram | kg
Kilograms per cubic meter | kg/m3
Kilograms per hour | kg/h
Kilograms per square meter | kg/m2
Kilometer | km
Kilometers per hour | km/h
Kilopascal | kPa
Kilotonne | kt
Kilovolt | kV
Kilowatt hour | kWh
Kilowatt hours per tonne | kWh/t
Kilowatt hours per year | kWh/a
Kilowatt | kW
Less than | <
Litre | L
Megavolt Ampere | MVA
Megawatt | MW
Meter | m
Meter above Sea Level | masl

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Metric ton (tonne) |
Microns | µm
Milligram | mg
Milligrams per litre | mg/L
Millilitre | mL
Millimeter | mm
Million cubic meters | Mm3
Million ounces | Moz
Million tonnes per Annum | Mtpa
Million tonnes | Mt
Million | M
Million Years | Ma
Newton | N
Ounce | oz
Parts per billion | ppb
Parts per million | ppm
Percent | %
Percent by Weight | wt%
Pound | lb
Square centimeter | cm2
Square kilometer | km2
Square meter | m2
Thousand tonnes | kt
Three Dimensional | 3D
Tonnes per day | t/d or tpd
Tonnes per hour | tph
Tonnes per year | tpa
Volt | V
Watt | W
Weight/volume | w/v
Weight/weight | w/w

Abbreviations and Acronyms

Acidity or Alkalinity | pH
Acid Base Accounting | ABA
Aluminum | Al
ALS | ALS
Antimony | Sb
Argillic | ARG
Argon | Ar
Arsenic | As
Association for the Advancement of Cost Engineering | AACE

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Atomic Adsorption | AA
Barium | Ba
Block Model | BM
Bond Abrasion Index | Ai
Bottle Roll | BR
Bed Volumes | BV
Business Opening and Operations Permit | GSM
Cadmium | Cd
Calcium Hydroxide | Ca(OH)2
Canadian Institute of Mining, Metallurgy, and Petroleum | CIM
Cemented Aggregate Fill | CAF
Certified Reference Material | CRM
Cobalt | Co
Coefficient of Variance | CV
Concentric Cylinder Rotational Viscometry | CCRV
Construction Management | CM
Copper | Cu
Copper Sulphate | CuSO4
Critical Solids Density | CSD
Cyanide | CN
Drift and Fill | DAF
Semi pure gold alloy | Doré
East | E
Eldorado Gold Corporation | Eldorado
Engineering, Procurement, Construction Management | EPCM
Environmental Impact Assessment | EIA
Environmental Impact Statement | EIS
Environmental Management Plan | EMP
European Goldfiels | EGL
European Union | EU
Factor of Safety | FOS
Feasibility Study | FS
Flocculant | FLOC
General and Administration | G&A
Geological Strength Index | GSI
Geosynthetic Clay Liner | GCL
Gold | Au
Gold Equivalent | Au Equiv
Hellas Gold SA | Hellas Gold
Hellenic Fertilizer Company | HFC
High Density Polyethylene | HDPE
High Grade | HG
Hydrochloric Acid | HCl
Hydrogen Oxide | H2O
In Stream Analysis System | ISA
Induced Polarization | IP
Inductively Coupled Plasma | ICP
Inductively Coupled Plasma Emission Spectroscopy | ICP-ES
Inductively Coupled Plasma Mass Spectrometry | ICP-MS
Inner Diameter | ID

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Internal Rate of Return | IRR
International Financial Reporting Standards | IFRS
International Organization for Standardization | ISO
Iron | Fe
Lead | Pb
Life-of-mine | LOM
Locked Cycle Test | LCT
Manganese | Mn
Mechanical, Piping, Electrical, Instrumentation | MPEI
Measured & Indicated | M&I
Mercury | Hg
Methyl Isibutyl Carbinol | MIBC
Ministry of Environment | MOE
National Instrument 43-101 | NI 43-101
Nearest Neighbour | NN
Nearest Neighbour Kriging | NNK
Net Acid Generation | NAG
Net Present Value | NPV
Net Smelter Return | NSR
Nickel | Ni
North | N
North East | NE
North West | NW
Operator Control Station | OCS
Ordinary Kriging | OK
Outer Diameter | OD
Paste Fill | PF
Polyvinyl Chloride | PVC
Potassic | POT
Potassium | K
Potential of Hydrogen | pH
Prefeasibility Study | PFS
Probability Assisted Constrained Kriging | PACK
Public Power Corporation | PPC
Quarter | Q
Qualified Person(s) | QP(s)
Quality assurance | QA
Quality control | QC
Quantitative Evaluation of Minerlas by Scanning Electron Microscopy | QEMSCAN
Quartz | Qz
Request for Quotations | RFQ
RL=Reduced Level | masl
Rock Quality Designation | RQD
Run of Mine | ROM
Selenium | Se
Silicon | Si
Silver | Ag
Sodium Cyanide | NaCN
Sodium Hydroxide | NaOH
Sodium Metabisulphite | Na2S2O5

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Sodium Metabisulphite | SMBS
South | S
South East | SE
South West | SW
Specific Gravity | SG
Spherical | SPH
Standard Reference Material | SRM
Strontium | Sn
Suub-level Caving | SLC
Sulfur | S
Sulfur Dioxide | SO2
Sulphide | S2-
Sulphuric Acid | H2SO4
Tailings Management Facility | TMF
Technical Study | TS
Transportable Moisture Limit | TML
Transverse Longhole Open Stoping | TLHOS
TVX Gold Inc. | TVX
Unconfirmed Compressive Strength | UCS
Underflow | U/F
Underground | UG
Universal Transverse Mercador | UTM
Uranium | U
Value Added Tax | VAT
West | W
Work Breakdown Structure | WBS
Xanthate Collector | SIPX
Zinc | Zn

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SECTION ● 1
Summary
Eldorado Gold Corporation (Eldorado), an international gold mining company based in Vancouver, British Columbia, owns and operates the Olympias mine in Greece through its 95% owned subsidiary Hellas Gold SA (Hellas Gold). Eldorado has prepared this technical report of the Olympias mine to provide an updated description of the geology and mineralization, mineral resources and mineral reserves, and mine and mill operations in light of significant additional technical studies as well as operating experience since the prior technical report, which had an effective date of July 14, 2011 (Technical Report on the Olympias Project Au Pb Zn Ag Deposit, Northern Greece, 2011).
Information and data for this report were obtained from the Olympias mine and Eldorado’s corporate office. The work entailed review of pertinent geological, mining, process and metallurgical data in sufficient detail to support the preparation of this technical report.
The qualified persons responsible for preparing this technical report as defined in National Instrument 43-101 (NI 43-101), Standards of Disclosure for Mineral Projects and in compliance with 43-101F1 (the “Technical Report”) are David Sutherland, P.Eng., Paul Skayman, FAusIMM, Ertan Uludag, P.Geo., Colm Keogh, P.Eng. and Sean McKinley, P.Geo., whom are all employees of Eldorado. All qualified persons have visited the Olympias gold mine.
1.1
Property Description and Location
The Property is located within the Kassandra Mines complex located on the Halkidiki Peninsula of Northern Greece. The Olympias mine lies 9 km north-northwest of the Stratoni port and loading facility, accessed by a paved road along the coast. The terrain is characterized by hills rising to about 600 m above sea level, with steeply incised valleys.
The Property consists of mining concession numbers F13 and F14, which have a combined area of 47.27 km2. Hellas Gold has been granted mining rights over these concessions until 7 April 2024. The concessions are conditionally renewable for a further two consecutive periods of 25 years each. Hellas Gold has ownership of a small portion of private land within the concessions.
In July 2011, the Ministry of Environment (MOE) formally approved the Environmental Impact Statement (EIS) submitted by Hellas Gold for the three Kassandra mines mine sites, being Olympias, Skouries, and Stratoni. This EIS is valid for 10 years and subject to renewal in 2021.
For production to commence, the MOE required the submission of a technical study. A study was submitted to the MOE and approved in early 2012. The installation permit for what was termed the Phase II process plant was issued on 22 March 2016. The Company received the operating permit for the Phase II plant in September 2017, allowing commencement of commercial production operations. Also, in September 2017 the Company received an extension of the installation permit and an interim operating permit for the Kokkinolakkas TMF, as well as the delayed installation permit for the paste backfill plant.
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Notifications for the Operation of the Olympias Paste Plant and Kokkinolakkas Tailings Management Facility (TMF) were formally submitted in 2018 and remain in force in line with new legislation that replaced previous operating permit issuance procedures.
1.2
Geology
The Western Tethyan orogenic belt in southeast Europe contains several major metallogenic provinces including the Serbo-Macedonian Metallogenic Province that hosts the Kassandra mining district. Crystalline basement within the district includes the upper Serbo Macedonian Vertiskos Unit and the lower Kerdilion Unit exposed within the southern Rhodope metamorphic core complex.
The Olympias deposit is located 6 km north of the Stratoni fault within the Kerdilion unit. Replacement-style sulfide orebodies are hosted by marble interlayered within a sequence of quartzo-feldspathic biotite gneiss, amphibolite and plagioclase microcline orthogneiss. The massive sulfide orebodies plunge shallowly to the southeast for over 1.8 km, subparallel to the orientation of F2 fold hinges and a locally developed L2 intersection lineation. The locations of the sulfide lenses, however, are largely controlled by strands of the ductile-brittle Kassandra fault and East fault and sub-horizontal shear zones that occur between the two faults.
Sulfide mineralogy of the Olympias deposit consists of coarse-grained, massive and banded lenses dominated by variable amounts of sphalerite, galena, pyrite, arsenopyrite, chalcopyrite and boulangerite. Gold occurs primarily in solid solution with arsenopyrite and pyrite.
Olympias is an example of a polymetallic carbonate replacement deposit. However it is somewhat unusual due to the high Au content of the deposit. Key characteristics of this class of deposit include carbonate host rocks, massive sulfide mineralization, spatial and temporal relationship with magmatism and zoned metal distribution.
1.3
Drilling, Sampling and Verification
Diamond drillholes continue to be the prime source of subsurface geologic and grade data for the Olympias deposit. The previous operator, TVX, drilled 764 drillholes for a total of 93,246 m. These are becoming less important as new information is acquired. Currently holes are drilled by Eldorado using contractors drilling HQ or NQ-size (63.5 mm or 47.6 mm nominal core diameter). The average drillhole depth is 100 m, as the holes are drilled from locations underground giving good intersection angles with the zones. There are currently 1,855 drillholes for 210,893 m in the database.
Core is delivered to secure core logging areas, and the core is logged in detail straight into a database using computer tablets. Lithology, alteration, structure and mineralization data is collected; core recovery data is also measured. Core photos are routinely taken of all the core, both wet and dry, using a camera stand to ensure consistent photographs. Collar and downhole survey data is collected. Downhole surveys are taken using a Reflex GyroTM or a Devico Deviflex. Both of these multishot instruments are calibrated annually.
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A dataset of measured bulk densities from over 900 mineralized samples is used to inform the resource block model.
Sampling of the core is carried out on 1 m intervals or to geological contacts. The core is sawn using an automated core saw and half is bagged for dispatch, with the remainder being placed in the core box for storage. Drill core samples are routinely sent to the ALS facility in Romania. They are bagged and packed in large sealed wooden bins before being trucked to ALS. The sample rejects are returned to the mine site in the same bins. The samples are prepared for assaying at the ALS facility.
All samples were assayed for gold by 30 g fire assay with an AAS finish, with Au values above 10 ppm determined by a gravimetric finish. Multi-element determination was carried out by inductively coupled plasma mass spectrometry (ICP-MS) analysis and / or inductively coupled plasma emission spectroscopy (ICP-ES) analysis.
Eldorado employs a comprehensive QA/QC program as part of the assaying procedure, involving regular insertion of Certified Reference Materials (CRMs), duplicates and blank samples. Site geologists regularly monitor the performance of CRMs, blanks and duplicates as the assay results arrive on site. In Eldorado’s opinion, the QA/QC results demonstrate that the Olympias mine’s assay database is sufficiently accurate and precise for the resource estimation.
An important measure of performance at any producing mine is reconciliation of the block model to the final mill production figures, adjusted for stockpiles as necessary. The reconciliation conducted at Olympias is detailed and thorough. It is currently providing a quarterly snapshot and demonstrating that the block model, and thus the mineral resources, are valid and robust. This validates the data underpinning the model and is, by association, a good verification of the work done.
1.4
Metallurgical Testwork
Historical and independent metallurgical test reports on samples from the Olympias deposit were reported to be commissioned by previous mine owners. The 2011 Technical Report concluded that, in general, the testwork confirmed the anecdotal evidence from historic operations that the ore responds well to flotation with high recoveries.
In 2015, further metallurgical testwork and mineralogical investigations were carried out on Fresh Ore and Old Ore samples from the Olympias deposit. The main sulphide minerals in these samples were galena, sphalerite, pyrite, and arsenopyrite. Other than quartz, all other minerals can be considered minor. The liberation of all sulphide minerals was excellent at a P80 of 120 µm. Flotation of galena, sphalerite and pyrite / arsenopyrite in a sequential flowsheet was found to be effective at producing lead, zinc and gold concentrates.
Lead flotation was found to work well with Aerophine 3418 A, with lime and sodium cyanide the only depressants needed. Copper activation and zinc flotation were found to be optimum at a pH of 11.8 with CuSO4 and SIPX. Flotation of pyrite and arsenopyrite was effective at a pH of 6.0 controlled with sulphuric acid and SIPX.
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The rejection of arsenic and antimony were found to be problematic in the Pb concentrate in different samples. It was found to be extremely difficult to reduce the content of arsenic in the Pb concentrate from Fresh Ore to less than approximately 3%. Antimony was not an issue with the Fresh Ore, but was identified with the Old Ore as being present in the Pb concentrate at a grade of approximately 3%, although arsenic was not an issue.
Tests confirmed good settling-thickening characteristics for the concentrates and tailings samples. Vacuum filtration of the Tails underflow sample showed high throughput and low residual cake moisture, which ranged from 13.6% to 22.3%. Pressure filtration on concentrate underflow samples indicated high throughput and low residual moisture.
Acid Base Accounting (ABA) testing indicated uncertain acid generation potential for the Olympias tailings; NAG testing reported no net acidity generated. Analysis of the NAG liquor reported arsenic at an environmentally significant concentration.
1.5
Mineral Resources
Mineral resource estimates for Olympias mine were made from a 3D block model utilizing MineSight 3D software. Project limits, in UTM coordinates, are 478105 to 479700 East, 4491165 to 4493480 North and -800 to +60m elevation. Cell size for the project was 5m east x 5m west X 5m high.
A grade based discriminant was developed to allow for more consistent interpretations to be made. This was accomplished by creating a simplistic value formulae based on the logic of a Net Smelter Return (NSR) formula that used a combination of metal prices and metal recoveries to act as weighting factors against each metal. This metric, a dollar value, proved to be an excellent surrogate for a comprehensive equivalent grade. Inspection of these resource defining values (RDV) showed that for the parameters used, a value of $50 best defined what one would classify as likely economically mineralized zones.
For the Olympias modeling, the deposit was divided into three zones: East, West, and Flats. Within each of these zones, modeling domains were created using the $50 RDV. Assays and composite samples were tagged by these domain shapes ahead of data analysis and grade interpolation. The assays were top-capped prior to compositing and were composited into 1 m composites within the wireframes.
Grade estimates for Au, Ag, As, Pb, Zn and Fe were interpolated using an inverse distance to the power of 4 (ID4) method. Nearest–neighbour (NN) grades were also interpolated as a declustered distribution to validate the estimation method. A multi-pass approach was instituted for interpolation. The first pass required a grade estimate to include composites from a minimum of two holes from the same estimation domain, whereas the second pass allowed a single hole to place a grade estimate in any uninterpolated block from the first pass. The metal models were validated by visual inspection, checks for global bias and local trends and for appropriate levels of smoothing (change-of-support checks).
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The mineral resources of the Olympias mine were classified using logic consistent with the CIM definitions referred to in the National Instrument 43-101-Standards of Disclosure for Mineral Projects (NI 43-101). The mineralization of the mine satisfies sufficient criteria to be classified into measured, indicated, and inferred mineral resource categories.
Olympias mine mineral resources, as of September 30, 2019, are shown in Table 1-1. The Olympias mine mineral resources are reported at a zero cut-off within wireframes based on $50 RDV.
Table 1-1: Olympias Mine Mineral Resources as of September 30, 2019

Classification | Tonnes (Kt) | Au (g/t) | Au (Koz) | Ag (g/t) | Ag (Koz) | Pb (%) | Pb (Kt) | Zn (%) | Zn (Kt)
Measured | 2,702 | 10.93 | 950 | 156 | 13,552 | 5.0 | 135 | 6.0 | 162
Indicated | 11,779 | 7.52 | 2,848 | 134 | 50,746 | 4.6 | 542 | 6.2 | 730
Measured and Indicated | 14,481 | 8.16 | 3,798 | 138 | 64,298 | 4.7 | 677 | 6.2 | 892
Inferred | 3,720 | 7.98 | 954 | 137 | 16,385 | 3.9 | 145 | 4.0 | 149

1.6
Mineral Reserves and Mining
The mineral reserve estimates conform to CIM Definition Standards (2014). All design and scheduling has been completed using the mineral resource model and estimate described in Section 14. Only measured and indicated resources have been used for mineral reserves estimation. The estimation assumes that the mining methods employed at the mine will be drift and fill (DAF) and transverse longhole open stoping (TLHOS).
The cut-off values supporting the estimation of underground mineral reserves were developed in 2018 and based on future projected operating costs at a steady-state production rate of 650,000 tonnes per annum. The operating cost assessment indicated that NSR values of $133/t for DAF mining and $116/t for TLHOS mining would adequately cover all site operating costs on a breakeven basis. The weight averaged operating cost can be estimated at $125/t considering the balance between TLHOS and DAF. These DAF and TLHOS cut-offs were used to create potentially mineable stope shapes from the NSR block model (NSR BM). The NSR BM was created by Eldorado and is based on metallurgical recovery experience and historical sales and includes recognition of transport and refining costs.
In the evaluation of underground mineral reserves, modifying factors were applied to the tonnages and grades of all mining shapes to account for dilution and ore losses. In the DAF stopes, a mining dilution factor of 13% and a mining recovery of 98% were estimated. In the TLHOS stopes a mining dilution factor of 14% and a mining recovery of 95% were estimated.
The mineral reserve estimate is summarized in Table 1-2 and has an effective date of September 30, 2019.
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Table 1-2: Olympias Mineral Reserves as of September 30, 2019

Class | Tonnes (kt) | Au (g/t) | Au (Moz) | Ag (g/t) | Ag (Moz) | Pb (%) | Pb (kt) | Zn (%) | Zn (kt)
Proven | 2,601 | 9.19 | 0.77 | 133 | 11.1 | 4.3 | 112 | 5.1 | 133
Probable | 10,324 | 6.47 | 2.15 | 115 | 38.2 | 4.0 | 413 | 5.3 | 547
Total | 12,925 | 7.02 | 2.92 | 119 | 49.3 | 4.1 | 525 | 5.3 | 680

Notes:
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Mineral reserves are included in measured and indicated mineral resources.
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Figures in the tables may not compute due to rounding.
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The mineral reserves are based on a planning cut-off grade of $133/t for DAF and $116/t for TLHOS, and $50/t for development ore.
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Cut-off grades are based on a gold metal price of $1,250/oz, silver metal price of $16/oz, zinc metal price of $2,400/t, and lead metal price of $2,000/t.
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Metallurgical recoveries are based on feed grade and metallurgical algorithms.
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Exchange rate used is €1.20 = US$1.00.
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Average mining dilution and mining recovery factors of 14% and 95%, respectively, for LHOS, and 13% and 98%, respectively, for DAF are assumed.
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Mineral reserves are reported on a 100% ownership basis; Eldorado owns 95% of Hellas Gold.
The Olympias mine is a 100% underground (UG) mining operation extracting ore from three zones: East, West and Flats and two sub-zones: Remnants, and North. These are shown in Figure 1-1 as an isometric view of the Olympias underground zones looking north-west.
Mining is currently at a rate of 360 ktpa. There is a production increase planned to take place between 2019 and 2023 to a steady-state value of 650 ktpa.
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Figure 1-1: Isometric View of Olympias Underground Zones
Ground support is a combination of shotcrete, split sets and swellex bolts of varying lengths. All mined out areas are backfilled either with paste fill or cemented aggregate fill.
Mining at Olympias will be a combination of drift and fill (DAF) and transverse longhole open stoping (TLHOS). TLHOS will be confined to areas of the Flats where geometry and ground conditions support the use of the higher productive method. The remainder of the mine will use DAF.
TLHOS will be excavated 10 m wide and 30 m high maximum. The maximum length varies depending on the height and average rock quality. Blind uppers will be used for drill and blast with no top accesses.
DAF mining utilizes the overhand mining method. Stopes are accessed on the foot wall side from the main ramp starting at the bottom of each 20 m high stoping block. Each lift is mined 5 m high, with each panel limited to 5 m wide. There are two methods of DAF stoping, longitudinal and multi-pass.
Longitudinal is utilized where the stopes are less than 10 m wide. A single pass is mined down the length of the stope and any remaining ore is slashed off the HW on retreat. The stope is then filled prior to taking down the back in the access for the next lift.
Multi-pass is utilized where the stopes are greater than 10 m wide. Mining starts by cross cutting the ore body from footwall to HW. Then a single drift is developed in the ore to the extents of the stope in both directions along the footwall. Transverse cuts are then made from footwall to HW at multiple locations such that multiple working faces are established. After each cut is made it is backfilled, and then a new cut is mined beside the previous one. Once the lift is completely mined out, the back of the access is taken down to start the next lift.
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All stopes are filled with backfill after excavation. Currently a combination of cemented aggregate fill (CAF) and paste fill is used. The paste fill system has been designed to produce 42 m3/hr of paste, which will meet all future backfill requirements at 650 ktpa production with 70% utilization. CAF is delivered to stopes by truck and pushed into place with loaders. Paste is delivered with positive displacement pumps via drill holes and pipes.
There are two declines currently in use, one accessing the West Zone down to the Flats and one accessing the East Zone down to the Flats. There are multiple cross-over drifts between the two declines. Both declines are currently being extended into the Flats and to the bottom of the mine.
Both ore and waste are hauled to surface utilizing 40 tonne haul trucks on the existing and expanding declines. This will continue to be the case after the production increase to a steady-state value of 650 ktpa.
The total of personnel currently working at the mine is 562. The mine operates 3 shifts a day, 20 shifts per week.
There is currently 23 large pieces of mobile mining equipment on site: four jumbos, two bolters, five trucks, six loaders, four transmixers and two shotcrete sprayers. To achieve the production increase to 650 ktpa, funding has been allocated to increase this fleet number to 33. The increase will consist of one jumbo, two bolters, two longhole drills, two trucks, and three loaders.
The ventilation system consists of a single exhaust raise with fan. Air intake is via the two declines, the shaft and the old workings. Two means of egress are provided by the two declines. Current flow is 115 m3/s; this will increase to 360 m3/s for the 650 ktpa production rate.
Currently packaged explosives are being used for all blasting. There are no magazines on site and explosives are brought to site daily by the supplier. The use of bulk explosives is being investigated as an opportunity. The construction of a new underground magazine is planned for 2020. Steady-state full production explosives consumption is estimated at 53 tonnes per month.
As an operating mine, infrastructure is well developed with existing process water, compressed air, electrical distribution, and dewatering systems. For the 650 ktpa expansion, a new compressor, dewatering station and underground shop is being installed. These activities are currently in progress.
1.7
Processing
The Olympias lead-zinc-gold-silver process plant has been developed in phases:
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Phase I involved the recommissioning of the plant, after a prolonged inactivity, and processing of the existing tailings. Phase I commenced in 2013 and was completed on commissioning of Phase II.
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Phase II was commissioned in late 2017 and is currently in operation. It included refurbishment and upgrading of the process facilities to process 430 ktpa of ore. The Phase II process facility consists of comminution, flotation and filtering to produce three saleable concentrates: lead / silver (lead), zinc and arsenopyrite / pyrite gold (gold). All concentrates are sold to worldwide markets. Tailings are used for underground backfill via the on-surface paste plant. Any tailings not used for underground mine backfill are filtered and trucked from the Olympias processing facility to the Kokkinolakkas tailings management facility (TMF) over public roads. Phase II will be in operation until commissioning of the Olympias Phase III expansion.
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The expansion Project involves upgrading of the existing Olympias process plant to handle a mine feed rate of 650 ktpa of ore, and upgrades to the port facilities at Stratoni. A permit is required from the Government of Greece to extend the mine life of the existing Phase II facilities and to increase the processing rate to 650,000 tpa. A further expansion may be required. Studies are currently underway to extract more value from the gold concentrate that is currently being sold to traders and smelters around the world. If a viable extraction method is developed, then construction of a new metallurgical facility would follow.
The treatment facility incorporates the following unit process operations:
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Three-stage crushing to produce 80% material passing 13 mm ore.
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Ore storage between crushing and grinding in a fine ore bin with a 1,155 t live capacity.
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Single-stage ball milling in closed circuit with hydrocyclones to produce 80% material passing 120 µm.
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Flash flotation to remove high grade lead from the recirculating load.
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Lead flotation employing the following circuits:
o
Roughing and scavenging.
o
Regrind of flash flotation and rougher / scavenger concentrate to 80% passing 15 µm size.
o
Three stages of cleaning and one stage of cleaner scavenging (in open circuit).
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Zinc flotation employing the following circuits:
o
Roughing and scavenging.
o
Regrind of rougher / scavenger concentrate to 80% passing 15 µm size of rougher / scavenger concentrate.
o
Three stages of cleaning and one stage of cleaner scavenging (in open circuit).
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Gold-pyrite flotation utilizing roughing and scavenging and a single stage of cleaning.
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Concentrate thickening, filtration, packaging, and storage prior to dispatch from the mine site by road.
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Tailings thickening and filtration prior to direct or reclaim addition to a cemented backfill plant.
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Tailings paste backfill.
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Reagent mixing, storage and distribution.
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Water and air services.
1.8
Infrastructure
As an operating mine, current infrastructure is robust and complete. The mine has access to the main highway system in Greece via paved roads to the mine site. Local services are provided via the towns of Olympiada and Stratoni, with additional services available through Thessaloniki.
Zinc concentrate is shipped via the port facility at Stratoni (owned by Hellas Gold). Lead and arsenopyrite concentrates are shipped via Thessaloniki. There is a plan in place to rehabilitate and upgrade the Stratoni port facility over the period from 2020 to 2023; this will allow for the shipping of all concentrates by bulk out of Stratoni with the associated cost savings.
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Water for the mine is obtained from underground dewatering, after treatment. Excess water from underground is discharged into the Mavrolakkas stream after settling and treatment to meet discharge standards. Currently, the capacity to handle 400 m3/hr is available; this is being increased to 650 m3/hr, which will be sufficient for the mine life. Service water is supplied via a local borehole in the regional aquifer.
Waste rock is either recycled underground for fill or is disposed of in the existing waste disposal facility. Tailings not used for pastefill are dewatered to 13% moisture content and transported by truck to the new tailings management facility at Kokkinolakkas near the Stratoni facilities, about 23 km by public road from the mine.
Existing surface facilities consist of a surface workshop, administration building, dry, shaft, and fuel storage (60,000 litres capacity). The workshop and fuel storage will be adequate for the production increase. The shaft is used for inspection of a legacy pump station only and there are no plans to rehabilitate the shaft further. Construction of a new geology preparation laboratory and technical services building, and expansion to the existing administration building are to be completed as part of the expansion.
Current power to site consists of a 20 kV 10 mVA pole line from the PPC grid. To facilitate the production increase, a new pole line at 150 kV 25 mVA, along with a new substation, will be constructed over the years 2020 and 2021. Backup power consists of 4,920 kW of diesel generation in multiple distributed generators. An additional 2,500 kW of generated power will be added for the production increase.
1.9
Market studies and contracts
The Olympias plant produces three concentrates: a pyrite concentrate containing gold, a lead concentrate that also contains silver and a zinc concentrate. Hellas Gold has negotiated multiple concentrate sales contracts with commodity traders.
Hellas Gold sells gold, lead and zinc concentrates to various customers from its Olympias operation. Currently contracts are used for short term revenue assumptions and long term averages are used over the life of mine.
1.10
Environmental
The EIS for the Kassandra mines mineral deposits project (Kassandra Project) includes an area of 26,400 ha in north-eastern Halkidiki (Macedonia Region). The Kassandra Project includes the Skouries, Olympias and Stratoni sites. No significant impact is expected on the landscape, geological environment, atmosphere, or water resources in the area. The overall impacts to date have been positive to the environment, as legacy tailings and concentrate storage are in the process of being removed to the new TMF facility at Kokkinolakkas, and the associated areas rehabilitated. The Kassandra Project provides significant economic and social impacts for the Halkidiki Prefecture, including:
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A significant contribution is made to the national economy.
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Significant infrastructure is constructed and equipped by local companies.
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Service industries in the local economy expand.
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New jobs are progressively created.
After the completion of all Kassandra operations, the project areas will be rehabilitated according to appropriate and approved land uses. All structures are to be removed or left in a state that they do not pose a risk to the environment or public. The environment will be returned to a state of a self-sustaining ecosystem and safe and stable biological conditions will be re-created.
Hellas Gold has provided a letter of guarantee for €50M in favour of the Greek State as an assurance that the funds necessary for rehabilitation works in relation to the mining and metallurgical facilities of the Kassandra mines project and the removal, cleaning and rehabilitation of the old disturbed areas from the historic mining activity in the wider area of the project. Additionally, a Letter of Guarantee to the MOE, in the amount of €7.5M, has been provided as security for the due and proper performance of the Kokkinolakkas TMF.
1.11
Capital and Operating costs
The total capital cost estimate includes the expansion to 650 ktpa in the process plant, as well as capital development of the underground mine and sustaining capital, which is primarily for development, mine, processing infrastructure, and maintenance. Expansion capital specifically includes process plant upgrades, water management facilities, Phase 1 and Phase 2 of the Stratoni port facility upgrade as allocated amongst the Kassandra assets, related indirect costs, EPCM, owner’s and contingency costs. Capital and sustaining capital costs are summarized in Table 1-3.
Table 1-3: Capital Cost Summary

Area | Growth ($ x 1,000) | Sustaining ($ x 1,000) | Total($ x 1,000)
Mining Dev't and Equipment | 9,616 | 89,471 | 99,087
Mine Infrastructure | 3,876 | 49,161 | 53,037
Process Plant | 18,892 | 14,356 | 33,248
Tailings + Water | 9,457 | 17,630 | 27,087
Surface Supports | 5,739 | 2,056 | 7,795
Stratoni Port Facility | 15,478 | 5,700 | 21,178
Others | 1,566 | 2,829 | 4,395
Total Direct Cost | 64,624 | 181,203 | 245,827
Indirect Cost | 6,894 | 10,990 | 17,884
Spares and Fills | 3,153 | - | 3,153
EPCM | 6,306 | 3,060 | 9,366
Owners Costs | 3,153 | - | 3,153
Contingency | 15,765 | - | 15,765
Total Installed Costs | 99,894 | 195,253 | 295,147

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Operating costs include allocations for underground mining, processing, tailings filtration, concentrate transport, tailings filtrations, materials handling, and stacking, water management, and general and administration (G&A).
Operating costs were estimated for each year of operation and are summarized in Table 1-4.
Table 1-4: Operating Costs Summary

Category | LOM ($M) | LOM average ($/t ore) | Steady State: 2024 to 2038 ($/t ore)
Mining costs | 908 | 69.6 | 48.9
Tailings and water management costs | 56 | 4.4 | 4.4
Processing costs | 652 | 50.4 | 50.4
G&A costs | 310 | 23.9 | 21.3
Total operating cost | 1,925 | 148.3 | 125.0

Current labour productivities are considered very low when compared to other international operations. Eldorado believes that the envisaged improvement initiatives are reasonable and achievable and has allowed for effective work hours per shift to gradually reach 5.5 hrs. This will be assisted by the increase in the number of available faces for DAF mining through design and sequence changes. The introduction of more cost-effective THLOS mining for approximately one third of the steady-state tonnage is also a major factor in the future operating costs schedules. All of these initiatives should be in place as steady state by 2024. To illustrate the positive impact of these initiatives on LOM operating costs, a separate column was created in Table 1-4 showing estimated costs from 2024 to 2038.
1.12
Economics
The economic model has been prepared on an annual life of mine basis. The effective date of the estimate is assumed to be December 31, 2019. The LOM for the current mineral reserves is projected to be 21 years.
The after-tax cash flow analysis shows that Olympias is a robust Project once the additional capital is applied and the project reaches a steady state level of production. The net present value (NPV) of the Project is estimated to be $1,067M, using a discount rate of 5%, with positive after-tax net cash flows projected in all years other than 2019, the initial year of expansion capital expenditure.
Key results for the project economics are provided in Table 1-5.
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Table 1-5: Key Economic Results

Olympias | Unit | Value
Total ore milled | kt | 12,925
Net revenue | $M | 4,501
Total Capital costs | $M | 295
Operating costs (total) | $M | 1,925
Transport, treatment and refining costs | $M | 408
Royalties | $M | 67
Corporate tax | $M | 423
Net post-tax cash flow | $M | 1,866
Post-tax NPV @ 5% discount rate | $M | 1,061
Post-tax NPV @ 8% discount rate | $M | 783

The economic model was subjected to a sensitivity analysis to determine the effects of changing metal prices, and of operating and capital cost assumptions on the projected financial returns. The results of the sensitivity analysis are provided in Table 1-6 and Table 1-7
The test of economic extraction for Olympias is demonstrated by means of this sensitivity analysis. At the mineral reserve metals prices stated, Olympias shows strong positive economics. The sensitivity analysis also shows that the Olympias economics are robust when evaluated using lower metal price assumptions, or higher operating and capital costs.
Table 1-6: Metal Price Sensitivity Analysis

Sensitivity Ranges
Parameters | Units | -15% | -7.5% | Project case | +7.5% | +15%
Gold price | $/oz | 1,200 | 1,300 | 1,400 | 1,500 | 1,600
Silver price | $/oz | 15.0 | 17.0 | 18.0 | 19.0 | 21.0
Lead price | $/t | 1,800 | 1,900 | 2,100 | 2,300 | 2,400
Zinc Price | $/t | 2,000 | 2,200 | 2,400 | 2,600 | 2,800
Results (after tax)
NPV 0% | $M | 1,327 | 1,604 | 1,866 | 2,164 | 2,415
NPV 5% | $M | 724 | 897 | 1,062 | 1,241 | 1,395
NPV 8% | $M | 519 | 655 | 783 | 925 | 1,046
Taxation | $M | 248 | 336 | 423 | 512 | 591
Royalties | $M | 42 | 52 | 66 | 76 | 120

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Table 1-7: Capital and Operating Cost Sensitivity Analysis

Sensitivity Ranges
Parameter | Units | -15% | Project case | 15%
LOM Capex | $M | 256 | 295 | 347
NPV 0% (after tax) | $M | 1,927 | 1,866 | 1,836
NPV 5% (after tax) | $M | 1,105 | 1,062 | 1,030
NPV 8% (after tax) | $M | 823 | 783 | 755
LOM operating costs | $/t ore | 126 | 148 | 170
NPV 0% (after tax) | $M | 2,100 | 1,866 | 1,663
NPV 5% (after tax) | $M | 1,205 | 1,062 | 929
NPV 8% (after tax) | $M | 899 | 783 | 678

1.13
Conclusions and recommendations
Olympias is a mine with a long history that has only been recently put back into production by Eldorado. Commercial production was achieved in December 2017. Project economics are robust. There are sufficient mineral reserves for a mine life of 21 years at a steady-state production rate of 650 ktpa. The mine has embarked on a series of initiatives to increase both productivity and production rates.
Key recommendations are as follows:
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Expand extents of mineralization in upper levels of the mine area through conversion of Inferred resources and exploration. Any additional ore in this region would have synergies with the existing infrastructure.
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Use of ore sorting or dense media technology to pre-concentrate the ore coming from underground before it is processed in the grinding and flotation circuits with the intention of maximizing the annual metal production from the Olympias mine
●
Ongoing research into an alternative processing method to extract additional value from the gold concentrate currently being sold directly to smelters and traders worldwide. Technical solutions do exist for extraction of much of the contained gold but these need to be able to demonstrate acceptable rates of return and to comply with environmental and safety requirements. If a viable extraction method is developed, then construction of the new metallurgical facility would follow.
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SECTION ● 2
Introduction
Eldorado Gold Corporation (Eldorado), an international gold mining company based in Vancouver, British Columbia, owns and operates the Olympias mine in Greece through its 95% owned subsidiary Hellas Gold SA (Hellas Gold). Eldorado has prepared this technical report of the Olympias mine to provide an updated description of the geology and mineralization, mineral resources and mineral reserves, and mine and mill operations in light of the long shelf life of the existing technical report, whose effective date was July 14, 2011 (Technical Report on the Olympias Project Au Pb Zn Ag Deposit, Northern Greece, 2011).
Information and data for this report were obtained from the Olympias mine and Eldorado’s corporate office. The work entailed review of pertinent geological, mining, process and metallurgical data in sufficient detail to support the preparation of this technical report.
When preparing reserves for any of its projects, Eldorado uses a consistent prevailing metal price methodology that is in line with the CIM Guidance on Commodity Pricing used in Resource and Reserve Estimation and Reporting. These were set for gold, silver, lead and zinc, as of September 2019 for Eldorado’s current mineral reserve work, at US$ 1,250/oz Au. US$ 16.00/oz Ag, US$2,000/t Pb and US$2,400/t Zn. All cut-off grade determinations, mine designs and economic tests of extraction used these prices in this technical report. In order to demonstrate the potential economics of a project, Eldorado may elect to use metal pricing closer to the current prevailing spot price and then provide some sensitivity around this price. For the Olympias gold mine, metal prices used for the potential economics evaluation were US$1,400/oz Au, US$18.00/oz Ag, US$2,100/t Pb and US$2,400/t Zn. This analysis (in Section 22 of this report) generally provides a better ‘snapshot’ of the project value at prevailing prices rather than limiting it to reserve prices, that might vary somewhat from prevailing spot prices. Eldorado stresses that only material that satisfies the mineral reserve criteria is subjected to further economic assessments at varied metal pricing.
The qualified persons responsible for preparing this technical report as defined in National Instrument 43-101 (NI 43-101), Standards of Disclosure for Mineral Projects and in compliance with 43-101F1 (the “Technical Report”) are David Sutherland, P.Eng., Paul Skayman, FAusIMM, Ertan Uludag, P.Geo., Colm Keogh, P.Eng. and Sean McKinley, P.Geo., whom are all employees of Eldorado.
David Sutherland, Project Manager, was responsible for overall preparation of the technical study and sections related to infrastructure and environment (report sections 1, 2, 3, 4, 5, 6, 18, 20, 24, 25, 26 and 27). He most recently visited the Olympias mine on October 2, 2019.
Colm Keogh, Operations Manager, Olympias Mine, was responsible for the mineral reserves and the preparation of related sections on mineral reserves calculation, mining methods and sections related to costs (report sections 15, 16, 21 and 22). He has worked at the Olympias mine since July 2019.
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Sean McKinley, Senior Geologist Resource Development was responsible for the preparation of the sections in this report concerned with geological information, exploration and drilling (report sections 7, 8, 9, 10 and 23). He most recently visited the Olympias mine on July 2 to 5, 2019.
Ertan Uludag, Resource Geologist, was responsible for the mineral resources and the preparation of related sections on sample preparation and analyses, data verification and mineral resource estimation (report sections 11, 12 and 14). He most recently visited the Olympias mine on August 21 to 27, 2019.
Paul Skayman, Special Advisor to the COO, was responsible for the preparation of the sections in this report that dealt with metallurgy and process operations and related costs and payability (report sections 13, 17 and 19). He most recently visited the Olympias mine on November 11, 2019.
This document presents a summary of the current and forecast operations at the mine.
Currency used is US$ throughout, unless stated.
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SECTION ● 3
Reliance on Other Experts
The Qualified Persons have relied, in respect of legal aspects, upon the work of the experts listed below. To the extent permitted under NI 43-101, the QPs disclaim responsibility for the relevant sections of the technical report.
The following disclosure is made in respect of this Expert:
●
Public Notary Ioanna V. Gabrieli-Anagnostalakis, on behalf of the Minister of Finance and the Minister of Development of the Government of Greece
Report, opinion, or statement relied upon:
●
Transfer Contract no. 22.138/12-12-2003 between the Greek State and Hellas Gold, specifically Appendix III which consists of Table 1, list of concessions.
Extent of reliance:
●
Full reliance
Portion of Technical Report to which disclaimer applies:
●
Section 4.2 Land Tenure
The following disclosure is made in respect of this Expert:
●
ENVECO S.A., Environmental Protection, Management and Economy S.A
Report, opinion, or statement relied upon:
●
ENVECO S.A., 2010, Environmental Impact Assessment (EIA) of the Mining-Metallurgical Facilities of Company Hellas Gold in Halkidiki
Extent of reliance:
●
Full reliance
Portion of Report to which disclaimer applies:
●
Section 20
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SECTION ● 4
Property Description and Location
4.1
Property Location
The Property is located within the Kassandra mines complex found within the Halkidiki Peninsula of the Central Macedonia Province in Northern Greece. The complex is comprised of a group of mining and exploration concessions, covering 317 km2, located approximately 100 km east of Thessaloniki. The concessions include the Olympias Mine, the Madem Lakkos and Mavres Petres mines (collectively known as Stratoni), and the Skouries copper-gold porphyry deposit. The Olympias and Mavres Petres mines are currently in production and the Skouries deposit is under development.
Olympias lies 9 km north-northwest of the Stratoni port and loading facility, accessed by a paved road along the coast. The area is centered on co-ordinates 474,000 E and 4,488,000 N of the Hellenic Geodetic Reference System HGRS ‘80, Ellipsoid GRS80 (approximately Latitude 40° 36’ E and Longitude 23°50’ N). Figure 4-1 shows all the concessions of the Kassandra mines complex in beige, with the Olympias property highlighted in red.
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Figure 4-1: Property Concessions
The Property is owned by Hellas Gold, with Eldorado holding a controlling interest of 95% in Hellas Gold.
4.2
Land Tenure
The Project is located within concession numbers F13 and F14, which have a combined area of 47.27 km2 (Table 4-1). Hellas Gold has been granted mining rights over these concessions until April 7, 2024. The concessions are conditionally renewable for a further two consecutive periods of 25 years each. Hellas Gold has ownership of a small portion of private land within the concessions.
Table 4-1: List of Concessions

Concession # | Title holder | Tenement type | Area (km2) | Expiry date
F13 | Hellas Gold | Mining Concession | 19.47 | April 7,2024
F14 | Hellas Gold | Mining Concession | 27.80 | April 7, 2024
Total |  |  | 47.27 |

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4.3
Permitting
In July 2011, the Ministry of Environment (MOE) formally approved the Environmental Impact Statement (EIS) submitted by Hellas Gold for the three Kassandra Mines mine sites, being Olympias, Skouries and Stratoni, which involves an area of 26,400 ha, in north-eastern Halkidiki (Macedonia Region). This EIS is valid for 10 years and subject to renewal in 2021.
For production to commence, the MOE required the submission of a technical study. A study was submitted to the MOE and approved in early 2012. The installation permit for what was termed the Phase II process plant was issued on 22 March 2016. Installation work was completed in May 2017, at which time commissioning and trial production commenced. The Company received the operating permit for the Phase II plant in September 2017, allowing commencement of commercial production operations. In September 2017, the Company also received an extension of the installation permit and an interim operating permit for the Kokkinolakkas tailings management facility (TMF), as well as the delayed installation permit for the paste backfill plant.
Hellas Gold has provided a €50.0 M Letter of Guarantee to the MOE as security for the due and proper performance of rehabilitation works in relation to the mining and metallurgical facilities of the Kassandra mines project, and the removal, cleaning and rehabilitation of the old disturbed areas from the historic mining activity in the wider area of the project. Additionally, a Letter of Guarantee to the MOE, in the amount of €7.5 M, has been provided as security for the due and proper performance of the Kokkinolakkas TMF.
Notifications for the Operation of the Olympias paste plant and Kokkinolakkas TMF were formally submitted in 2018 and remain in force in line with new legislation that replaced previous operating permit issuance procedures.
4.4
Royalties and Taxes
Based on current Greek legislation, royalties are applicable on active mining titles. The royalty is calculated on a sliding scale tied to metal prices. At the price index selected for the Project evaluation, Hellas Gold would pay a royalty of approximately 2% on Au, 1% on Ag, 0.5% on Pb, and 1% on Zn revenues.
The corporate income tax rate for Greek companies in 2018 was 29%. This is legislated to drop by 1% per annum until it reaches 25% in fiscal year 2022.
4.5
Environmental Liabilities
The closure and environmental rehabilitation activities for Kassandra Mine Mineral Deposits Project relate to the following facilities:
●
Open pit and underground mine
●
Integrated waste management facility (IWMF)
●
Process facilities and infrastructure
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To meet the requirements of the reclamation program, decommissioning, closure, and reclamation of the affected areas must be undertaken.
Hellas Gold has provided €57.5 M in Letters of Guarantee to the MOE as security for the due and proper performance of rehabilitation works in relation to the mining and metallurgical facilities of the Kassandra Mines project and the removal, cleaning and rehabilitation of the old disturbed areas from the historic mining activity in the wider area of the project.
Hellas Gold has also provided insurance coverage in accordance with Presidential Decree 148/2009 (Government Gazette 190/A/29.9.2009) for environmental liability.
4.6
Other
To the extent known, there are no other significant factors and risks that may affect access, title, or the right or ability to perform work on the property. A permit is required from the Greek Government to extend the permitted operating life of the current Olympias facility as well as a permit to increase annual throughput to 650,000 tpa.
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SECTION ● 5
Accessibility, Climate, Local Resources, Infrastructure and Physiography
5.1
Location and Accessibility
The Property is located in the Halkidiki peninsula in the Northern Macedonia region about 100 km by road from Thessaloniki, the second largest city in Greece. Thessaloniki has one of the largest ports in Greece and an international airport (Figure 5-1).
The Property is readily accessible year-round by the national road network with paved secondary roads within 1 km of site. The national road network in the area is among the best in northern Greece, with a major highway (E90), extending east from Thessaloniki to approximately 15 km north of the property. Olympias lies 9 km north-northwest of the Stratoni port and loading facility, accessed by a paved road along the coast, and paved mine road to site.
Figure 5-1: Location of the Olympias Property
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5.2
Infrastructure and Local Resources
The area is well served by main 115 kV power from the Public Power Corporation (PPC), with power supply right to the mine site. Communications are good. Telephone and broadband are available in the area. Hellas Gold has the existing microwave link at Stratoni, giving additional access to broadband data communications.
Process water is reclaimed from the tailings thickening and filtration circuit and backfill clarifier overflow. A minimum amount of make-up water is supplied from the raw water tank. Potable water is produced from a borehole located in the regional aquifer.
Skilled labour, some of whom worked on the mine historically is available in the village of Olympiada, which is located about 1 km from the mine. Other nearby villages also have a pool of skilled and semi-skilled labour available. The company owns the all the land on which existing and proposed mine infrastructure is located, and this is sufficient to allow for future operation of the mine as designed, without further land acquisition.
The mine site comprises mine offices and change rooms, trackless decline, shaft, underground development, and processing consisting of crushing plant, mill, flotation plant, and backfill plant.
Figure 5-2 shows the location of the Olympias operations and the Kokkinolakkas TMF, which is on the Stratoni property and common to both operations.
Figure 5-2: Olympias Site Layout
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5.3
Climate and Physiography
The Halkidiki Peninsula climate is generally mild with limited rainfall. Over 300 days or around 3,000 hours of sunshine are recorded annually. Average temperatures have limited fluctuations during the year. The lowest temperatures occur during December to February, ranging between 3.5°C to 19°C, while the highest temperatures occur during summer months, ranging between 23°C and 34°C. Temperatures below 0°C are limited to the mountainous areas. Rainfall varies from 19 mm to 40 mm in a month, with October to December the wettest months. Operations can continue all year round.
The terrain is characterized by hills rising to about 600 metres above sea level (masl) with steeply incised valleys.
The area is wooded with oak, beech and pine being the principal species, while inland there are vineyards and farmlands. The main farming products are grapes, honey, and olives.
5.4
Surface Rights
Hellas Gold owns the land containing the entire surface infrastructure associated with the Olympias Project. This is sufficient to allow the future operation of the mine without further land acquisition.
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SECTION ● 6
History
6.1
Introduction
There is a long history of mining in the area. Ancient mining reached a peak during the time of Philip II of Macedon and Alexander the Great, in the period 350 to 300 BC. The lead-rich ores from the Madem Lakkos mine at Stratoni were smelted for silver and the Olympias ores were processed for their high gold content. It has been estimated, from the volume of ancient slags that about 1.0 Mt of ore were extracted from each locality during this period. It is believed that by 300 BC, the bulk of the ores above the water table at Olympias had been exploited, though the Stratoni mine continued in production through the Roman, Byzantine, and Ottoman periods.
6.2
Ownership and Work Carried Out
Milestones in the history of the Property are shown in Table 6-1.
Table 6-1: Summary of the history of the Property

Year | Commentary
Historic times | Bulk of ores at Olympias above water table were extracted by 300 BC.
1933 | Shaft sunk to 74 m depth with some drifting.
1954 | Owners commenced exploration; thin, discontinuous sulphide lenses encountered (and many ancient workings).
1965-66 | Further drilling intersected 10 m of lead-zinc mineralization 20 m below the 1933 shaft.
1970 | Ownership transferred to Hellenic Fertilizer Company; ramp was started, and production commenced in West Orebody.
1974-84 | Mine was developed to mine lead and zinc. Shaft was sunk to the -312 m level; high grade mineralization of East orebody intersected; highly profitable mining using sub-level caving; eventual transition to less profitable drift-and-fill mining due to excessive dilution, ground subsidence and water problems.
1991 | Hellenic Fertilizer Company went into receivership; mine continued production under subsidy from Greek government.
1995 | Ownership transferred to TVX Gold Inc. (TVX); production suspended to allow for drilling to define mineral resources.
1998-99 | TVX completed drilling campaign (760 holes, 91,319 m) and issued a Mineral Resource estimation; initial feasibility study completed.
2004 | Aktor acquired mining concessions holding 317 km2, including the Olympias and Skouries, deposits together with Stratoni (the Kassandra Mines) through its subsidiary Hellas Gold. The Hellas Gold acquisition of the Kassandra Mines was ratified by parliament and passed into law in January 2004 (National Law no. 3220/2004).
European Goldfields (EGL) acquired its initial ownership percentage interest in Hellas Gold from Aktor through its wholly owned subsidiary European Goldfields Mining (Netherlands) B.V.
2007 | EGL increased share ownership of Hellas Gold to 95% (with 5% held by Aktor).
2011 | EIS approved by Greek government.
2012 | Eldorado acquired the project via the acquisition of EGL. Commenced tailings re-treatment and rehabilitation of the underground mine shortly thereafter.

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6.2.1
Work Pre 1970
In 1933, a shaft was sunk to 74 m depth at Olympias by the then owners, with lateral development to intersect a marble / gneiss contact. The drift encountered ancient workings, which were largely exhausted.
In 1954, the owners commenced exploration in the Olympias area. Their drilling encountered numerous discontinuous lenses of sulphide mineralization 10 cm to 20 cm in width, and many old workings were intersected. In 1965, further drilling commenced and, in 1966, a 10 m intersection of lead‐zinc mineralization was encountered 20 m below the 1933 shaft, which proved to be exploitable.
6.2.2
Hellenic Fertilizer Company 1970 – 1991
By the early 1970s ownership of the mines had transferred to the Hellenic Fertilizer Company, (HFC). A ramp was started in 1970 and full production commenced in the West orebody. From 1974 to 1984 the shaft was sunk to 312 m below mean sea level (msl) and high-grade mineralization, termed the East orebody, was intersected at 254 m below msl. The first ten years of production proved highly profitable when sub‐level caving (SLC) could be used, but excessive dilution, ground subsidence and problems with water influx resulted in a change to more expensive drift and fill (DAF) mining. During this period lead and zinc concentrates were sold and gold-bearing pyrite-arsenopyrite concentrate was stockpiled. The stockpile of gold bearing pyrite-arsenopyrite concentrate was sold to off‐takers by EGL.
HFC carried out extensive programs of surface and underground drilling in order to define orebody dimensions and to explore the surrounding area. Partial logs are available for this work but none of the original cores are available in labelled boxes; none of the holes were surveyed and no assays, certifiable or otherwise, have been found. It is believed that ore was identified solely by visual assessment of the core. Where available, the partial logs of this work were entered by EGL into the database but only in order to guide exploration work and for use in the modelling of major geological units.
HFC went into receivership in 1991 and the Government subsidized the operation until development and production fell behind schedule. Following three separate international auctions, TVX successfully won the bidding for the Property in 1995.
6.2.3
TVX 1995 – 1999
Shortly after TVX obtained control of the Property, production at the Olympias Project was suspended so that drilling could be conducted to confirm and expand the Mineral Resources.
TVX commenced an intense program of drilling in 1996. An interim Mineral Resource assessment was completed in June 1998. By February 1999, TVX had completed a drill program comprising 760 holes totalling 91,319 m. The Mineral Resource was updated in December 1998 to incorporate all drilling completed to that date. In 1998 and 1999, Kvaerner Metals produced a feasibility study on the Olympias Project. During 1999 ‐ 2000, SNC Lavalin of Toronto was commissioned to conduct a basic engineering study on the Olympias Project.
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6.2.4
European Goldfields 2004 – 2014
The Property was acquired by EGL (through its 95% owned subsidiary Hellas Gold) in 2004. No exploration was carried out by EGL; however, sufficient reviews of existing data allowed EGL to file the pre-existing NI43-101 Technical Report with an effective date of July 14, 2011 (Technical Report on the Olympias Project Au Pb Zn Ag Deposit, Northern Greece, 2011).
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SECTION ● 7
Geological Setting and Mineralization
7.1
Regional Geology
The Western Tethyan orogenic belt in southeast Europe contains several major metallogenic provinces including the Serbo-Macedonian Metallogenic Province (SMMP) that hosts the Kassandra mining district in northeast Greece (Janković, 1997; Baker 2019). The Western Tethyan orogen comprises a series of magmatic belts encompassing Cretaceous to Paleogene subduction related arc magmatism through to post-collisional Neogene magmatism (Richards, 2015; Moritz and Baker, 2019). In northeast Greece the orogeny developed during the Late Cretaceous to early Eocene as the Serbo-Macedonian Apulian and Pelagonian microcontinents converged on the previously accreted Rhodope continental fragments at the Eurasian margin (Pe-Piper and Piper, 2006). Crystalline basement within the Kassandra mining district includes the upper litho-tectonic Serbo-Macedonian Vertiskos unit and the lower litho-tectonic Kerdilion unit exposed within the southern Rhodope metamorphic core complex (Figure 7-1).
The SMMP forms a northwest-trending zone of base and precious metal deposits including a large Au-endowment (~25 Moz; Baker, 2019) that is associated with Oligocene to Miocene magmatic complexes including porphyry (Skouries, Greece; Illovitza, Bucim, North Macedonia; and Tulare, Serbia) and carbonate replacement deposits (Olympias, Mavres Petres, Madem Lakkos and Piavitsa in Greece), as well as the Plavica high sulfidation epithermal deposit (North Macedonia). The mineral deposits formed during post-collision extension and emplacement of intermediate to felsic magmas with high K calc-alkalic to shoshonitic composition and localized ultra-potassic mafic magmas (Borojevic Sostaric et al., 2012; Siron et al., 2016). The heterogeneity of the Cenozoic magmas likely resulted from crystal fractionation, assimilation and mixing of melted depleted mantle metasomatized by Cretaceous subduction processes, and partial melting of lower crustal rocks.
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Figure 7-1: Regional Geology of the Kassandra Mining District (modified from Siron et al, 2018)
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7.2
Local Geology
The Kassandra mining district contains a geological endowment of > 8 Moz of gold in measured and indicated resources at the Olympias Mine (14.4 Mt @ 8.2 g/t Au, 138 g/t Ag, 4.7% Pb and 6.2% Zn; this report) and the Skouries deposit (289.3 Mt @ 0.58 g/t Au and 0.43 % Cu; Eldorado Gold Corp. Annual Information Form (AIF), March 29 2019, SEDAR website). Additional measured and indicated resources are reported for the Mavres Petres (Stratoni) Mine (0.8 Mt @ 185 g/t Ag, 7.2% Pb and 10.1% Zn; Eldorado Press Release November 15, 2019, SEDAR). The Stratoni fault is a major tectonic feature and important mineralizing corridor located in the center of the Kassandra mining district (Siron et al., 2016, 2018; Figure 7-1). The east-west oriented, ductile to brittle fault zone extends >12 km from the coast at Stratoni to the village of Varvara in the west. It is a moderately south-dipping normal fault that separates gneiss and marble of the Kerdilion unit to the north from amphibolite and schist of the Vertiskos unit to the south (Figure 7-1). The historically mined Madem Lakkos deposit, the currently mined Mavres Petres deposit, and the undeveloped Piavitsa deposits are hosted dominantly in marble within and adjacent to the fault. The fault zone crosscuts the lower portion of the late Oligocene (25.4 ± 0.2 Ma) Stratoni granodiorite stock but is cut by a Miocene glomerophyric monzonite porphyry dike at Piavitsa (20.62 ± 0.13 Ma) constraining latest major fault movement and related hydrothermal mineralization to the late Oligocene to early Miocene (Siron et al., 2016). The Olympias deposit occurs approximately 6km north of the Stratoni fault with sulfide ore occurring within marble layers of the Kerdilion unit adjacent to two major normal faults, the Kassandra fault and the East fault. Crosscutting structural relationships and 40Ar/39Ar geochronology on ore-related muscovite alteration (22.6 ± 0.3 Ma) indicate that carbonate replacement mineralization at Olympias also occurred during the latest Oligocene, coincident with the early stage of post-deformation magmatism in the region (Siron et al., 2018).
Metamorphic rocks of the Kerdilion unit consist of quartzo-feldspathic hornblende-biotite gneiss, marble, amphibolite, localized bodies of megacrystic plagioclase-microcline orthogneiss, and granite gneiss (Kalogeropoulos et al., 1989; Gilg and Frei, 1994). The lithologies have an arcuate geometry, striking in a north-south direction in the north and becoming east-west near the Stratoni fault (Siron et al., 2016). Middle Jurassic to Early Cretaceous zircon U-Pb and Pb-Pb ages from granitic gneisses of the Kerdilion unit range from 164-134 Ma and are interpreted as primary igneous ages (Himmerkus et al., 2011). The host rock sequence, however, is likely Carboniferous to Permian based on inherited zircon cores derived from the orthogneisses (Himmerkus et al., 2011). Pegmatitic dikes and sills occur throughout the Kerdilion unit and represent anatectic partial melting of the metamorphic rock from about the middle Paleocene to middle Eocene (Wawrzenitz and Krohe, 1998; Kalogeropoulos et al., 1989). The pegmatites are largely absent south of the Stratoni fault.
Graphitic garnet-bearing quartz-biotite gneiss and schist are spatially associated with the Stratoni fault zone and amphibolite occurs within the hanging wall with variably serpentinized pyroxenite. The Vertiskos unit occurs south of the Stratoni fault and hosts the Skouries porphyry deposit. The unit is a monotonous sequence of quartz-rich feldspathic to muscovite-biotite-bearing gneiss and schist. Minor calcareous schist, marble, and amphibolite are also thinly interlayered within the metamorphic sequence. Zircon U-Pb ages show that the micaceous schist ranges from Neoproterozoic (686-576 Ma) to Ordovician (464-450 Ma), which is consistent with the age of the Pan-African Pirgadikia and Vertiskos units of the Serbo-Macedonian terrane (Himmerkus et al., 2007).
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Cretaceous to mid-Eocene ductile deformation accompanied by lower amphibolite-grade metamorphism and overprinting retrograde greenschist metamorphism affected the Kerdilion and Vertiskos units. A regionally prominent penetrative shallow dipping S1 foliation is defined by alignment of peak metamorphic minerals (e.g., biotite or amphibole). Subsequent high-strain transposition resulted in tight to isoclinal F2 folds locally accompanied by axial planar S2 cleavage. A later lower-strain deformation event superimposed a spaced steeply dipping S3 foliation on the pre-existing fabrics. This event is associated with km-scale upright and open east-plunging F3 folds evident as district-scale antiforms in the footwall of the Stratoni fault (Siron et al., 2016).
A series of discrete magmatic events are recognized in the region including the Triassic Arnea granite suite (228 ± 5.6 Ma) within the Vertiskos unit, and Late Cretaceous to early Eocene (68 ± 1 Ma to 53 ± 4 Ma) subduction-related calc-alkaline granites at Ierissos, Ouranoupolis, and Grigoriou on the Athos Peninsula. These latter granitic intrusions exhibit a weak tectonic fabric suggesting emplacement during the waning stages of regional deformation. Post-collision Oligocene-Miocene magmatism coincided with the main mineralizing events in the Kassandra mining district. Late Oligocene magmatism ranges from early monzogabbro to monzodiorite to later-stage granodiorite (Siron et al., 2016). These intrusions typically display medium-grained equigranular textures through to porphyritic phases with crowded textures dominated by feldspar phenocrysts. Most unaltered late Oligocene intrusions are high-K calcalkaline, and the intrusions occur along a north-northeast trend defined by the alignment of igneous centers and orientation of dikes. A suite of early Miocene intrusions, including Skouries (~20 Ma; Hahn et al. 2012), have porphyritic textures and are quartz monzonite to syenite in composition. Phenocrysts are prismatic consisting of plagioclase and megacrystic K-feldspar, fine-grained euhedral biotite, and relict amphibole. Rounded quartz phenocrysts occur in minor abundance and K-feldspar and quartz comprise the groundmass with accessory zircon, magnetite, and pyrite. The quartz monzonite and syenite intrusions belong to the high-K calc-alkaline to weakly shoshonitic magma series, and their emplacement was controlled by pre-existing structures such as fold axial surfaces and faults.
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SECTION ● 8
Deposit Types
8.1
Deposit Geology
The Olympias deposit is located 6 km north of the Stratoni fault within the Kerdilion unit (Figure 8-1). Replacement-style sulphide orebodies are hosted by marble interlayered within a sequence of quartzo-feldspathic biotite-gneiss, amphibolite and plagioclase microcline orthogneiss (Kalogeropoulos et al., 1989; Siron et al., 2016; Figure 8-2). The massive sulphide orebodies plunge shallowly to the southeast for over 1.8 km (Figure 8-3), subparallel to the orientation of F2 fold hinges and a locally developed L2 intersection lineation (Siron et al., 2018). The locations of the sulphide lenses, however, are largely controlled by strands of the ductile-brittle Kassandra fault and East fault and sub-horizontal shear zones that occur between the two faults (Figure 8-2).
The sulphide ore bodies are grouped into three major domains: East Zone, West Zone, and Flats Zone (Figure 8-3). Two smaller sub-zones, the Remnants and North zones, are considered as part of the West Zone for the purpose of resource estimation. The East Zone ore occurs dominantly in the footwall to the steeply northeast-dipping East fault with the sulphide lenses hosted by marble at or below the upper gneiss-marble contact. The stratabound ore lenses dip shallow to moderately to the northeast and strike for ~ 1.1 km. Ore lenses typically pinch and swell with thickness ranging up to 10 m and widths up to 130 m. The West Zone ore bodies are controlled by the Kassandra fault and have a steep (~ 60 °) NE dip that shallows at depth as it merges into the Flats. The West Zone ore has a strike extent of ~ 1.2 km and is up to 25 m thick with a down dip extent up to 200 m. The Flats ore body extends to the east from the West Zone ore and dips shallowly to the NE, however, in detail it comprises a series of stacked lenses that also extend to the west into the footwall of the Kassandra fault. The Flats ore body has a strike extent of over 1 km and is locally up to 15 m thick, although more commonly ranges from 3 to 10m thick, and ore lenses are tens of metres wide to up to 100 m in places.
Sulphide mineralogy of the Olympias deposit consists of coarse-grained, massive and banded lenses dominated by variable amounts of sphalerite, galena, pyrite, arsenopyrite, chalcopyrite and boulangerite. Accessory minerals include pyrrhotite, marcasite, stibnite, mackinawite, enargite, freibergite, bournonite, geocronite, cubanite, bornite and covellite (Nicolaou and Kokonis, 1980; Kalogeropoulos and Economou, 1987; Kalogeropoulos et al., 1989; Siron et al. 2016). Secondary ion mass spectroscopy studies indicate that Ag and Au primarily occur in solid solution with their respective host minerals; Ag in galena and Au in both arsenopyrite and pyrite.
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Figure 8-1: Geology of the Olympias mine area (modified from Siron et al., 2018)
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Figure 8-2: Simplified Geological Cross Section through the Olympias Deposit (facing NW)
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Figure 8-3: Three-dimensional oblique View of the Major Ore Bodies in the Olympias Mine looking west-northwest
Mine nomenclature classifies the mineralization into eight ore types. Ore types 1 to 3 are base metal‐pyrite dominant, ore types 6 and 7 are arsenopyrite-rich and silica bearing, ore type 8 is manganese-rich and ore types 4 and 5 are sub-economic pyritic wall rock alteration.
Ore types 1, 2 and 3 are gradational and reflect end-members of galena‐sphalerite dominant (Ore Types 1) to pyrite dominant (Ore Type 2) to transitional mixed galena-sphalerite-pyrite (Ore Type 3). Arsenopyrite is common in all three ore types but is not the dominant sulphide. The ore types typically occur as massive to banded sulphide zones with medium to coarse grained sphalerite‐galena‐pyrite‐arsenopyrite and calcite gangue. Ore Types 1 to 3 are dominant in the Flats. Ore Type 7 is arsenopyrite‐rich and has the highest gold content. The mineralization is typically siliceous with massive to banded sulphide dominated by blocky to acicular arsenopyrite with lesser pyrite, galena and sphalerite. Ore Type 7 is locally gradational to Ore Type 3, and banded zones commonly comprise intergrown Ore Types 1 to 3 and 7. Ore Type 7 is dominant in the East Zone. Ore Type 6 is a paragenetically younger quartz-rich sulphide assemblage that locally overprints the early replacement massive sulphide ore horizons. Ore Type 6 can vary from banded siliceous zones to extensive intervals of grey siliceous matrix breccia that contains angular altered wallrock fragments. These quartz-rich sulphide bodies consist of interlocking, euhedral and growth-zoned quartz accompanied by interstitial arsenopyrite and boulangerite with subordinate pyrite, galena and sphalerite. The breccia matrix consists of dark gray chalcedonic quartz containing disseminated, euhedral pyrite, fibrous boulangerite and bladed arsenopyrite. In places Ore Type 6 grades into Ore Type 7, and commonly these quartz-rich ore types are surrounded by lower grade quartz-rhodochrosite alteration of the marble (Ore Type 8). Ore Type 6 and 8 are most commonly developed in the northern and eastern portions of the Olympias deposit, and in and around the Kassandra fault.
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8.2
Deposit Model
Olympias is an example of a polymetallic carbonate replacement deposit, however it is somewhat unusual due to the high Au content of the deposit (cf. Siron et al. 2019). Key characteristics of this class of deposit include carbonate host rocks, massive sulphide mineralization, spatial and temporal relationship with magmatism and zoned metal distribution (Megaw, 1998). Geochronological and paragenetic relationships described by Siron et al. (2016, 2018) confirm that Olympias was coeval with late Oligocene (~ 25-22 Ma) magmatism in the region although the source intrusion has not been identified. Nonetheless, carbon and oxygen isotopic data support a magmatic fluid source (Siron et al. 2019), and the zonation towards Mn-rich alteration in the upper levels of the deposit (North Zone) are consistent with shallow, distal portions of a magmatically derived carbonate replacement system.
Similar carbonate replacement deposits of Oligocene age elsewhere in the West Tethyan belt include the massive sulphide deposits of Maden-Thermes district of southern Bulgaria and northeastern Greece, as well as the deposits in the Lavrion district of southern Greece and skarn and replacement sulphide deposits in the Trepca district of Kosovo (Siron et al., 2019 and references therein). However, for the most part these deposits are gold-poor. Global examples of Au-rich polymetallic carbonate replacement deposits include deposits in the Eureka (Nevada), Bingham (Utah) and Leadville (Colorado) districts of the USA.
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SECTION ● 9
Exploration
Eldorado has not undertaken any significant exploration work at the Olympias.
There have been geological studies carried out on the deposit as follows:
●
Surface geological mapping in the area surrounding the minesite.
●
A study of the deposit geology and structural controls, and a separate petrographic study were carried out in 2013.
●
In 2018 a review of the geological controls of the deposit was carried out. There was also a review of the logging database; this work simplified the logging codes to bring consistency to the process and recommended further studies on the alteration rock geochemistry.
●
Some Terraspec (Hyperspectral) SWIR measurements have been carried out on core to measure alteration footprints to assist in future exploration targeting.
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SECTION ● 10
Drilling
10.1
Introduction
Diamond drillholes are the sole source of subsurface geologic and grade data for the Olympias deposit. There is a large amount of historical drilling in the database. Data generated prior to the drilling by TVX (prior to 1996) was not utilized for Mineral Resource estimation purposes, as there is no means of validating that data.
The previous operator, TVX, drilled 764 drillholes for a total of 93,246 m.
A list of drillholes completed up to 30 June 2019 is shown in Table 10-1.
Table 10-1: Drilling by Year

Year | No. of Holes | Total Metres | Average Depth | Comments
Pre 2000 | 764 | 93,246 | 122 | Pre-Eldorado drillholes -TVX only
2014 | 10 | 1,929 | 193 | Eldorado drilling 2014-2019
2015 | 68 | 8,377 | 123 |
2016 | 228 | 21,778 | 96 |
2017 | 311 | 32,583 | 105 |
2018 | 295 | 33,073 | 112 |
2019 | 179 | 19,907 | 111 | Up to 30 June
Total | 1,855 | 210,893 | 114 |

The core drilled by TVX was located on site in secure containers when Eldorado acquired the property. In 2012-13, Eldorado relogged over 70,000 metres, representing 75% of the TVX drill core, to gain an understanding of the deposit geology and to validate the geological model. The relogging effort facilitated the present ore type classification referred to in Section 8. Since 2014, delineation drilling has been concentrated on the East and West (Kassandra) Zones, in and around areas of mining. This drilling has been carried out predominantly on a 15 x 15 metre grid spacing; local areas have been drilled on a grid spacing of 10 x 10 metres where more detail was necessary. Areas covered by this delineation drilling provide the data used in updating the resource model, superseding the older TVX drill information. Use of TVX data remains in areas that are yet, undrilled by Eldorado.
Since 2014, drilling has been carried out by Greek and Canadian contractors drilling HQ or NQ-size core (63.5 mm or 47.6 mm nominal core diameter). The current contractors are Geotest-GAIA JV from Greece, who drill NQ core, and Cabo Drilling from Canada, who drill both HQ and NQ core. The average hole depth since 2014 has been around 100 m. Fans of drillholes are designed to maximize good intersection angles to the structure and mineralized zones. The mine geology team supervises the rigs and has weekly meetings with the contractors.
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10.2
Core Handling and Sampling Procedures
Drillers place the core into sturdy, locally-made, wooden core boxes. The driller keeps track of the drilling depth and places wooden marker blocks, later nailed in place, at the end of each run to indicate the depth from the collar. Core boxes are clearly labelled. The underground core is delivered to the core shed west of the offices on the mine site, and is logged in a secure area.
The core is logged in detail on tablet computers using the logging software LogchiefTM from Maxgeo. Data are then uploaded into an acQuire database. Data collected includes lithology, alteration, mineralization (including ore types), RQD, core recovery, and other geotechnical factors for input into the Q rating system.
Core photos are routinely taken both on wet and dry core, using a camera stand to ensure consistent photographs.
10.3
Surveys
Mine surveyors set out the location of drill collars prior to the setup of the drill. A survey is taken at the actual collar once the drill is set up. Downhole surveys are also routinely taken using multishot instruments, either a Reflex GyroTM or a Devico Deviflex. Calibration is done annually.
10.4
Core Recovery
Core recoveries are recorded in the geotechnical logs for all drillholes. The overall average recovery is 88.6%, which is considered to be reasonable.
10.5
Bulk Density Determinations
In 2013, Eldorado measured bulk density on over 900 samples of mineralization from historical drillcore using the Archimedes water immersion method (waxed samples) and incorporated the results into the Eldorado database. For a subset of 617 samples, for which there was a complete assay dataset that included Fe, a calculated density was derived as follows:
●
All Zn from assay was assigned to sphalerite; the composition of Olympias sphalerite is assumed to be 86% Zn and 14% Fe based on an historical mineralogical study.
●
All As from assay was assigned to arsenopyrite (along with a portion of the total Fe).
●
All Pb from assay was assigned to galena. While boulangerite is known to be present in Olympias ores, it is not volumetrically significant enough to affect the densities to a great degree and is not included in the density calculation.
●
All remaining Fe (after assignments of Fe to sphalerite and arsenopyrite) was assigned to pyrite.
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●
Mineral percentages for sphalerite, galena, arsenopyrite and pyrite were derived from the assay data and were added together to yield the total sulfide mineral content. Gangue mineral content was calculated by subtracting the total sulfides from 100%.
●
The calculated bulk density was then derived by accumulating the proportion of each of the minerals multiplied by its mineral density (gangue was assigned a density of 2.69 based on measured data from unmineralized samples).
A comparison of the measured versus calculated bulk densities was carried out using a quantile-quantile plot. It was determined that most calculated densities were about 6% higher than measured and that this disparity was most likely due to voids that could not be accounted for in a calculated bulk density. Thus, each block in the resource model has a calculated bulk density (based on the interpolated metal grades for each block) that is then downwardly adjusted by 6% to reflect the typical volume of voids.
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SECTION ● 11
Sample Preparation, Analyses and Security
11.1
Introduction
A number of the samples used in the current resource estimates originated from the 1996 – 98 drill campaign carried out by TVX. The procedures for sampling, analysis and security are described in previous Technical Report on Olympias (2011). Eldorado has reviewed the report and data and agreed with that report’s conclusions that those drill data are acceptable to be used for mineral resource estimation. The remainder of this section describes the sample preparation and quality assurance / quality control (QA/QC) results of the work carried out by Eldorado, in particular during 2018 and 2019.
11.2
Sample Preparation and Assaying
Drill core samples are routinely sent to the ALS facility in Romania. They are bagged and packed in large sealed wooden bins before being trucked to ALS. The sample rejects are returned to the mine site in the same bins. The samples are prepared for assaying at the ALS facility.
Upon delivery to the laboratory the samples were crushed to 90% passing minus 3 mm and prepared according to the following protocol.
●
A 1 kg subsample was riffle split from the crushed minus 3 mm sample and pulverized to 90% passing 75 µm (200 mesh).
●
A 200 g subsample was split off by taking multiple scoops from the pulverized 75 µm sample.
●
The 200 g subsample was placed in a kraft envelope, sealed with a folded wire or glued top, and prepared for shipping. The rest of the pulverized sample was then stored in plastic bags.
●
All equipment was flushed with barren material and blasted with compressed air between each sampling procedure. Regular screen tests were done on the crushed and pulverized material to ensure that sample preparation specifications were being met.
All samples were assayed for gold by 30 g fire assay with an AAS finish, with Au values above 10 ppm determined by a gravimetric finish. Multi-element determination was carried out by Inductively Coupled Plasma Mass Spectrometry (ICP-MS) analysis and / or Inductively Coupled Plasma Emission Spectroscopy (ICP-ES) analysis. (Table 11-1).
11.3
QA/QC Program
11.3.1
Overview
Assay results were provided to Eldorado in electronic format and as paper certificates. The QA/QC procedure included inserting either a CRM, blank and duplicate into the samples stream every 10th samples. Upon receipt of assay results, values for Certified Reference Materials (CRMs) and field blanks were tabulated and compared to the established CRM pass-fail criteria:
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Automatic batch failure if the CRM result is greater than the round-robin limit of three standard deviations.
●
Automatic batch failure if two consecutive CRM results are greater than two standard deviations on the same side of the mean.
●
Automatic batch failure for each element if the field blank or analytical blanks are over the values given in Table 11-2.
Table 11-1: Summary of Assay Methods

ALS
Code | Description
Au-AA25 | Ore grade Au 30 g FA AA finish
Au-GRA21 | Ore grade Au 30 g FA AA finish
ME-ICPORE | Oxidizing digestion w/ ICP-AES finish

Table 11-2: Blank Failure Criteria for Au, Ag, Pb and Zn

Element | Field Blank results | Analytical Blank Results
Au (g/t) | 0.2 > | 0.01>
Ag (g/t) | 20 > | 1 >
Pb % | 0.2 > | 0.01 >
Zn % | 0.2 > | 0.01 >

If a batch failed, it was re-assayed until it passed. Override allowances were made for barren batches. Batch pass/failure data were tabulated on an ongoing basis, and charts of individual reference material values with respect to round-robin tolerance limits were maintained.
The number of samples, duplicates, CRMs and blanks are presented in Table 11-3 along with the compliance figures. The QA/QC compliance of above 5% in each case aligns with accepted norms. A total of 147 batches were sent to the lab during 2018 – 2019.
Table 11-3: Summary of QA/QC sampling in 2018 – 2019

| Original Samples | Blanks | Gold CRMs | Base metal CRMs | Total CRMs | Pulp dups | Coarse dups | Total dups
Number | 8,847 | 653 | 299 | 332 | 631 | 557 | 512 | 1069
Compliance |  | 7.40% | 3.40% | 3.80% | 7.10% | 6.30% | 5.80% | 12.10%

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11.3.2
CRM Results
CRMs used represent all economic minerals at Olympias. For each economic mineral, there should be three corresponding standards:
●
At around the expected cut-off grade of the deposit
●
At the expected average grade of the deposit
●
At a higher grade
In Eldorado’s opinion, both the grade range of the CRMs used, and the percentage of CRM samples submitted are appropriate.
Table 11-4 summarizes the details of the CRMs that were inserted by Eldorado in 2018 and 2019.
Table 11-4: Summary of CRM expected values and standard deviations

Standard ID | Au | Ag | Pb | Zn | No. of Assays
Expected Value (g/t) | SD | Expected Value (g/t) | SD | Expected Value (%) | SD | Expected Value (%) | SD
G310-8 | 7.970 | 0.290 |  |  |  |  |  |  | 42
G903-9 | 11.260 | 0.447 |  |  |  |  |  |  | 39
G908-7 | 4.820 | 0.182 |  |  |  |  |  |  | 23
G909-6 | 0.560 | 0.014 |  |  |  |  |  |  | 30
G912-4 | 1.907 | 0.081 |  |  |  |  |  |  | 54
G915-4 | 9.160 | 0.350 |  |  |  |  |  |  | 26
G317-2 | 12.970 | 0.410 |  |  |  |  |  |  | 26
G915-6 | 0.674 | 0.042 |  |  |  |  |  |  | 28
G318-10 | 4.580 | 0.170 |  |  |  |  |  |  | 31
GBM310-14 |  |  | 59.6 | 2.50 | 8.9465 | 0.3355 | 17.9106 | 0.8084 | 84
GBM911-11 |  |  | 10.2 | 1.10 | 0.1722 | 0.0084 | 0.1422 | 0.0067 | 75
GBM913-11 |  |  | 32.5 | 2.30 | 3.5846 | 0.1541 | 7.8969 | 0.2849 | 74
OREAS 133a |  |  | 99.9 | 2.42 | 4.9000 | 0.1620 | 10.8700 | 0.3540 | 41
OREAS 134a |  |  | 201.0 | 7.00 | 12.7900 | 0.7660 | 17.2700 | 0.5530 | 30
Oreas 137 |  |  | 25.9 | 1.18 | 0.6730 | 0.0200 | 4.9200 | 0.0800 | 28

Figure 11-1 shows the results for a selection of the gold CRMs. They show good performance with variation about the CRM mean values within acceptable ranges. Note in each CRM figure below, that the red line indicates three times the standard deviation, the yellow line indicates two times the standard deviation, the black line represents the expected value and the blue dots are the results of the assayed samples which were sorted by assay date, and blue dashed line shows either 2 or 4 moving average of consecutive samples.
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Figure 11-2 to Figure 11-4 shows a selection of plots showing results for the Ag, Pb, and Zn CRMs.
Figure 11-1: Gold CRM Charts, 2018-2019 Olympias Mine
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Figure 11-2: Ag, Pb, Zn GBM310-14 Charts, 2018-2019 Olympias Mine
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Figure 11-3: Ag, Pb, Zn GBM913-11 Charts, 2018-2019 Olympias Mine
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Figure 11-4: Ag, Pb, Zn OREAS134A Chart, 2018-2019 Olympias Mine
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11.3.3
Blank Samples
Blank samples monitor for contamination during sample preparation and assay process. A total of 653 blank samples were submitted for Au, Ag, Pb, and Zn in 2018 - 2019. Figure 11-5 show the results of the blank samples for Au, Ag, Pb, and Zn, respectively No warnings or failures were identified in the blank samples.
Figure 11-5: Au, Ag, Pb, Zn Blank Performance by Metal
11.3.4
Duplicate Samples
Eldorado implemented a program, which monitored data from regularly submitted coarse reject duplicates and pulp duplicates. Coarse, uncrushed duplicate samples monitor sampling variance, including that arising from crushing, analytical variance and geological variance. The pulp duplicates monitor the assaying procedure. In 2018 – 2019, 512 coarse duplicates and 557 pulp duplicates were submitted for analysis. The duplicate data for Olympias reproduce well and are shown in percentile rank or absolute difference charts for Au and Zn in Figure 11-6 to Figure 11-9. The coarse reject charts show that all data fall well within the 20% limits at the 90th percentile of the population (Eldorado’s threshold for coarse reject duplicates). The pulp duplicate data shows similar compliance to its threshold of 10%.
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Figure 11-6: RPD plot for coarse duplicates Au
Figure 11-7: RPD plot for coarse duplicates Zn
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Figure 11-8: RPD plot for pulp duplicates Au
Figure 11-9: RPD plot for pulp duplicates Zn
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11.4
Concluding Statement
In Eldorado’s opinion, the sampling, sample preparation, security, and analytical procedures, as demonstrated by the QA/QC results, show that the Olympias mine’s assay database, in particular for data since 2017, is sufficiently accurate and precise for resource estimation.
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SECTION ● 12
Data Verification
12.1
Introduction
Data verification is regularly carried out by Eldorado, and comprises the following:
●
Review of available data
●
Review of QA/QC protocols and QA/QC performance
●
Regular discussions and interviews with Olympias mine personnel regarding data collection and database compilation
●
Regular site visits to the minesite since
●
Cross checking of assays in the Eldorado sample database with original laboratory certificates
●
Reconciliation of production to the block model estimates
Eldorado considers that these regular reviews adequately assess the Olympias database.
12.2
Database checks
Eldorado made checks of a selection of the original laboratory assay certificates against the database used for estimation. Eldorado has been operating at the site from 2014 but did some re assay of the historic drill core in 2013 as well. There is a total of 25,325 records in the database that were sent for analysis by Eldorado Gold as of end of 2019. 3,009 of these were standards and blanks and 1,716 duplicate records, these constitutes 11.9% and 6.80% of the total database respectively. As a result of these checks, data supporting the Olympias resource work are sufficiently free of error and adequate for resource estimation.
12.3
Model to Mill Reconciliation
An important measure of performance at any producing mine is reconciliation of the block model to the final mill production figures, adjusted for stockpiles as necessary.
Reconciliation studies are carried out on site on a regular basis. Comparison is made between the depleted block model and the mill month-end figures. There is an adjustment for stockpiles that is applied to the mill figure, and dilution from multiple sources is taken into account when calculating the mine delivered tonnage.
For Olympias, the dilution is calculated from the difference between the mineralized wireframe and the as-built survey of the mined stope. The paste fill and waste contributions to the ore stream are then determined.
The first period, April 2017 to June 2018, tracked initial production. These results are shown in Table 12-1.The second, January to December 2019, represents the first year of regular quarterly reconciliation studies. Table 12-2 shows the summary of those results.
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Table 12-1: Reconciliation for April 2017 to June 2018

| Tonnes | Au g/t | Ag g/t | Pb% | Zn% | Au oz | Ag oz | Pb tonnes | Zn tonnes
Mined | 313,000 | 7.76 | 78 | 2.4 | 3.5 | 78,100 | 784,900 | 7,500 | 11,000
Mill | 312,000 | 7.83 | 83 | 2.7 | 3.7 | 78,500 | 832,600 | 8,400 | 11,500
Mill / Mined | 100% | 101% | 106% | 113% | 106% | 101% | 106% | 112% | 105%

Table 12-2: Reconciliation for January to December 2019

| Tonnes | Au g/t | Ag g/t | Pb% | Zn % | Au oz | Ag oz | Pb tonnes | Zn tonnes
Mined | 285,000 | 7.78 | 89 | 2.8 | 4.1 | 71,300 | 813,600 | 7,900 | 11,700
Mill | 288,000 | 7.30 | 92 | 2.8 | 3.7 | 67,700 | 854,600 | 8,200 | 11,100
Mill / Mined | 101% | 94% | 104% | 102% | 94% | 95% | 105% | 103% | 95%

The reconciliation is detailed and thorough. It is currently providing a quarterly snapshot and demonstrating that the block model, and thus the mineral resources, are valid and robust. This validates the data underpinning the model and is, by association, a good verification of the work done.
12.4
Concluding Statement
The data collection by Eldorado is done in a diligent fashion. The data transfer to the database is by a mix of electronic and manual means, with many internal checks. The production reconciliation that is based on the data is very good. In Eldorado’s opinion, the data are adequate and verified for the purposes of the technical report.
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SECTION ● 13
Mineral Processing and Metallurgical Testing
13.1
Historical Metallurgical Testing
A number of extensive metallurgical reports were commissioned by TVX Hellas SA. The testwork confirmed that the ore responds well to flotation with high recoveries. Other metallurgical testwork and mineralogical investigations have been carried out on samples from the Olympias deposit. These are summarized in the following sections.
13.2
SGS Metallurgical Test Report No. 14607-001 – Final, 2015
In 2015, SGS Lakefield conducted a metallurgical test program on two samples originating from the Olympias Mine. Details and results of this test program were issued in a report titled “An Investigation into Flotation Flowsheet Development of Two Samples from the Olympias Mine for HELLAS GOLD S.A. Project 14607-001 – Final Report, September 23, 2015” (the 2015 Report).
The two samples for this test program were originally denoted as Fresh Ore and Old Ore from the Upper West Zone of the Olympias Mine. The objectives of the program were to define the ore characteristics and confirm the process flowsheet for restarting the mining operation. The scope of the testwork program included head characterization, grindability, flotation, solid-liquid separation, and tailings characterization. A key focus of the test program was to evaluate smelter penalty elements, arsenic and antimony, in the lead and zinc concentrates.
13.2.1
Sample Characterization
The head assays of Old Ore and Fresh Ore samples are summarized in Table 13-1.
The samples contained a significant amount of lead, zinc and sulphur. Of particular interest is the grade of arsenic, which was 10.5% for the Fresh Ore and 5.90% for the Old Ore. The gold content is related to the content of arsenic and, based on the ratio of Au/As, the Fresh Ore was classified from the Upper West Zone while the Old Ore was classified from the Upper East Zone.
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Table 13-1: Old Ore and Fresh Ore Head Assays

Sample ID | Pb % | Zn % | Cu % | Fe % | Cd % | Sb % | Au g/t | Ag g/t | As % | 1S= % | 2S %
Old Ore | 10.2 | 8.43 | 0.39 | 24.5 | 0.047 | 0.4 | 12.4 | 277 | 5.9 | 28.5 | 29.8
Fresh Ore | 5.04 | 4.49 | 0.11 | 27.7 | 0.02 | 0.1 | 22.8 | 159 | 10.5 | 28.2 | 29.2
Sample ID | Al g/t | Ba g/t | Be g/t | Bi g/t | Ca g/t | Co g/t | Cr g/t | K g/t | Mg g/t | Mn g/t | Mo g/t
Old Ore | 1780 | 24.5 | 0.06 | <20 | 18600 | <4 | 169 | 823 | 1110 | 1410 | <5
Fresh Ore | 1630 | 5.8 | 0.04 | <20 | 4950 | <4 | 201 | 766 | 360 | 689 | <5
Sample ID | Na g/t | Ni g/t | P g/t | Se g/t | Sn g/t | Sr g/t | Ti g/t | Tl g/t | U g/t | V g/t | Y g/t
Old Ore | 31 | <20 | <70 | <30 | 108 | 10.9 | 46.2 | <30 | <50 | <40 | 0.3
Fresh Ore | 18 | <20 | <70 | <30 | 163 | 3.94 | 51.9 | <30 | <50 | <40 | <0.2

Notes:
1 Sulphide Sulphur
2 Total Sulphur
13.2.1.1
Mineralogical Examination
Each of the Fresh Ore and Old Ore samples, after stage-pulverizing to 80% passing 120 µm, was screened into five size fractions (+106 µm, -106/+75 µm, -75/+45 µm, -45/+10 µm, and -10 µm) and submitted for mineralogical examination using Quantitative Evaluation of Minerals by Scanning Electron Microscopy (QEMSCAN).
Mineralogical modal analyses are summarized in Table 13-2 and Table 13-3 respectively. As evident from the head assays, each sample is dominated by sulphide minerals.
For the Fresh Ore sample, the sample is dominated by pyrite (37.3%), arsenopyrite (23.1%), and quartz (23.3%) with lesser amounts of galena (6.29%) and sphalerite (6.94%).
For the Old Ore sample, the sample contains pyrite (36.3%), arsenopyrite (12.5%), and quartz (16.9%), but contains higher quantities of galena (12.8%) and sphalerite (13.9%).
Calcite is present in fairly low quantities in the Fresh Ore (1.17%) and Old Ore (4.32%) samples, and all remaining minerals are present in contents less than 1%.
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Table 13-2: Mineral Modals – Fresh Ore

Sample | Fresh Ore
Fraction | Combined | +106 um | -106/+75 um | -75/+45 um | -45/+10 um | -10 um
Mass Size Distribution | 100 | 27.5 | 18.1 | 21.2 | 21.8 | 11.4
Calc'd grain size | 24 | 100 | 65 | 39 | 17 | 7
Mineral Mass (%) | Net | Net | Fraction | Net | Fraction | Net | Fraction | Net | Fraction | Net | Fraction
Pyrite/Marcasite | 37.3 | 13.9 | 50.6 | 6.29 | 34.7 | 7.01 | 33.0 | 6.62 | 30.4 | 3.48 | 30.6
Arsenopyrite | 23.1 | 5.47 | 19.9 | 4.97 | 27.4 | 5.61 | 26.4 | 4.75 | 21.8 | 2.32 | 20.4
Sphalerite | 6.94 | 1.06 | 3.87 | 1.28 | 7.05 | 1.65 | 7.78 | 1.94 | 8.91 | 1.00 | 8.82
Galena | 6.29 | 0.76 | 2.75 | 0.77 | 4.26 | 1.35 | 6.34 | 2.16 | 9.92 | 1.25 | 11.0
Chalcopyrite | 0.32 | 0.06 | 0.22 | 0.04 | 0.21 | 0.05 | 0.24 | 0.07 | 0.34 | 0.09 | 0.82
Other-Cu-Sulphides | 0.02 | 0.00 | 0.01 | 0.00 | 0.01 | 0.00 | 0.01 | 0.00 | 0.02 | 0.00 | 0.04
Boulangerite | 0.11 | 0.02 | 0.07 | 0.02 | 0.09 | 0.02 | 0.11 | 0.02 | 0.09 | 0.03 | 0.30
Fe-Oxides | 0.09 | 0.01 | 0.02 | 0.01 | 0.04 | 0.01 | 0.04 | 0.02 | 0.10 | 0.04 | 0.38
Quartz | 23.3 | 5.89 | 21.4 | 4.44 | 24.5 | 5.12 | 24.1 | 5.39 | 24.8 | 2.50 | 22.0
Plagioclase/K-Feldspar | 0.16 | 0.03 | 0.12 | 0.02 | 0.12 | 0.03 | 0.16 | 0.04 | 0.21 | 0.02 | 0.21
Micas/Clays | 0.68 | 0.10 | 0.37 | 0.06 | 0.35 | 0.11 | 0.52 | 0.20 | 0.94 | 0.20 | 1.78
Other Silicates | 0.06 | 0.01 | 0.05 | 0.01 | 0.06 | 0.01 | 0.05 | 0.02 | 0.08 | 0.01 | 0.09
Calcite | 1.17 | 0.13 | 0.47 | 0.18 | 0.98 | 0.21 | 1.01 | 0.35 | 1.59 | 0.30 | 2.64
Dolomite | 0.30 | 0.02 | 0.06 | 0.03 | 0.19 | 0.04 | 0.17 | 0.13 | 0.58 | 0.09 | 0.75

Table 13-3: Mineral Modals – Old Ore

Sample | Old Ore
Fraction | Combined | +106 um | -106/+75 um | -75/+45 um | -45/+10 um | -10 um
Mass Size Distribution | 100 | 28.8 | 15.8 | 19.6 | 22.1 | 13.8
Calc'd grain size | 21 | 98 | 57 | 36 | 17 | 6
Mineral Mass (%) | Net | Net | Fraction | Net | Fraction | Net | Fraction | Net | Fraction | Net | Fraction
Pyrite/Marcasite | 36.3 | 14.7 | 51.2 | 5.93 | 37.6 | 6.50 | 33.2 | 6.23 | 28.1 | 2.87 | 20.9
Arsenopyrite | 12.5 | 3.50 | 12.2 | 1.92 | 12.1 | 2.63 | 13.4 | 2.69 | 12.1 | 1.79 | 13.0
Sphalerite | 13.9 | 2.54 | 8.83 | 2.43 | 15.4 | 3.29 | 16.8 | 3.69 | 16.6 | 1.97 | 14.3
Galena | 12.8 | 1.53 | 5.32 | 1.75 | 11.1 | 2.58 | 13.2 | 4.06 | 18.3 | 2.86 | 20.8
Chalcopyrite | 0.53 | 0.07 | 0.26 | 0.06 | 0.38 | 0.07 | 0.34 | 0.12 | 0.55 | 0.21 | 1.51
Tetrahedrite/tennantite | 0.13 | 0.00 | 0.02 | 0.02 | 0.14 | 0.04 | 0.19 | 0.05 | 0.22 | 0.01 | 0.11
Other-Cu-Sulphides | 0.12 | 0.00 | 0.01 | 0.01 | 0.07 | 0.02 | 0.11 | 0.04 | 0.20 | 0.04 | 0.31
Boulangerite | 0.34 | 0.05 | 0.19 | 0.06 | 0.36 | 0.06 | 0.31 | 0.09 | 0.41 | 0.08 | 0.58
Fe-Oxides | 0.19 | 0.07 | 0.23 | 0.01 | 0.06 | 0.01 | 0.07 | 0.02 | 0.09 | 0.08 | 0.55
Quartz | 16.9 | 5.38 | 18.7 | 2.79 | 17.7 | 3.22 | 16.5 | 3.41 | 15.4 | 2.08 | 15.1
Plagioclase/K-Feldspar | 0.23 | 0.04 | 0.15 | 0.03 | 0.19 | 0.04 | 0.19 | 0.04 | 0.19 | 0.08 | 0.55
Micas/Clays | 0.72 | 0.10 | 0.33 | 0.07 | 0.44 | 0.11 | 0.56 | 0.21 | 0.95 | 0.24 | 1.71
Other Silicates | 0.09 | 0.02 | 0.08 | 0.01 | 0.06 | 0.01 | 0.05 | 0.02 | 0.07 | 0.03 | 0.22
Calcite | 4.32 | 0.59 | 2.06 | 0.61 | 3.87 | 0.86 | 4.38 | 1.25 | 5.66 | 1.01 | 7.31
Dolomite | 0.81 | 0.11 | 0.37 | 0.07 | 0.47 | 0.11 | 0.56 | 0.19 | 0.87 | 0.33 | 2.42

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13.2.1.2
Mineral Liberation and Association
The main minerals that are recoverable by flotation are galena, sphalerite, pyrite, and arsenopyrite, to produce lead, zinc, and gold sulphide concentrates. The liberations of these minerals are summarized in the following sections.
Galena Liberation
The liberation of galena in the Fresh and Old Ore samples at grind size of 80% passing 120 µm is summarized in Figure 13-1and Figure 13-2, respectively.
Figure 13-1: Galena Liberation – Fresh Ore
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Figure 13-2: Galena Liberation – Old Ore
As the data indicates, both samples are well liberated, with the amount of free and liberated grains being 90.4% and 93.7% for the Fresh and Old Ore samples, respectively. In both cases, the proportion of locked and sub-middling particles reduces considerably in the finer fractions, and this suggests that regrinding would be required to attain high concentrate grades.
For the Fresh Ore sample, galena is mainly associated with pyrite (4.48%) and arsenopyrite (3.34%), with smaller amounts associated with sphalerite (0.90%) and complex particles (0.49%). Galena contained in the Old Ore sample is mainly associated with pyrite (3.00%), but significantly less so with arsenopyrite (0.60%), sphalerite (1.20%), and complex particles (0.62%).
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Sphalerite Liberation
The liberation of sphalerite at grind size of 80% passing 120 µm is summarized in Figure 13-3 and Figure 13-4 for the Fresh and Old Ore samples, respectively.
Figure 13-3: Sphalerite Liberation – Fresh Ore
The charts demonstrate that the degree of sphalerite liberation is higher than galena, with the content of free and liberated grains being 95.8% in the Fresh Ore and 96.2% in the Old Ore sample, respectively. The amount of locked and middling particles in the coarser size fractions indicates that regrinding would be required to attain high concentrate grades.
For the Fresh Ore and Old Ore samples, the amount of sphalerite associated with pyrite is 1.51% and 1.60%, respectively, with the association with other minerals being less than 1%.
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Figure 13-4: Sphalerite Liberation – Old Ore
Pyrite Liberation
The liberation of pyrite is summarized in Figure 13-5 and Figure 13-6 for the Fresh and Old Ore samples, respectively. The liberation of pyrite is very good, with the proportions being 97% and 96% for the Fresh and Old Ore samples, respectively. As the liberation is consistent across the size fractions, regrind would not be required.
Arsenopyrite Liberation
The liberation of arsenopyrite is 96% and 92% in the Fresh and Old Ore samples, respectively. The coarsest size fraction (+106 µm) has the lowest degree of liberation, but at 91% and 86% for the Fresh and Old Ore samples, the liberation is quite good.
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Figure 13-5: Pyrite Liberation – Fresh Ore
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Figure 13-6: Pyrite Liberation – Old Ore
Electron Microprobe Analysis
Electron microprobe analysis was conducted to determine element deportment in various minerals and mineral identification. The mineral assay determined by microprobe analysis is summarized in Table 13-4. An important objective of this work was to identify minerals that contain arsenic and antimony that can report to the lead and zinc concentrates.
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Table 13-4: Electron Microprobe Mineral Assay Summary

Mineral | Sample | %
S | As | Zn | Ag | Fe | Pb | Sb | Cu | Total
Pyrite | Fresh Ore | 52.1 | 1.40 | 0.026 | 0.0043 | 46.2 | 0.0 | 0.0045 | 0.019 | 99.8
Old Ore | 52.3 | 1.22 | 0.047 | 0.0017 | 46.0 | 0.0 | 0.0020 | 0.010 | 99.5
Global Average | 52.2 | 1.31 | 0.036 | 0.0030 | 46.1 | 0.0 | 0.0032 | 0.014 | 99.7
Galena | Fresh Ore | 13.4 | 0.010 | 0.046 | 0.18 | 0.072 | 85.9 | 0.27 | 0.0004 | 99.8
Old Ore | 13.3 | 0.019 | 0.057 | 0.13 | 0.067 | 85.4 | 0.27 | 0.0100 | 99.3
Global Average | 13.4 | 0.015 | 0.051 | 0.15 | 0.070 | 85.6 | 0.27 | 0.0052 | 99.6
Arsenopyrite | Fresh Ore | 22.4 | 41.6 | 0.021 | 0.006 | 35.7 | 0.0 | 0.042 | 0.0009 | 99.7
Old Ore | 22.4 | 41.6 | 0.110 | 0.004 | 35.7 | 0.0 | 0.069 | 0.0026 | 99.8
Global Average | 22.4 | 41.6 | 0.064 | 0.005 | 35.7 | 0.0 | 0.055 | 0.0018 | 99.8
Sphalerite | Fresh Ore | 33.5 | 0.0042 | 59.0 | 0.0098 | 6.97 | 0.0 | 0.0003 | 0.220 | 99.7
Old Ore | 33.6 | 0.0000 | 57.9 | 0.0060 | 8.43 | 0.0 | 0.0026 | 0.052 | 100.0
Global Average | 33.6 | 0.0021 | 58.5 | 0.0078 | 7.72 | 0.0 | 0.0015 | 0.130 | 99.9
Boulangerite / Bournonite | Fresh Ore | 18.2 | 0.24 | 0.42 | 0.0130 | 0.91 | 56.0 | 24.1 | 0.013 | 99.9
Old Ore | 18.9 | 0.44 | 0.13 | 0.0000 | 0.15 | 50.9 | 24.6 | 4.27 | 99.3
Global Average | 18.6 | 0.35 | 0.26 | 0.0061 | 0.50 | 53.2 | 24.4 | 2.34 | 99.6
Tetrahedrite | Old Ore | 25.1 | 1.01 | 3.66 | 1.16 | 3.76 | 0.0 | 27.4 | 37.2 | 99.3

The only mineral present that contains appreciable amounts of arsenic is arsenopyrite. Tetrahedrite and pyrite contain a relatively high amount of arsenic at approximately 1%, but due to the likely small amount of these minerals being present in the concentrate, they are not expected to be problematic. For antimony, galena contains 0.27%. Other minerals containing antimony at significant levels are boulangerite, bournonite, and tetrahedrite. These antimony minerals would be expected to report to the lead concentrate and could cause excessive antimony contamination.
13.2.2
Grindability testwork
13.2.2.1
Unconfined Compressive Strength Test
One rock from each sample was used for unconfined compressive strength (UCS) testing (Table 13-5). The compressive strength was 83.5 MPa and 35.6 MPa for the Fresh and Old Ore samples, respectively. Based on these data, the Fresh Ore sample was classified as ‘soft’, while the Old Ore sample was classified as ‘very soft’.
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Table 13-5: UCS Test Results

Sample | Length (cm) | Diameter (cm) | Compressive strength (MPa) | Specific Gravity (t/m3)
Fresh Ore | 6.81 | 3.74 | 83.5 | 3.90
Old Ore | 9.20 | 3.76 | 35.6 | 4.63

13.2.2.2
Bond Rod Mill Grindability Tests
The test results for Bond rod mill grindability testing are summarized in Table 13-6.
Table 13-6: Bond Rod Mill Grindability Test Results

Sample name | Mesh of grind | Feed F80 (mm) | Product P80 (mm) | Work index (kWh/t)
Fresh Ore | 14 | 11,294 | 832 | 6.8
Old Ore | 14 | 9,320 | 872 | 6.7

The Fresh Ore sample had a rod mill work index of 6.8 kWh/t and was classified as ‘very soft’, while the Old Ore sample had a similar rod mill work index of 6.7 kWh/t and was also classified as ‘very soft’.
13.2.2.3
Bond Ball Mill Grindability Test
The Bond ball mill grindability test was conducted at two closing mesh grind sizes, 65 mesh and 80 mesh (212 µm and 180 µm), to attain a product size distribution within the grind target between 120 µm to 150 µm. The test results are summarized in Table 13-7.
Table 13-7: Bond Ball Mill Grindability Test Results

Sample name | Mesh of grind | Feed F80 (µm) | Product P80 (µm) | Work index (kWh/t)
Fresh Ore | 65 | 2,357 | 168 | 8.0
80 | 2,357 | 145 | 8.5
Old Ore | 65 | 2,182 | 171 | 8.9
80 | 2,182 | 148 | 9.2

The Fresh Ore sample had similar ball mill work indexes of 8.0 kWh/t and 8.5 kWh/t at the mesh of grinds of 65 and 80, respectively, with classification as ‘very soft’ in both cases. The Old Ore sample also had similar ball mill work indexes of 8.9 kWh/t and 9.2 kWh/t at the mesh grinds of 65 and 80, respectively, and was also classified as ‘very soft’ in both cases.
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13.2.2.4
Bond Abrasion Tests
The Fresh Ore and Old Ore samples were submitted for the Bond abrasion index test. The Fresh Ore sample was considered ‘medium abrasive’ with an abrasion index of 0.274 g, while the Old Ore sample was also considered ‘medium abrasive’ with an abrasion index of 0.244 g.
13.2.3
Flotation Testwork
Both the Fresh Ore and Old Ore samples were subjected to flotation testwork, with the purpose of producing separate lead, zinc, and gold concentrates.
Testwork included an initial flotation tests under conditions from past flotation testwork at primary grind of 80% passing 100 to 120 µm and isopropyl xanthate as a collector in all three circuits. A series of rougher and open-circuit cleaning tests; and a locked cycle test (LCT) were completed with the Fresh Ore sample. A total of 45 batch flotation tests were completed, using Fresh Ore sample (40 tests) and Old Ore sample (5 tests)
13.2.3.1
Fresh Ore Sample
Lead Rougher Flotation
A variety of variables were evaluated including the following: Collector type (Aerophine 3418A or SIPX) and dosage, ZnSO4 dosage, lime dosage, primary grind size, and rougher flotation residence time.
The primary grind size was generally in the order of 80% passing 110 µm ~ 158 µm. The effect of various parameters on lead recovery and selectivity against other metals was relatively small. Lead recovery ranged from 91.8% to 94.1%, with an average of 92.9%.
Following conditions provided the best results:
●
12 minutes of rougher flotation time
●
750 g/t of lime and 50 g/t of NaCN added to the mill
●
pH maintained at 9.5
●
35 g/t of Aerophine 3418A distributed to the first three rougher concentrates
Lead Cleaner Flotation
Twenty-three (23) tests were completed that involved lead cleaner flotation, with the purpose of reducing the amount of arsenic reporting to the final lead concentrate. Rejecting arsenic to the desired level of 0.5% was not achieved during this testwork program and the level of arsenic in the final lead concentrate was seldom less than 3%. The content of antimony in the lead concentrate was not an issue with the Fresh Ore sample and was generally less than 0.5%.
Many varying conditions and reagents were trialed in an attempt to reject arsenic. These included: regrind size, addition of lime, high sodium cyanide dosage, depressants such as guar gum, Aero 7260HFP, hydrogen peroxide, sodium citrate, sodium metabisulphite, extended regrind with multiple stages, and reverse flotation of arsenopyrite from lead concentrate.
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Based on the results from all rougher and open-circuit cleaner tests, it was concluded that entrainment is the main factor for arsenic reporting to the lead concentrate. Flotation of arsenopyrite associated with galena was also partially responsible. It was estimated that, for arsenic grade in the lead concentrate to reach 0.5%, arsenic head grade in the feed should be lower than 5% and the final concentrate lead grade should be above 80%. When the concentrate lead grade is less than 80%, arsenic content in the concentrate will increase accordingly.
Lead Concentrate Characterization
The size-by-size analysis was completed on two lead concentrate samples. One concentrate was high grade from initial lead cleaning test, while the second sample was lower grade from the combined lead concentrate. Size-by-size analysis of the lead concentrate is provided in Table 13-8.
Table 13-8: Lead Concentrate Size-by-Size Analysis

size range | Wt. % | Assay % | Distribution %
Pb | Cu | Zn | Fe | As | S | Pb | Cu | Zn | Fe | As | S
+53 µm | 10.9 | 36.6 | 0.38 | 4.88 | 15.3 | 7.80 | 20.9 | 6.9 | 14.1 | 14.9 | 17.7 | 16.2 | 12.8
38 ~ 53 µm | 18.1 | 49.8 | 0.29 | 4.49 | 12.6 | 7.33 | 19.5 | 15.7 | 17.8 | 22.7 | 24.2 | 25.2 | 19.9
25 ~ 38 µm | 15.2 | 55.7 | 0.25 | 4.21 | 10.9 | 6.52 | 18.4 | 14.7 | 12.9 | 17.9 | 17.6 | 18.9 | 15.8
14 ~ 25 µm | 31.1 | 64.9 | 0.21 | 2.51 | 7.70 | 4.97 | 16.4 | 35.3 | 22.3 | 21.9 | 25.5 | 29.5 | 28.8
‐14 µm | 24.8 | 63.4 | 0.39 | 3.28 | 5.64 | 2.17 | 16.2 | 27.4 | 32.9 | 22.7 | 14.9 | 10.2 | 22.6
Head (Calc) | 100 | 57.3 | 0.29 | 3.57 | 9.39 | 5.25 | 17.7 | 100 | 100 | 100 | 100 | 100 | 100
Head (Direct) | 100 | 56.1 | 0.28 | 3.60 | 9.19 | 5.44 | 18.0 | NA | NA | NA | NA | NA | NA

These data show that lead grade is lowest in the coarser size fractions, while arsenic grade is highest, indicating that liberation may be an issue. Mineralogical examination indicated that the above lead concentrate contained 65.4% galena, 14.1% arsenopyrite, 9.13% pyrite, 6.36% sphalerite, and 2.93% quartz. Arsenic is present as arsenopyrite, with a minor amount associated with pyrite.
The arsenopyrite is fairly well liberated with 81% of arsenopyrite grains being greater than 90% liberated. Rejection of the liberated arsenopyrite should be possible, while further regrinding might be necessary to liberate the arsenopyrite associated with galena.
Zinc Rougher Flotation
Zinc flotation conditions included activation using copper sulphate and flotation using a weaker xanthate collector (SIPX) to limit recovery of pyrite and arsenopyrite, high pH of 11.8. Similar to the lead rougher, a series of variables were investigated, including collector dosage, CuSO4 dosage, slurry pH, and primary grind size.
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Unlike the lead flotation, zinc flotation appeared to be more sensitive to flotation variables. Important variables were SIPX dosage, CuSO4 dosage, and lime dosage (pH).
60 g/t SIPX was the best compromise between zinc recovery and selectivity. Lime dosage appeared to have limited impact on zinc recovery. However, the selectivity of zinc flotation over pyrite/arsenopyrite was improved considerably at the high lime dosage of 1,800 g/t. The higher dosage of 600 g/t of CuSO4.5H2O provided the greatest zinc recovery and selectivity compared to the lower CuSO4.5H2O dosages of 300 and 500 g/t.
Coarser grind sizes of 141 and 158 µm resulted in zinc recovery ranging from 81.9% to 84.5%. The zinc recovery increased to 88.6% at the finer grind size.
Following conditions provided the best results for zinc rougher flotation: Lime dosage of 1,890 g/t (pH 11.8), CuSO4.5H2O dosage of 600 g/t, and SIPX dosage of 60 g/t.
It was found that using soda ash as opposed to lime in the primary grind avoided the formation of an extremely hydrophobic phase following CuSO4 activation. The use of soda ash in the primary grind in two tests demonstrated that the requirement for CuSO4 could be significantly reduced.
Zinc Cleaner Flotation
The best zinc open-circuit cleaner flotation test achieved zinc recovery of 74.2% at concentrate grade of 55.9% zinc. The contents of arsenic and antimony in the zinc concentrate were low at 0.46% and 0.038%, respectively.
Pyrite and Arsenopyrite Flotation
Following the flotation of lead and zinc, the gold is recovered by flotation of remaining sulphide minerals into a bulk sulphide concentrate with concentrate grade of 20 – 25 g/t Au. The main gold bearing minerals are predominately pyrite and arsenopyrite. Flotation is carried out by reducing slurry pH to 6 with sulphuric acid, and with a large dosage of SIPX.
The highest recovery of gold was 90.6%. A single stage cleaning test resulted in very little upgrading for gold grade from 34.0 g/t to 34.8 g/t. The cleaner stage recovery was quite high at 99.7%.
The preferred conditions of pyrite and arsenopyrite flotation were pH 6.0, 250 g/t SIPX, and MIBC as frother.
Locked-Cycle Flotation Test
A locked-cycle test was completed with the Fresh Ore sample. 1 kg test charges were used. Target and actual grind sizes are provided in Table 13-9.
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Table 13-9: LCT-1 Grind Size Summary

Sample | Target P80, µm | Actual P80, µm
Lead Concentrate | 25 | 26
Zinc Concentrate | 25 | 26
Gold Pyrite/Arsenopyrite Concentrate | 109 | 102
Final Tail | 103 | 106

The flowsheet is summarized in Figure 13-7 and the conditions are summarized in Table 13-10.
Figure 13-7: Locked-Cycle Test Flotation Flowsheet
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Table 13-10: LCT-1 Flotation Conditions

Stage | Reagents added, g/t | Time, minute | pH
Lime | NaCN | 3418A | MIBC | CuSO4 | H2SO4 | SIPX | Grind | Cond. | Froth
Lead Rougher | 900 | 50 | 35 | 20 |  |  |  | 7.5 | 4 | 6.5 | 9.5
Lead Regrind | 50 | 20 |  |  |  |  |  | 3 |  |  | 8.2
Lead Cleaner | 100 | 30 | 5 | 35 |  |  |  |  | 4 | 7 | 9.5
Zinc Rougher | 156 |  |  | 20 | 600 |  | 60 |  | 3 | 6.5 | 11.8
Zinc Regrind | 250 |  |  |  | 50 |  |  | 3 |  |  | 10.9
Zinc Cleaner | 980 |  |  | 25 |  |  | 5 |  | 4 | 7.5 | 11.8
Py / AsPy Rougher RRougherRougherRougher |  |  |  | 30 |  | 2990 | 250 |  | 3 | 6.5 | 6.0
Py / AsPy Cleaner |  |  |  | 10 |  | 150 |  |  | 2 | 3.5 | 6.0

The locked-cycle test demonstrated reasonably good stability overall.
The metallurgical performance for Fresh Ore sample from the locked-cycle test is summarized in Table 13-11.
Table 13-11: Locked-Cycle Test Metallurgical Projection – Fresh Ore

Product | Wt | Assays, %, g/t | % Distribution
% | Pb | Zn | Au | Ag | S | As | Sb | Pb | Zn | Au | Ag | S | As | Sb
Pb 3rd Cleaner Conc | 5.7 | 69.5 | 1.63 | 7.76 | 1830 | 14.5 | 3.22 | 0.34 | 83.6 | 2.0 | 1.9 | 75.8 | 2.9 | 1.8 | 30.1
Zn 3rd Cleaner Conc | 7.5 | 6.64 | 53.6 | 1.34 | 207 | 31.2 | 0.49 | 0.046 | 10.5 | 85.9 | 0.4 | 11.3 | 8.3 | 0.4 | 5.4
AsPy / Py Cleaner Conc | 62.1 | 0.38 | 0.72 | 35.8 | 26.3 | 39.6 | 16.1 | 0.064 | 5.0 | 9.5 | 96.1 | 11.9 | 87.4 | 96.1 | 62.0
Final Tail | 24.6 | 0.18 | 0.49 | 1.44 | 6.0 | 1.56 | 0.73 | 0.007 | 0.9 | 2.6 | 1.5 | 1.1 | 1.4 | 1.7 | 2.5
Head (back calculated) | 100 | 4.75 | 4.70 | 23.1 | 138 | 28.2 | 10.4 | 0.064 | 100 | 100 | 100 | 100 | 100 | 100 | 100
Head (direct) | 100 | 5.04 | 4.49 | 22.8 | 159 | 29.2 | 10.1 | 0.10 | NA | NA | NA | NA | NA | NA | NA

High-grade lead and zinc concentrates were produced containing 69.5% Pb and 53.6% Zn, respectively. Concentrate recoveries were 83.6% for lead and 85.9% for zinc in lead and zinc concentrates, respectively.
The lead concentrate contained 3.22% arsenic and 0.34% antimony. In view of 10.4%, arsenic in the feed, 3.22% arsenic in the lead concentrate is quite good. For the plant operation, arsenic content in the mill feed is controlled normally to less than 5%. Because arsenic in the lead concentrate is proportional to arsenic in the mill feed, less than 2% arsenic in the lead concentrate is expected if arsenic content in the mill feed is less than 5%.
The zinc concentrate contained 10.5% Pb, which is high and suspected to be the result of elevated level of lead in the lead cleaner scavenger tail reporting to the zinc circuit. Re-routing this stream to the lead rougher circuit is an option to consider for improvement. Due to the lower recovery of zinc in the zinc circuit, a considerable amount of zinc reported to the gold concentrate. Improving zinc recovery by higher collector dosages during the cleaning stage or re-routing the zinc cleaner scavenger tail to the zinc rougher were noted as options to consider for improvement.
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The pyrite / arsenopyrite concentrate grade was extremely high containing 39.6% S, 35.8 g/t Au, which represents 96.1% gold recovery.
13.2.3.2
Old Ore Sample
The flotation testing with the Old Ore sample was limited to five flotation tests. The Old Ore sample had considerably higher grade than the Fresh Ore sample, containing almost twice the amount of lead and zinc than the Fresh Ore sample. However, the Old Ore sample contained approximately half the amount of arsenic compared to the Fresh Ore sample. All flotation tests responded well with 30 ~ 40 g/t Aerophine 3418A and 75 ~ 80 g/t sodium cyanide addition, except for one test (F7) which resulted in excessive arsenic recovery at increased collector and reduced cyanide additions. There is similarity in the zinc and arsenic selectivity, but the mass pull to achieve the same recovery was significantly higher due to the substantially higher grade of lead in the feed.
For zinc rougher flotation, due to the higher feed grades, a higher mass pull would be required to attain the same level of recovery. pH 11.8 was used for activation and resulted in considerably greater selectivity for zinc over arsenic.
The performance of pyrite/arsenopyrite flotation was relatively poor, due to high recovery of gold-bearing sulphides in the zinc circuit. Nevertheless, the stage-recovery of gold approaches 100% for all tests
Due to the high head grade of lead and low head grade of arsenic compared to the Fresh Ore sample, the content of arsenic in the lead concentrates was quite low at 0.47 ~ 0.70%. However, the content of antimony in the lead concentrate was quite high at 2.32% ~ 2.64%. Antimony floated strongly together with lead. The overall lead recovery for the Old Ore sample was 74.4%, despite higher lead rougher recovery of 94.4%.
It appears that the addition of lime considerably reduced lead recovery. The low arsenic and high antimony grade in the lead concentrate was due to the lower arsenic head grade and higher antimony head grade in the feed.
13.2.3.3
Large Batch Flotation
Large batch flotation was completed on 300 kg of Fresh Ore sample to produce 90 kg of tailings for backfill testing by a third-party laboratory. The large batch flotation test resulted in lower concentrate grades and recoveries than anticipated. This was attributed to the fairly coarse lead regrind P80 of 44 µm compared to the 25 µm used during the smaller-scale laboratory testwork. Solid-liquid separation testwork was conducted on the products from the large batch flotation test, including final tail, gold/pyrite/arsenopyrite cleaner concentrate, zinc cleaner concentrate and lead cleaner concentrate.
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13.2.3.4
Solid-Liquid Separation Testing
Flocculant scoping and static settling testwork indicated that all samples responded well to BASF Magnafloc 10 flocculant. Static settling tests showed underflow (U/F) densities of 66% w/w for the final tail, and 71% ~ 79% w/w for the concentrates. The results are presented in Table 13-12.
Table 13-12: Static Settling Results Summary

Sample ID | Dosage flocc't g/t | Feed1 %w/w | U/F2 %w/w | CSD U/F3 %w/w | TUFUA4 m2/t/day | Supernatant5 visual | TSS mg/L
Final tail | 9 | 10 | 66 | 68 | 0.04 | Clear | <10
Gold/pyrite/arsenopyrite concentrate | 5 | 15 | 75 | n.d | 0.01 | Clear | 39
Zinc concentrate | 9 | 15 | 71 | 74 | 0.03 | Clear | 19
Lead concentrate | 8 | 15 | 79 | 80 | 0.02 | Clear | 18

Notes:
All values were calculated without a safety factor. Common conditions: Raked, ambient temperature. Flocculants: BASF Magnafloc 10
1 Diluted Thickener Feed.
2 Ultimate Underflow Density.
3 Underflow Density as predicted by the Critical Solids Density (CSD) as detailed in the Rheology Test Section.
4 Thickener Underflow Unit Area corrected with actual CSD.
5 Supernatant Clarity after 60 minutes of elapsed settling time.
Dynamic settling tests were conducted on all samples except for the zinc cleaner concentate due to insufficient mass. Table 13-13 summarizes the optimized dynamic settling operating conditions and results, confirming the U/F densities obtained in the static settling tests.
Table 13-13: Optimized Dynamic Settling Results Summary

Sample ID | Dosage flocc't | Feed1 | U/F2 | U/F3 extended | TUFUA4 | Solids loading | Net Hyd. loading | Res. time solid vs UF | O/F | TSS
g/t | % wt. | % wt. | % wt. | m2/t/d | t/m2/h | m3/m2/d | h | visual | mg/L
Final tail | 10 | 10 | 64.5 | 72.4 | 0.10 | 0.416 | 3.70 | 0.84 | Clear | 62
Gold/pyrite/arsenopyrite concentrate | 10 | 15 | 80.6 | 87.0 | 0.07 | 0.592 | 3.30 | 1.1 | Clear | 10
Lead concentrate | 12 | 15 | 79.1 | 83.4 | 0.08 | 0.520 | 2.88 | 1.4 | Clear | 22

Notes:
All values were calculated without a safety factor. Flocculants: BASF Magnafloc 10 underflow raking, ambient temperature.
1 Autodiluted Thickener Feed.
2 Underflow (UF) Density.
3 Underflow (UF) Density after 30 minutes of extended thickening (raked, no feed).
4 Thickener Underflow Unit Area.
Concentric cylinder rotational viscometry (CCRV) rheology was conducted on the final tail, zinc concentrate, and lead concentrate underflow samples, while vane rheology measurements were conducted on the gold/pyrite/arsenopyrite concentrate underflow sample due to its fast settling nature. Table 13-14 summarizes the critical solids densities (CSD) and corresponding yield stress values.
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Table 13-14: Underflow Rheology Results Summary

Sample I.D. | CSD1, %w/w solid | Yield stress, Pa Un - sheared
Final tail | 67.6 | 21
Zinc concentrate | 74.0 | 14
Lead concentrate | 80.0 | 10

Notes:
1 CSD: Rheology-determined Critical Solids Density.
Vane rheology was conducted at pulp densities from 85.8% to 87.7% w/w solids using a Haake VT550 viscometer. The resulting vane-yield stress values ranged from 222 Pa to 989 Pa at 86.1% to 87.7% w/w solids (see Table 13-15). At 85.8% w/w solids, the sample settled and did not provide a yield point.
The critical solid density (CSD) is the solids density at which a small increase of the solid density causes a significant decrease of flowability of the slurry. The CSD value is also predictive of the maximum underflow solids density achievable in a commercial thickener and pumpability achievable in practice.
Table 13-15: Vane Rheological Data Summary – Gold/Pyrite/Arsenopyrite Concentrate

Test No. | Size, P80 (µm) | SG1 (g/cm3) | Temp (0C) | Solids % | Shear stress 2Peak (Pa)
1 | 103 | 5.10 | 20.0 | 87.7 | 989
2 | 103 | 5.10 | 20.0 | 86.9 | 527
3 | 103 | 5.10 | 20.0 | 86.1 | 222
4 | 103 | 5.10 | 20.0 | 85.8 | No yield

Notes:
The gold/pyrite/arsenopyrite concentrate samples were flocculated with Magnafloc 10 at a dosage of 10 g/t.
Shear Stress value was produced by subjecting the slurry sample to a constant rotation at 0.022 s-1 for 4 minutes.
1 SG - Dried solids specific gravity, measured with pycnometer
2 Maximum (peak) initial shear stress achieved.
Results of vacuum filtration tests on the final tail underflow sample are summarized in Table 13-16. Using a Testori 6583 TC filter cloth, a cake moisture of 13.6% w/w was achieved from a 65% w/w pulp density with a dry time of 30 seconds. A qualitative confirmation test at lower filter feed solids (60% w/w) achieved a similar cake moisture of 15% w/w.
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Table 13-16: Vacuum Filtration Results Summary of the Final Tail

Operating conditions | Filter throughput1
Feed solids %w/w | Vacuum level, inch Hg | Form time (s) | Dry time (s) | Form / dry ratio | Cake thickness (mm) | Throughput, dry kg/m2h | Cake2 moisture %w/w
65 | 23.4 | 5 | 5 | 1.00 | 20.0 | 10329 | 21.5
4 | 4 | 1.06 | 14.5 | 9111 | 19.5
4 | 6 | 0.65 | 15.0 | 7016 | 17.4
5 | 15 | 0.32 | 15.0 | 3853 | 16.3
4 | 20 | 0.21 | 15.0 | 3140 | 14.7
3 | 30 | 0.10 | 10.5 | 1594 | 13.6
6 | 2 | 4.00 | 20.0 | 13397 | 22.3
60 | 24 | 6 | 30 | 0.20 | 15.0 | 2123 | 15.0

Notes:
1 Filter throughput predictions versus test conditions using raw test data. Results are not for sizing of any specific type of filter. Refer to individual test results for additional sizing information.
2 Common cake characteristics for all samples: Cohesive, Good cake release from cloth.
Conditions and results of pressure filtration tests on the concentrate underflows are summarized in Table 13-17.
A qualitative confirmation test at lower filter feed solids was conducted using the best conditions of the pressure filtration test. Cake moistures ranged from 3.8% to 5.1% w/w for the gold/pyrite/arsenopyrite concentrate, from 5.2% to 7.2% w/w for the zinc concentrate, and from 5.0% to 6.0% w/w for the lead concentrate.
13.2.3.5
Environmental Characterization
The final tailing sample from the large batch flotation testing was subjected to a basic environmental test program, including elemental analysis (qualitative X-ray diffraction, whole rock, and Aqua regia digest analyses), toxicity characteristic and synthetic precipitation leaching procedure (TCLP and SPLP) extractions, modified acid base accounting (ABA), net acid generation (NAG) testing, and NAG liquor analysis.
Elemental analyses indicate that the tailing sample was predominantly comprised of quartz with minor amounts of calcium, iron, and aluminum. Environmentally significant concentrations of arsenic, lead, and zinc were also observed.
Analysis of the TCLP leachate reported arsenic and lead at concentrations expected to be of environmental concern. It should be noted that since the TCLP is a very aggressive extraction procedure, the limits applicable to this test procedure are much higher than those used for the SPLP or NAG liquor analyses. With the exceptions of pH, arsenic, and lead, the SPLP extraction leachate reported all of the typically controlled parameters well within the limits characteristically applied to this test procedure.
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Table 13-17: Pressure Filtration Results Summary for the Concentrates

Sample identification | Filter cloth | Operating conditions | Filter throughput1
Feed solids %w/w | Pressure level, PSI | Form time (s) | Dry time (s) | Form / dry ratio | Cake thickness (mm) | Throughput, dry kg/m2h | Cake2 moisture, % w/w
Gold/pyrite/ arsenopyrite concentrate | Micronics 861 | 85.0 | 5.5 | 1.7 | 92 | 0.02 | 31.5 | 3672 | 3.8
1.2 | 59 | 0.02 | 24.5 | 4461 | 4.0
1.1 | 45 | 0.02 | 20.4 | 4784 | 3.8
1.1 | 31 | 0.03 | 14.5 | 5002 | 3.8
6.9 | 1.7 | 62 | 0.03 | 30.4 | 5052 | 4.3
1.3 | 50 | 0.03 | 25.5 | 5518 | 4.2
1.5 | 49 | 0.03 | 20.0 | 4293 | 4.0
1.0 | 36 | 0.03 | 15.0 | 4503 | 4.0
Gold/pyrite/ arsenopyrite concentrate | 80.0 | 1.7 | 33 | 0.05 | 21.0 | 6046 | 5.1
Zinc concentrate | Micronics 889 | 74.0 | 5.5 | 3.5 | 85 | 0.04 | 32.5 | 2906 | 6.1
2.4 | 55 | 0.04 | 25.0 | 3489 | 6.1
2.1 | 37 | 0.06 | 20.0 | 4124 | 6.3
1.4 | 27 | 0.05 | 14.5 | 4029 | 6.0
6.9 | 2.3 | 56 | 0.04 | 30.5 | 4113 | 6.0
1.8 | 55 | 0.03 | 25.5 | 3476 | 5.7
1.8 | 56 | 0.03 | 20.4 | 2858 | 5.2
1.3 | 33 | 0.04 | 14.0 | 3364 | 5.1
Zinc concentrate | 64.0 | 1.9 | 39 | 0.05 | 20.4 | 3764 | 7.2
Lead concentrate | Micronics 861 | 78.0 | 5.5 | 3 | 72 | 0.04 | 32.0 | 4635 | 5.2
2 | 55 | 0.04 | 24.5 | 4724 | 5.1
2 | 42 | 0.04 | 20.5 | 4969 | 5.0
1 | 30 | 0.04 | 14.0 | 5200 | 5.0
6.9 | 2 | 45 | 0.05 | 29.5 | 6697 | 5.2
2 | 38 | 0.04 | 24.3 | 6671 | 5.7
1 | 26 | 0.05 | 19.0 | 7295 | 5.9
1 | 24 | 0.04 | 14.0 | 6268 | 6.0
Lead concentrate | 68.0 | 3 | 40 | 0.08 | 19.0 | 4660 | 5.7

Notes:
1 Filter throughput predictions versus test conditions using raw test data. Throughputs are calculated based on cycle time of form and dry only. Results are not for sizing of any specific type of filter. Refer to individual test results for additional sizing information.
2 Common cake characteristics for all samples: Cohesive, Good cake release from cloth.
Modified ABA testing results show the final tailing sample as uncertain about acid generation potential. Although NAG testing reported no net acidity generated, due to the elevated sulphide content, the single peroxide addition may not have completely oxidized the sulphide present. Sequential NAG tests using multiple peroxide additions were noted as being required to confirm this assumption. Analysis of the NAG liquor reported arsenic at an environmentally significant concentration.
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13.2.3.6
Transportable Moisture Level Testing
A portion of the lead, zinc, and gold concentrates produced from the large batch flotation testwork was used for determining the transportable moisture level. Results are summarized in Table 13-18.
Table 13-18: Transportable Moisture Level Test Summary

Sample | P80, µm | Flow moisture, % | Transportable moisture, %
Lead Concentrate | 25 | 11.4 | 10.3
Zinc Concentrate | 23 | 13.8 | 12.4
Gold/pyrite/arsenopyrite Concentrate | 109 | 9.2 | 8.3

13.3
Conclusions and Recommendations
13.3.1
Conclusions
●
The main sulphide minerals for the Fresh Ore and Old Ore samples were galena, sphalerite, pyrite, and arsenopyrite. Other than quartz, all other minerals can be considered minor.
●
The liberation of all sulphide minerals was excellent at a P80 of 120 µm.
●
Flotation of galena, sphalerite and pyrite / arsenopyrite in a sequential flowsheet was found to be effective at producing lead, zinc and gold concentrates.
●
Lead flotation was found to work well with Aerophine 3418A, with lime and sodium cyanide.
●
Copper activation and zinc flotation were found to be optimum at a pH of 11.8 with 600 g/t CuSO4.5H2O and 60 g/t SIPX.
●
Flotation of pyrite and arsenopyrite was effective at pH 6.0 and 250 g/t SIPX.
●
The rejection of arsenic in the Fresh Ore sample and antimony in the Old Ore sample was found to be problematic in the lead concentrate. It was extremely difficult to reduce the content of arsenic in the lead concentrate from the Fresh Ore to less than 3%. Antimony was not an issue with the Fresh Ore sample but was identified with the Old Ore sample as being present in the lead concentrate at a grade of approximately 3%.
●
The locked cycle test indicated a high amount of lead (6.64%) reporting to the zinc concentrate and a high amount of zinc (0.72%) reporting to the gold concentrate.
●
A large batch flotation campaign was completed to produce tailings for backfill testing by a third-party. Issues with zinc flotation were identified. The lead and zinc concentrates produced were relatively low grade and the recovery was low.
●
All flotation products responded well to BASF Magnafloc 10 flocculant.
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●
Tests confirmed good thickening characteristics. Thickener overflow was visually clear with total suspended solids (TSS) of 62 mg/L, 10 mg/L, and 22 mg/L for the final tail, gold/pyrite/arsenopyrite concentrate and lead concentrate, respectively.
●
Final tail, zinc concentrate and lead concentrate underflows displayed Bingham plastic rheological behaviour characterized by yield stress.
●
The critical solid density was ~68% for the final tail underflow, ~74% for the zinc concentrate underflow, and ~80% for the lead concentrate underflow. The yield stress for the gold/pyrite/arsenopyrite concentrate ranged from 22 to 989 Pa between 86.1% and 87.7% solids w/w.
●
Vacuum filtration of the final tail underflow showed high throughput and low residual cake moisture, which ranged from 13.6% to 22.3%.
●
Pressure filtration on all concentrate underflows indicated high throughput and low residual moisture of 3.8%~4.3% for the gold/pyrite/arsenopyrite concentrate, 5.1%~6.3% for the zinc concentrate, and 5.0%~6.0% for the lead concentrate.
The environmental characterization of the final tailing sample indicated the following:
●
The final tailing was comprised primarily of quartz with minor amounts of calcium, iron, and aluminum and environmentally significant concentrations of arsenic, lead, and zinc.
●
The TCLP leachate reported arsenic and lead at concentrations expected to be of environmental concern.
●
The SPLP extraction leachates reported a strongly alkaline pH value and concentrations of arsenic and lead that may be environmentally significant.
●
ABA testing indicated uncertain acid generation potential for the final tailing.
●
NAG testing of the final tailing reported no net acidity generated.
●
Analysis of the NAG liquor reported arsenic at an environmentally significant concentration.
13.3.2
Recommendations from Metallurgical Testwork
●
Column flotation which involves a high degree of froth washing was recommended to aid in the rejection of arsenic in the lead concentrate in cleaner flotation stages.
●
Mineralogy of the lead concentrate indicated that a portion of the remaining arsenic is present as arsenopyrite locked with galena grains. The use of two regrinding stages in the lead cleaning circuit was found to benefit arsenic rejection and this was recommended to be investigated further as well as optimization of depressant schemes.
●
The high content of antimony in the lead concentrate from the Old Ore sample was due to the high content of antimony-bearing minerals such as boulangerite, bournonite, and tetrahedrite that appear to report to the lead concentrate. These minerals contain copper and lead, and thus were expected to behave similarly to chalcopyrite and galena. Further testwork for depression of antimony bearing minerals was recommended as the lead concentrate containing appreciable amounts of antimony would face significant penalty charges.
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●
Further study was recommended on replacing the lime addition with soda ash in the primary grind and a lower dosage of CuSO4 to overcome the formation of an extremely dry and flat froth causing low zinc rougher recovery.
●
In light of the high pH requirements for selective flotation in the zinc rougher circuit, it was recommended that flotation of pyrite / arsenopyrite at the naturally elevated pH be evaluated in order to reduce sulphuric acid consumption.
●
The locked cycle test identified the issues with excessive lead and zinc reporting to the zinc and gold concentrates, respectively. Further evaluation of the flowsheet, specifically the cleaner scavenger conditions and retention times with locked cycle testing, was recommended
●
Both Old Ore and Fresh Ore samples had challenges with rejection of arsenic and antimony, which are smelter penalty elements. Further testing with different samples was recommended to ascertain how widespread the issue of arsenic and antimony reporting to the lead concentrate is, and if any other issues arise with zinc and gold flotation circuits.
●
It was noted that during dynamic thickening testing, there was an observed sensitivity of the flow response of the underflows to the solid’s residence time. This could lead to bed over-compaction or underflow plugging during extended residence times. Thickener design and operation was advised to be carefully controlled to mitigate any issues relating to this phenomenon.
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SECTION ● 14
Mineral Resource Estimates
14.1
Introduction
Mineral resource estimates for Olympias mine were made from a 3D block model utilizing MineSight 3D software. Project limits, in UTM coordinates, are 478105 to 479700 East, 4491165 to 4493480 North and -800 to +60m elevation. Cell size for the project was 5 m east x 5 m west X 5 m high.
Three areas (East, West, and Flats) have been included in the mineral resource estimate. Wireframes were prepared in LeapFrog software to model these three zones. Block grades for Au, Ag, As, Pb, Zn, and Fe were estimated using inverse distance to the power of 4 (ID4). SG and Sulfur grades calculated by formula and are of the function of the Pb, Zn, As and Fe grade estimations.
14.2
Mineralized Domains
The historical approach to creating mineralized domains at Olympias involved defining limits to what was thought as the dominant mineralization style: massive sulphide bodies. The relogging effort of the pre-existing core and the growing database of infill drilling and mapped stopes showed this assumption to be simplistic. The sulphide mineralization at Olympias, as described in Section 8, comprises many styles and these styles often change abruptly over the scale of mining. Instead, a grade based discriminant would allow for more consistent interpretations to be made. The challenge was to find the formulation of this discriminant because the Olympias deposit contains multiple revenue-contributing metals: Au, Ag, Pb and Zn. In the end, a simplistic value formulae based on the logic of a Net Smelter Return formulae, that used a combination of metal prices and metal recoveries to act as weighting factors against each metal, showed to be an excellent surrogate for a comprehensive equivalent grade. Inspection of these resource defining values (RDV) showed that for the parameters used, a value of $50 best defined what one would classify as likely economically mineralized zones. Seequent’s Leapfrog GeoTM software was used to model the mineralized domain wireframes.
For the Olympias modeling, the deposit was divided into three zones: East, West, and Flats, shown in plan view in Figure 14-1. Within each of these zones, modeling domains were created using the $50 RDV. Assays and composite samples were tagged by these domain shapes ahead of data analysis and grade interpolation.
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Figure 14-1: Ore Zones, Olympias Mine
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14.3
Data Analysis
The dominant assay sample length inside the mineralization shapes is 1 m. This then became the compositing length. Assay samples were composited to 1 m lengths within each mineralization lenses. During compositing, samples less than 0.25 m in length were merged with the previous sample.
Assays were also composited to 5 m lengths within each mineralization lens. The 5 m composites are used in the nearest neighbour estimation for model validation purposes.
Statistical properties of uncapped 1m composite data for gold, silver, lead and zinc, by zone are shown in Table 14-1 to Table 14-4
Table 14-1: Uncapped Gold Composite Statistics for each Zone

ZONE | No of Comps | Mean | CV | SD | q25 | q50 | q75 | Maximum
EAST | 3511 | 14.29 | 0.97 | 13.81 | 4.15 | 10.44 | 20.56 | 137.1
FLAST | 2889 | 7.46 | 1.03 | 7.66 | 2.16 | 5.24 | 10.54 | 111.7
WEST | 5086 | 8.52 | 1.03 | 8.75 | 3.01 | 6.14 | 11.15 | 140.3

Table 14-2: Uncapped Silver Composite Statistics for each Zone

ZONE | No of Comps | Mean | CV | SD | q25 | q50 | q75 | Maximum
EAST | 3510 | 166 | 1.2 | 197 | 23 | 97 | 238 | 1460
FLAST | 2889 | 87 | 1.5 | 134 | 8 | 31 | 112 | 1410
WEST | 5078 | 139 | 1.3 | 175 | 21 | 76 | 193 | 1580

Table 14-3: Uncapped Lead Composite Statistics for each Zone

ZONE | No of Comps | Mean | CV | SD | q25 | q50 | q75 | Maximum
EAST | 3511 | 5.16 | 1.16 | 5.97 | 0.78 | 3.14 | 7.30 | 43.0
FLAST | 2889 | 3.18 | 1.46 | 4.64 | 0.31 | 1.35 | 4.12 | 39.2
WEST | 5083 | 4.47 | 1.23 | 5.48 | 0.71 | 2.56 | 6.38 | 56.4

Table 14-4: Uncapped Zinc Composite Statistics for each Zone

ZONE | No of Comps | Mean | CV | SD | q25 | q50 | q75 | Maximum
EAST | 3511 | 5.65 | 1.08 | 6.10 | 0.60 | 3.42 | 8.93 | 33.1
FLAST | 2889 | 4.83 | 1.13 | 5.47 | 0.65 | 2.76 | 7.14 | 35.3
WEST | 5083 | 5.37 | 1.08 | 5.79 | 0.88 | 3.45 | 7.96 | 41.2

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14.3.1
Evaluation of Extreme Grades
Outlier sample grades can cause overestimation in the resource model if left untreated. Extreme grades were examined for gold, silver, lead and zinc by means of cumulative probability plots and histograms for each element. Generally, 1.0 to 1.5% of the samples showed susceptibility to extreme grades. The cap value and number of samples capped are shown in Table 14-5.
Table 14-5: Top-Capping Grades

| Au (g/t) | Ag (g/t) | Pb (%) | Zn (%) | As (%) | Fe (%)
Cap Value | 50 | 900 | 25 | 25 | 25 | 35
No. Samples | 12,673 | 12,654 | 12,667 | 12,669 | 12,665 | 11,725
No. Capped | 153 | 134 | 174 | 172 | 190 | 173
Percent capped | 1.21% | 1.06% | 1.37% | 1.36% | 1.50% | 1.48%

Statistical properties of capped 1m composite data for gold, silver lead and zinc, by zone, are shown in Table 14-6 to Table 14-9.
Table 14-6: Capped Gold Composite Statistics for each Zone

ZONE | No of Comps | Mean | CV | SD | q25 | q50 | q75 | Maximum
EAST | 3511 | 13.83 | 0.87 | 12.09 | 4.15 | 10.4 | 20.55 | 50
FLATS | 2889 | 7.41 | 0.99 | 7.32 | 2.16 | 5.24 | 10.54 | 50
WEST | 5086 | 8.42 | 0.95 | 8.02 | 3.01 | 6.14 | 11.15 | 50

Table 14-7: Capped Silver Composite Statistics for each Zone

ZONE | No of Comps | Mean | CV | SD | q25 | q50 | q75 | Maximum
EAST | 3510 | 163.1 | 1.1 | 186.9 | 22.6 | 96.5 | 237.5 | 900
FLATS | 2889 | 86.5 | 1.5 | 130 | 8 | 31 | 112.2 | 900
WEST | 5078 | 136.4 | 1.2 | 163.6 | 21 | 76.4 | 193.1 | 900

Table 14-8: Capped Lead Composite Statistics for each Zone

ZONE | No of Comps | Mean | CV | SD | q25 | q50 | q75 | Maximum
EAST | 3511 | 5.04 | 1.09 | 5.51 | 0.78 | 3.14 | 7.27 | 25
FLAST | 2889 | 3.13 | 1.4 | 4.37 | 0.31 | 1.35 | 4.12 | 25
WEST | 5083 | 4.36 | 1.14 | 4.97 | 0.71 | 2.56 | 6.36 | 25

Table 14-9: Capped Zinc Composite Statistics for each Zone

ZONE | No of Comps | Mean | CV | SD | q25 | q50 | q75 | Maximum
EAST | 3511 | 5.62 | 1.07 | 6 | 0.6 | 3.42 | 8.93 | 25
FLAST | 2889 | 4.79 | 1.11 | 5.34 | 0.65 | 2.75 | 7.14 | 25
WEST | 5083 | 5.32 | 1.05 | 5.59 | 0.88 | 3.45 | 7.95 | 25

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14.4
Block Model
The mineral resource was estimated using a block model with limits and size as shown in Table 14-1. MineSight software was used to model the deposit. The block size selected based on the minimum unit for drift and fill mining. Ore blocks were coded to have domain percentages to honor the wireframe boundaries; coded with zones and domain values to control hard boundary between mineralization units and to define various search parameters for the interpolation stage.
Table 14-10: Block Model Parameters

| Minimum (m) | Maximum (m) | Block size (m) | Number of blocks
East | 478,105 | 479,700 | 5 | 319
North | 4,491,165 | 4,493,480 | 5 | 463
Elevation | -800 | 60 | 5 | 172

14.5
Grade Estimation
Grade estimates for Au, Ag, As, Pb, Zn and Fe were interpolated using an inverse distance to the power of 4 (ID4) method. Nearest–neighbour (NN) grades were also interpolated as a declustered distribution to validate the estimation method.
The search orientation and radii were based on drillhole spacing and the orientation and size of each mineralized zone.
A multi-pass approach was instituted for interpolation. The first pass required a grade estimate to include composites from a minimum of two holes from the same estimation domain, whereas the second pass allowed a single hole to place a grade estimate in any uninterpolated block from the first pass.
For the East zone, three estimation passes were used to ensure each block had all grades estimated. During the first pass estimation, the search distances used was 60 x 60 x 10 m in X, Y and Z directions respectively. At this step of the estimation, model blocks received a minimum number of 4 composite samples, maximum of 3 composite from a single hole and total maximum number of composites adjusted as 12. The second pass, which aimed to capture more of the surrounding nearby samples for the interpolation to provide a good local estimation, had a smaller search ranges of 30 x 30 x 5 m in X, Y and Z direction. Maximum number of samples per hole varied between 2 to 5 for different elements, and maximum samples varied from 12 to 24 composite samples. The minimum number of samples required varied between 3 to 6 samples for the different elements. A final large search was employed to place grade for those remaining uninterpolated blocks using a 10-sample maximum for the interpolation and 4-sample maximum per hole.
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The Flats zone contains seven separate mineralization lenses. For each lens a large search radius was used as a first pass to ensure each block had a grade estimated. For this pass, blocks received a minimum of 4 and a maximum of 18 composites.
Search distances for the second pass run consisted of 70X x 70Y x 20Z m and lay in the direction of mineralization. Blocks required to have minimum 5 and maximum 16 composite samples for this interpolation, with a maximum 4 samples allowed from a single hole.
In the West zone, the short range estimation ellipse, comprised 70X x 70Y x 15Z m. Blocks required to have minimum 4 and maximum 18 composites for this interpolation, with a maximum 3 samples allowed from a single hole. The longer ranged pass had a minimum 4 and maximum 12 composites protocol.
In East and West zones, an outlier restriction was used to control the effects of high-grade composites for gold in local areas of less dense drilling. The restricted distances ranged 20 and 55 m with threshold gold grades of 30 g/t (West) and 45 g/t (East).
Bulk density in the block model is function of the estimated Pb, Zn, As and Fe grades. Values for each model block were calculated following the description of the method in Section 10.
14.5.1
Areas of Previous Mining
Areas of previous mining are located in the East and West Zones. Mined out areas do not form part of the stated mineral resource estimate. Figure 14-2 is a plan showing the location of the mineral resources and areas mined out.
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Notes:
●
Yellow colour equals area of mineral resources.
●
Green colour represents areas of pre-Eldorado mining.
●
Red colour shows areas mined by Eldorado.
●
Mine development shown with blue lines.
Figure 14-2: Location of Areas of previous mining
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14.6
Block Model Validation
14.6.1
Visual Inspection
Eldorado completed a detailed visual validation of the resource model. Models were checked for proper coding of drill hole intervals and block model cells, in both section and plan. Coding was found to be properly done. Grade interpolation was examined relative to drill hole composite values by inspecting sections and plans. The checks showed good agreement between drill hole composite values and model cell values. Examples of representative views of block model grades (Au) and drill hole composites for each zone are shown in Figure 14-5.
Figure 14-3: 3D Oblique View of EAST Zone comparing Block Au Grades to Composite Au Grades
Figure 14-4: 3D Oblique View of FLATS Zone comparing Block Au Grades to Composite Au Grades
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Figure 14-5: 3D Oblique View of WEST Zone comparing Block Au Grades to Composite Au Grades
14.6.2
Model Checks for Bias
Eldorado checked the block model estimates for global bias by comparing the average metal grades (with no cut-off) from the ID4 models with means from nearest-neighbour estimates. The nearest-neighbour estimator declusters the data and produces a theoretically unbiased estimate of the average value when no cut-off grade is imposed. It is a good basis for checking the performance of different estimation methods. Results, summarized in Table 14-11 show no global bias in the estimates.
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Table 14-11: Global Model Mean Grade Gold Values (g/t) by Domain

ZONE | ID4 | NN Estimate | Difference
| AU | AUNN | %
EAST | 13.82 | 13.66 | 1.2%
FLATS | 6.97 | 6.93 | 0.5%
WEST | 8.61 | 8.29 | 3.7%
| AG | AGNN | %
EAST | 169 | 172 | -2.0%
FLATS | 136 | 140 | -2.8%
WEST | 149 | 150 | -0.6%
| ZN | ZNNN | %
EAST | 6.4 | 6.2 | 2.1%
FLATS | 6.5 | 6.8 | -4.4%
WEST | 5.3 | 5.3 | -0.3%
| PB | PBNN | %
EAST | 5.4 | 5.5 | -1.1%
FLATS | 4.7 | 4.9 | -3.9%
WEST | 4.7 | 4.7 | 0.0%
| AS | ASNN | %
EAST | 6.8 | 6.7 | 1.3%
FLATS | 3.5 | 3.5 | -1.5%
WEST | 4.0 | 3.8 | 6.2%
| FE | FENN | %
EAST | 14.7 | 14.6 | 0.8%
FLATS | 14.6 | 14.7 | -0.4%
WEST | 15.4 | 15.2 | 1.1%

14.6.3
Swath Plots
Eldorado also checked for local trends in the grade estimates. This was done by plotting the mean values from the nearest-neighbour estimate versus the ID4 modeled results for benches, eastings and northings in 20 m swaths. Examples for Au, Pb and Zn, along eastings of all three zones are shown in Figure 14-6 to Figure 14-8. Since 5 m composites were used for nearest neighbour (NN) estimation, distribution is already smoothed; therefore, model estimates should be closer or slightly smoother than the nearest-neighbour estimate. The observed trends behave as predicted and show no significant trends in the estimates in models.
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Figure 14-6: Au SWATH Plots
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Figure 14-7: Pb SWATH Plots
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Figure 14-8: Zn SWATH Plots
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14.7
Mineral Resource Classification
The mineral resource was classified as measured, indicated and inferred using logic consistent with the CIM Definition Standards referred to in NI 43-101. The mineral resource was classified as measured, indicated or inferred, based on:
●
Location and number of drillholes
●
Location of blocks close to areas that were mined
The measured mineral resources were generally located near areas of active mining and within approximately 15 m of at least three drillholes. The indicated mineral resources were generally located a maximum of 45 m from two drillholes (actual average distance is 30 to 35 m). Blocks that honored these conditions for measured and indicated resources were examined in longitudinal section. Polygons were digitally drawn around contiguous areas of appropriately tagged blocks for each class type. These shapes were subsequently used to formally classify the model blocks as measured or indicated mineral resources. All remaining interpolated blocks were classified as inferred mineral resources.
Figure 14-9 to Figure 14-11 show plan views of classified blocks relative to drill holes for each zone.
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Figure 14-9: East Zone Mineral Resource Classification
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Figure 14-10: West Zone Mineral Resource Classification
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Note: Measured Mineral Resources are not defined in the Flats Zone.
Figure 14-11: Flats Mineral Resource classification
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14.8
Mineral Resource Estimate
The Olympias mineral resources as of September 30, 2019 are shown in Table 14-12. All material contained within the mineralization domains, defined by a RDV of $50, are reported.
Table 14-12: Olympias Mineral Resources, as of September 30, 2019

Classification | Tonnes (Kt) | Au (g/t) | Au (Koz) | Ag (g/t) | Ag (Koz) | Pb (%) | Pb (Kt) | Zn (%) | Zn (Kt)
Measured | 2,702 | 10.93 | 950 | 156 | 13,552 | 5.0 | 135 | 6.0 | 162
Indicated | 11,779 | 7.52 | 2,848 | 134 | 50,746 | 4.6 | 542 | 6.2 | 730
Measured and Indicated | 14,481 | 8.16 | 3,798 | 138 | 64,298 | 4.7 | 677 | 6.2 | 892
Inferred | 3,720 | 7.98 | 954 | 137 | 16,385 | 3.9 | 145 | 4.0 | 149

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SECTION ● 15
Mineral Reserve Estimates
The mineral reserve estimates conform to CIM Definition Standards referred to in NI 43-101. All design and scheduling have been completed using the mineral resource model and estimate described in Section 14. The mineral reserves are as of September 30, 2019.
Only measured and indicated resources have been used for mineral reserves estimation. The estimation assumes that the mining methods employed at the mine will be drift and fill (DAF) and transverse longhole open stoping (TLHOS).
The proven and probable underground mineral reserves for the Olympias deposit are summarized in Table 15-1.
Table 15-1: Olympias Mineral Reserves as of September 30, 2019

Class | Tonnes (kt) | Au (g/t) | Au (Moz) | Ag (g/t) | Ag (Moz) | Pb (%) | Pb (kt) | Zn (%) | Zn (kt)
Proven | 2,601 | 9.19 | 0.77 | 133 | 11.1 | 4.3 | 112 | 5.1 | 133
Probable | 10,324 | 6.47 | 2.15 | 115 | 38.2 | 4.0 | 413 | 5.3 | 547
Total | 12,925 | 7.02 | 2.92 | 119 | 49.3 | 4.1 | 525 | 5.3 | 680

Notes:
●
Mineral reserves are included in measured and indicated mineral resources.
●
Figures in the tables may not compute due to rounding.
●
The mineral reserves are based on a planning cut-off grade of $133/t for DAF and $116/t for TLHOS, and $50/t for development ore.
●
Cut-off grades are based on a gold metal price of $1,250/oz, silver metal price of $16/oz, zinc metal price of $2,000/t, and lead metal price of $1,800/t.
●
Metallurgical recoveries are based on feed grade and metallurgical algorithms.
●
Exchange rate used is €1.20 = US$1.00.
●
Average mining dilution and mining recovery factors of 14% and 95%, respectively, for LHOS, and 13% and 98%, respectively, for DAF are assumed.
●
Mineral reserves are reported on a 100% ownership basis; Eldorado owns 95% of Hellas Gold.
15.1
Cut-Off Grade
The cut-off values supporting the estimation of underground mineral reserves were developed in 2018 and based on future projected operating costs at a steady-state production rate of 650,000 tonnes per annum. The operating cost assessment indicated that NSR values of $133/t for DAF mining and $116/t for TLHOS mining would adequately cover all site operating costs on a breakeven basis. The weight averaged operating cost can be estimated at $125/t considering the balance between TLHOS and DAF. These DAF and TLHOS cut-offs were used to create potentially mineable stope shapes from the NSR block model (NSR BM). The NSR BM was created by Eldorado and is based on metallurgical recovery experience and historical sales and includes recognition of transport and refining costs.
It is recognized that the Olympias mine is currently in a ramp up phase and has not yet achieved the proposed production rate and mining cost targets. Table 15-2 contains target costs over the peak production phase (2024 – 2038) when Olympias will be operating at full 650 ktpa capacity.
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Table 15-2: Target Operating Costs

Category | Target
Mining Cost | 53.30
Process Cost | 50.40
G&A Cost | 21.30
Total | 125.00

The operational improvement plan includes the following elements:
●
Increased equipment availability
●
Upskilling and multiskilling the workforce
●
Increase in manpower utilization
●
Maximizing the use of paste backfill
●
Infrastructural projects to increase underground mining efficiency:
o
Underground workshop
o
Slick line to deliver shotcrete
o
Fuel line to underground
o
Underground slimes filtration system
o
Introduction of remote operations and central control system
Eldorado has reviewed the Olympias plans for operational improvements and considers them adequately achievable to support the current mineral reserves estimate. The impact of operational improvements will be closely monitored over 2020 and later years, failure to achieve the target cost performance by 2023 would potentially reduce mineral reserves and shorten the mine life. This is a base case scenario and any subsequent metallurgical processing would enable Eldorado to run another reserve calculation.
Table 15-3 presents the estimated impact to reserves and mine life should it become necessary to raise the breakeven cut-off grade in the future.
Table 15-3: Estimated Impact of Raised Breakeven Cut-off Grade on Reserves and LOM

Class | Tonnes | Tonnes | Tonnes | Tonnes | Tonnes
(kt) | (kt) | (kt) | (kt) | (kt)
Proven | 2,601 | 2,592 | 2,543 | 2,479 | 2,418
Probable | 10,324 | 9,797 | 9,467 | 8,940 | 8,630
Total | 12,925 | 12,389 | 12,010 | 11,419 | 11,048
Cut-off grade (averaged $/t) | 125 | 135 | 145 | 155 | 165
Years of Mine Life | 21 | 20 | 19 | 18.5 | 18

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Eldorado has also reviewed the material planned to be mined during the ramp-up period prior to the steady-state production value of 650 ktpa being achieved and, therefore, prior to the full benefits of the operational improvements being realized. The value of this material validates its inclusion as part of the estimated mineral reserves.
15.2
Mining Shapes
DAF stope shapes were generated by Eldorado using the software mineable shape optimizer (DSO) to produce stope shapes based on the NSR cut-off of $133/t, minimum stope width of 5 m, a fixed stope height of 5 m, and a maximum FW / HW angle of 45º. The TLHOS stope shapes were created by Eldorado using a cut-off of $116/t to generate grade shells. The grade shells were used to manually design the stope shapes.
On the outliers of the ore-body, some marginally economic areas were not included in the mine plan and, therefore, is not included in the mineral reserves. Capital and operating development cost estimates were used to determine economic viability for these areas.
15.3
Dilution and Recovery Factors
In the evaluation of underground mineral reserves, modifying factors were applied to the tonnages and grades of all mining shapes to account for dilution and ore losses. In the DAF stopes, a mining dilution factor of 13% and a mining recovery of 98% were estimated. The sources of dilution for DAF will be floor mucking on fill (estimated 0.1 m), and side wall dilution of fill or waste (estimated 0.1 m on each wall) from overbreak and sloughing.
In the TLHOS stopes (predominantly blind-uppers) a mining dilution factor of 14% and a mining recovery of 95% were estimated. The sources of dilution for TLHOS were considered for primary and secondary stopes. In the primary stopes, there will be waste rock overbreak on the back of the blind uppers stopes and end wall fill dilution from the preceding TLHOS (when applicable). The floor will invariably be on country rock and associated dilution is anticipated to be minimal. For the secondary stopes, there will be waste rock overbreak in the back, and end wall and side wall fill dilution from the preceding secondary stope and adjacent primary stope, respectively.
Ore losses (mining recovery factors) are related to the practicalities of extracting ore under varying conditions, including difficult mining geometry, problematic rock conditions, losses of ore into backfill, and blasting issues. Mining recovery has been estimated to be 98% by weight for DAF and 95% for TLHOS.
Figure 15-1 and Table 15-4 present 2019 reserves dilution estimation.
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Figure 15-1: 2019 Reserves Dilution Estimation
Table 15-4: 2019 Reserves Dilution Estimation

| Sidewall Backfill | Backfill Dilution (%) | Overbreak | Total
Drift and fill
Ore Zone | Frequency | Width (m) | Height (m) | Backfill (m2) | Sidewall | Floor | TOTAL | All | All
Upper East | 50% | 0.3 | 5.0 | 0.75 | 1.9% | 4.8% | 6.7% | 6.0% | 12.7%
Lower East | 75% | 0.3 | 5.0 | 1.13 | 2.8% | 4.8% | 7.6% | 6.0% | 13.6%
Flats D&F | 75% | 0.3 | 5.0 | 1.13 | 2.8% | 4.8% | 7.6% | 6.0% | 13.6%
Middle West | 75% | 0.3 | 5.0 | 1.13 | 2.8% | 4.8% | 7.6% | 6.0% | 13.6%
South West | 50% | 0.3 | 5.0 | 0.75 | 1.9% | 4.8% | 6.7% | 6.0% | 12.7%
LHOS
Flats LH | 50% | 2 | 10.0 | 10.00 | 6.3% | 0.0% | 6.3% | 8.0% | 14.3%

15.4
Reserve Resource Reconciliation
Table 15-5 shows mineral resource to mineral reserve tonnes and metal content conversion factors.
Total mineral reserve tonnes (proven plus probable) are approximately 89% of mineral resource (measured plus indicated) tonnes. The tonnage conversion from measured to proven is 96% and that for Indicated to probable is 88%. Total metal conversion percentages for gold, silver, lead, and zinc are 77%, 77%, 78%, and 76% respectively.
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Table 15-5: Conversion of Mineral Resources to Mineral Reserves

Olympias | Mineral Resources | Mineral Reserves* | Conversion factors*
Measured | Indicated | M+I | Proven | Probable | P+P | Prov / Meas | Prob / Ind | P+P / M+I
Tonnes | kt | 2,702 | 11,779 | 14,481 | 2,601 | 10,324 | 12,925 | 96% | 88% | 89%
Au | g/t | 10.93 | 7.52 | 8.16 | 9.19 | 6.47 | 7.02 | 81% | 75% | 77%
Ag | g/t | 156 | 134 | 138.0 | 133 | 115 | 119 | 82% | 75% | 77%
Pb | % | 5.0 | 4.6 | 4.7 | 4.3 | 4.0 | 4.1 | 83% | 76% | 78%
Zn | % | 6.0 | 6.2 | 6.2 | 5.1 | 5.3 | 5.3 | 82% | 75% | 76%

Notes:
*Tonnes and metal content.
Numbers may not compute exactly due to rounding.
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SECTION ● 16
Mining Methods
16.1
Introduction
The Olympias mine is an underground mining operation extracting ore from three zones: East, West and Flats, and two sub-zones: Remnants, and North. These are shown in Figure 16-1 as an isometric view of the Olympias underground zones looking north-west.
Figure 16-1: Isometric View of Olympias Mining Areas
The East zone is one of three current active mining areas at Olympias mine. It extends approximately 700 m horizontally and has a vertical height of 300 m (-375 masl to -75 masl). The zone dips from 20º to 25º and is mined exclusively using DAF as the mining method. Existing development in the East zone is established at 20 m to 25 m sublevels. The lower East zone (below -240 masl) has not been developed but is included in both short and long-range mine plans.
The West zone is the second active mining area at Olympias; it is also mined entirely as DAF. The zone dips between 45º to 65º and has a horizontal extent of approximately 900 m and a vertical rise of approximately 260 m (-500 masl to -240 masl). Existing development between -280 masl and -240 masl accesses the West zone DAF stopes. Active levels below -280 masl are also currently being mined.
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The third active mining area is the Flats zone. This zone dips from 150 to 200, has a horizontal extent of approximately 1,000 m and elevation difference of approximately 355 m (-275 masl to -630 masl). The current mining area in the Flats is at -280 masl and mining is via DAF. Approximately half of the total mineral reserve tonnage in the Flats is planned to be mined using blind up-hole TLHOS. The Flats currently has limited existing development.
The majority (~90%) of the mineral reserve tonnes will be extracted from these three zones (East, West, and Flats). The remaining ore will be mined from the two other zones, North and Remnants, which are further described below.
The Remnant sub-zone (part of the overall West zone that was mined historically) has a dip of 30º to 50º and is situated on the West side of the Olympias orebody between elevations -235 masl and -130 masl. The zone includes the remaining areas of previously mined DAF stopes, as well as unmined areas that will be removed using the DAF mining method.
The smallest of the mining areas is the North sub-zone that has a dip between 15º and 25º and is situated from -65 masl to -15 masl. This area will be mined as DAF and has no existing development.
Main access to underground is through the existing portal at 70 masl to the north, as well as a portal at 55 masl to the east. An existing shaft (previously utilized for hoisting) extends from surface down to -315 masl. This shaft is not, nor is planned to be, used to transport ore / waste to surface. It is planned to be utilized for ventilation throughout the LOM. The existing underground development plus an increased emphasis on future development will allow the mine to increase the annual production rate from the current 360,000 tonnes up to 650,000 tonnes by 2024. As development progresses, additional stope mining levels will be accessed from lower parts of the mine, in the East, West and Flats, to achieve the production increase to 650,000 tpa.
An average mining rate of 1,800 tonnes per day of ore is planned to reach the steady-state of 650,000 tonnes of ore per year. Mine life is projected to be 21 years for the total mineral reserve of 12.9 million tonnes.
16.2
Geotechnical Considerations
16.2.1
General Description
The ground conditions in the West zone vary considerably. The majority of the poor ground is within the orebody or at the HW contacts due to the presence of the Kassandra fault and strong to intense clay alteration. The fault and weak rock mass strength may cause instability or HW overbreak / sloughing. The West Zone is deemed not to be suitable for longhole mining and the drift and fill method “DAF” is used throughout, with the length of the cut varying depending upon the conditions. The DAF method is also used in the East Zone and in the Remnants area where the dip and the orebody geometry dictate the preferred method rather than ground conditions.
The overall rock mass quality in the Flats zone appears good based on field observation and review of core photos and the analysis of three recently completed geo-technical drill holes in 2019. If the ground conditions observed and the cores examined and from photos, are representative of the overall geotechnical conditions in the Flats zone, longhole mining can be anticipated to be appropriate.
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16.2.2
Longhole Stope Stability
Information obtained from the three 2019 geotechnical drill holes contributed towards a rock mass rating system to quantify the geotechnical properties in prospective longhole areas. Key data from these drill holes are shown in Table 16-1.
Table 16-1: Results from logging of geotechnical drill holes in Flats Zone

Domains | Density (KN/m3) | UCS (MPa) | Poisson’s Ratio | Young’s Modulus (GPa) | Rock Quality Designation (RQD %) | Rockmass Rating (Q system)
GN (Gneiss) | 27 | 132 | 0.31 | 30 | 86.2 | 22.79
MA-10A (Marble) | 26,5 | 74 | 0.31 | 32 | 76.7 | 19.01
MA-IWZ (Marble) | 26.8 | 79 | 0.27 | 38
MZ (Massive Sulphides | 37.5 | 62 | 0.28 | 27 | 45.3 | 12.28
PE (Pegmatite) | 25.6 | 132 | 0.25 | 29 | 77.5 | 44.38

These data allowed for an empirical stope stability analysis based on the stability graph method. This method considers the geometry of the stope plane (the wall, the back or the end) as described by its hydraulic radius, and the geotechnical condition expressed as a modified rock mass rating designation N’, which takes into account the orientation and persistence of joint sets, the orientation of the stope plane which is being evaluated, and other geotechnical parameters. The results obtained from the data are plotted against empirical data collected and normalized to fit the model. This creates a spectrum of stable to unstable spans and are a means of determining the predicted stability and the type of ground support if required.
The inputs into the stability graph method used to analyze the proposed longhole stopes in the Flats zone based on the average geotechnical parameters are shown in Table 16-2.
Table 16-2: Inputs to the Stability Graph Method

Stope Span | Span | Length | UCS (Mpa) | Stress (Mpa) | RQD | Jn | Jr | Ja | Q’ | A | B | C | N’ | Stope HR | Stable HR
Backs | 10 | 20 | 80 | 36 | 87 | 4 | 1.5 | 1.0 | 32.5 | 0.12 | 0.3 | 2.0 | 2.4 | 3.3 | 3.3
Walls | 25 | 20 | 60 | 36 | 60 | 4 | 1.5 | 2.0 | 11.3 | 0.1 | 1.0 | 8.0 | 9.0 | 5.6 | 5.3
Ends | 25 | 10 | 60 | 36 | 60 | 4 | 1.5 | 2.0 | 11.3 | 0.1 | 1.0 | 8.0 | 9.0 | 3.6 | 5.3

The resultant stability graph is shown in Figure 16-2. The analysis shows that the proposed longhole stoping geometries will be stable with up to 20 m height and a span width of up to 10m. The results using the averages from the rock mass ratings suggest that support will not be necessary to maintain these stope geometries although both the sidewall and the backs of the stopes points lie close to the support necessary transition zone boundary.
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Figure 16-2: Empirical Stope Stability Analysis using 50th percentile Geotechnical Data
Since no longhole open stoping has been done at Olympias mine since the current Phase 2 has been operating (it was used by previous owners) and is currently scheduled to take place in Olympias in the Flats zone. Eldorado is looking to test longhole stoping to confirm the geotechnical behavior of likely areas for LHOS.
16.2.3
Ground Support
Field observations supports the current Olympias ground support system. The implementation of ground control at the mine appears adequate, with no unsecured ground being noted and support performance apparently good. Regular QA/QC (rock bolt pull testing and shotcrete compressive testing) is conducted to ensure quality installation of the ground support and the effectiveness of the support system.
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Based on the rock conditions at Olympias mine, the ground support class of Fair to Good may be predominantly applied. The underground openings appear stable, being generally supported with 50 mm of shotcrete and 2.4 m long Swellex bolts on a 1 m by 1.5 m spacing. The dominant condition in ore is similar to slightly weaker than that of the hosting waste rock.
Pull testing on 1% of installed bolts is carried out and the overall bond performance of Swellex bolts in poor ground is satisfactory, which implies that the poorer quality of rock mass is of limited width and extent. In particular it appears that any clay alteration is not extensive and the frictional bolts have a good bond strength at the interface with competent rock. Therefore, the integrated ground support (shotcrete and frictional bolts) is seen to be effective in the observed weak ground and mixed ground conditions.
16.3
Mining Methods
The mining methods planned for future operations at Olympias mine are DAF (the current method) and TLHOS. DAF is envisaged to be used in all mining areas (East, West, Flats, Remnants, and North), whereas TLHOS is projected for use in the Flats. The overall geometry of the Olympias orebody and the planned locations of the two types of mining methods are shown in Figure 16-3 as an isometric view of the Olympias underground zones looking north-west. The beige colouring indicates planned areas of DAF, the magenta TLHOS.
Figure 16-3: Isometric View of Olympias Underground – Mining Methods
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16.3.1
Blind-Uppers Transverse Longhole Open Stoping (TLHOS)
Blind-uppers TLHOS will be utilized only in the Flats zone. Flats are gently dipping (15º to 20º) tabular lenses with thicknesses up to 30 m. TLHOS is to be used in over 50% of this zone.
An isometric view of typical primary and secondary TLHOS stopes is shown in Figure 16-4. The ore is accessed through a FW drive and by individual accesses to each stope from the drive. The accesses are driven to the ore contact, then the remaining development in ore is driven at 5.0 m W x 5.0 m to the FW contact. A blind-uppers (conventional or large diameter) raise is driven at the end of the TLHOS followed by longhole drilling of the production rings.
Slot blasting will occur once the raise and production rings are completed and after the backfill curing of the preceding stope. The ore will be mucked from the stope undercut drive with a mixture of manual mucking when the brow is choked with ore, and remote scoop operation when it is not choked-off. Confirmation of complete extraction of the stope by operations supervision will initiate technical services to complete the stope cavity monitor survey (CMS). After the CMS, an engineered fill fence will be constructed back from the brow of the stope and paste filling will proceed until the stope is tight-filled. Other stopes will be mined following the same sequence.
Based on the current geotechnical assessment the primary and secondary stope dimensions will be 10 m wide (constant), with the maximum height being 30 m (undercut floor elevation up to the back of the blind uphole stope). The designed maximum unsupported stope strike length varies based on the stope height and average rock quality for the stope back.
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Figure 16-4: View of typical primary secondary Longhole Stopes in the Flats
16.3.2
Drift and Fill (DAF)
DAF at Olympias commences once the decline reaches the footwall drive or Level access elevation of the orebody, usually midway along its strike length (see representative sketch as Figure 16-5). DAF is an overhand mining method. The stope sequence begins with the lowest 5.0 m high lift. Then each subsequent lift requires the back of the level access to be slashed down (‘take-down-back or TDB) to reach the next lift. There are four lifts between levels for a total rise of 20 m from each access.
Future DAF mining is expected to be either in single or multiple passes depending on orebody geometry; this is further described below.
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Figure 16-5: Longitudinal DAF
16.3.2.1
Longitudinal DAF
Generally, for FW to HW thickness of 10 m or less the stope will be developed as longitudinal DAF. The mining begins by driving the Level access to the FW contact of Lift 1 and then the drive is extended flat (zero gradient) to the HW contact of the ore. Next, the ore is mined longitudinally in a single pass along strike in both directions to the limits of the orebody. Any remaining ore on the HW side will be slashed out on retreat and then the drift will be backfilled with pastefill (PF) or lesser cemented aggregate fill (CAF). After the fill has cured, the Level access will be slashed down to provide access for the next lift, and the process will be repeated for subsequent lifts.
16.3.2.2
Multi-Pass DAF
Multi-pass DAF is generally for orebody widths greater than 10 m from FW to HW. The mining begins by driving the Level access to the FW contact of Lift 1 (Figure 16-6). Then the drift is extended to the HW contact of the orebody. The Primary drift will be mined (note: generally, all multi-pass DAF drives are planned to be 5.0 m H x 5.0 W) on the FW contact to the extremities of the orebody. The drift will then be filled mainly with PF. After the fill has cured, the Secondary drift will be developed parallel to the Primary drift, with fill on one side and ore on the other side of the drift. After filling and curing is complete for the Secondary, a Tertiary drift is driven beside the Secondary drift, followed by filling and curing. This process is repeated until the HW is reached, which is the Quaternary drift in Figure 16-6. Once the fill has cured for the Quaternary drift, the Level Access will be slashed down to reach the next lift and the process will be repeated for the remaining lifts of multi-pass DAF.
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Figure 16-6: Multi-Pass DAF Stope
16.4
Backfill
Olympias is currently using CAF and PF for underground backfill operations. The PF system is intended to provide most future backfill for the mine.
The CAF is prepared in a batch plant adjacent to the portal. Aggregate and sand are combined with 8% binder to produce CAF with a 7-day strength of 1 MPa and 28 day-strength of 4.0 MPa. The CAF is backhauled in 40 t trucks and tipped at the entrance to the DAF production area. The CAF is then placed with a loader and, using a pusher arm attachment, tight filled to the back of the drift.
The PF system has been designed to meet future backfill requirements up to the 650 ktpa production rate and operating at up to 70% utilization. PF will be used in DAF and longhole stopes. Using dewatered flotation tailings and cement, the plant can produce 42 m3/hr of PF. When backfill is not required, dewatered tailings are stockpiled and reclaimed with a front-end loader during the next fill run. Paste can be placed in continuous runs of up to 2,500 m3.
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PF is delivered by positive displacement pump via drill holes and pipelines to the stopes. For control purposes, pressure is monitored in the pipelines by instrumentation. PF is retained in stopes using shotcreted barricades. Tight filling is achieved by managing delivery via in-stope pipelines or drill holes, and location of barricades.
The current operational permit dictates a backfill directive for Olympias that recommends using a strength of 4 MPa for all backfill placed underground. Using best available design practice, Eldorado calculates that a maximum strength for 5 m x 5 m wide drift sill beams, at a Factor of Safety (FOS) of 3.0, should be 1.4 MPa and suggests that the 4.0 MPa requirement is overly conservative.
The backfill directive also states that vertical walls of cured backfill must not be exposed until the strength has reached 1.0 MPa. Using best available design practice, Eldorado calculates the maximum strength for 5 m high x 5 m wide drifts to be only 300 kPa at a FOS of 1.5.
Eldorado recommends that Olympias mine initiate a backfill geotechnical study to optimize the strength requirements of undercut sills and drift sidewalls using a combination of empirical and numerical modelling techniques. It further recommends that Olympias then prepare a technical brief for submission to the relevant authorities on an optimized backfill strength design, using current industry best practice, for stable openings of cured fill.
16.5
Mine Design
16.5.1
Underground Access
A FW decline from the surface Portal at +70 m down to ‐300 masl connects with the new Olympias decline (the West decline) that will eventually be extended to the base of the orebody at ‐630 masl of the Flats zone. The FW decline and the West decline extension that is currently being developed, are used as a main haulage route.
A second main decline is currently being driven from -300 masl on the east side of the Flats zone and will be extended down to the -540 masl. This East decline will be used as a haulage ramp for the Flats and lower part of the East Zone.
The existing development above the Flats zone has twin declines and cross-over drifts for personnel and equipment to provide alternate haulage routes and secondary egress. Similamasly, the undeveloped East and West declines will also have multiple cross-over drifts in the Flats zone to provide alternate haulage routes and secondary egress from the bottom of the mine to surface.
16.5.2
Material Handling
Currently all ore and waste is hauled via 40 tonne trucks to surface. A material handling trade-off study was completed in 2018 and determined that, for the 650 ktpa expansion, this continues to be the most cost-effective solution for handling ore and waste.
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16.5.3
Personnel
The mine currently operates 24 hours per day, 7 days per week on 3 shifts per day. One shift is lost per week (Sunday night) for a total of 20 worked shifts per week. Support staff at the mine works only a single shift.
The operation has a full complement of workers at the moment and extra underground workforce will be required (approximately 40 personnel) to operate at the 650,000 tpa proposed.
16.5.4
Underground Mining Fleet
The current mobile equipment list for underground mining is shown in Table 16-3. For the production increase to 650 ktpa, additional equipment will be required, including longhole drills, jumbo, bolters, trucks and loaders. These are also listed in Table 16-3.
Table 16-3: Olympias Mine Equipment List

Equipment | Current | 650 ktpa
Jumbo – 2 boom | 4 | 5
Bolter | 2 | 4
Long hole drill – ITH | 0 | 2
Haul trucks – 40 tonne | 5 | 7
LHD – 14 tonne | 4 | 7
LHD – utility | 2 | 2
Transmixers | 4 | 4
Shotcrete Sprayers | 2 | 2

16.6
Ventilation
16.6.1
Ventilation Strategy
The ventilation design is based on an exhausting system configuration with the main surface fans located at the single exhaust raise. At present, fresh air for Olympias is supplied from the +59 Portal, +70 Portal, Shaft and Fires area. Exhaust is via the main exhaust raise at -173 level via three 132 KW fans. This arrangement enables a total of 150 m³/s to be available for total mine ventilation.
The planned LOM ventilation strategy, with a main exhaust raise an additional exhaust through +70 Portal, will increase the available airflow for the mine to approximately 360 m³/s.
Return air is exhausted from the mine through internal raises adjacent to each ore block before feeding the -173 Level for exhaust into the primary exhaust raise and the ramp leading towards +70 Portal.
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See Figure 16-7 for a representation of the ventilation design for Olympias mine to meet a production rate of 650 ktpa.
Figure 16-7: LOM Ventilation Strategy Planned
Ventilation for each production level is designed such that fresh air will be sourced from level accesses and FW drives. Delivery of fresh air to work areas will be via auxiliary fan and duct, with contaminated air from development and production level activities returning to an internal return air raise.
Two means of egress are provided for each production area of the mine. The primary means of egress is via the ramp system. Secondary emergency egress is provided in internal raises by means of installed ladderways, and also via crossover drifts.
16.6.2
Mine Airflow
In terms of the quantity of air to be supplied for the mine, Part 3, Article 75 of the Greek Mining Regulations states that, “The volume of clean air at all the work places must be at least 5.66 m³ for every first minute and worker and 2.3 m³ for every first minute and internal combustion engine power (0.051 m³/s/kW), regardless the difference in operating time of machinery”. The diesel engine ratings are based on technical specifications for major equipment.
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In order to meet this requirement, it will be necessary for Olympias to move away from the use of high horsepower four-wheel drive vehicles and towards lower horsepower tractors for personnel transport. Furthermore, Olympias is recommended to consider the introduction of all-electric equipment, particulamasly jumbos, which are already tried and proven in other mines.
16.6.3
Emergency Preparedness
In development of the ventilation strategy for Olympias, consideration has been given to the potential for mine emergencies. As such, the following criteria have been established:
●
In general, ramps will be in fresh air once developed.
●
On almost all levels, escape can be either to a ramp or to the escape ladderway in the internal raises.
●
In each ramp, escape may either be up the ramp or down the ramp to a safe area.
●
Portable refuge chambers are installed for flexibility of location at the most appropriate points in the mine.
●
Whilst the primary means of communication will be by radio, a stench system will be in place for introduction of ethyl mercaptan into the portal air stream in the event of fire.
16.7
Drill and Blast
Olympias mine currently uses packaged explosives products for all blasting practices in the underground mine. There is no existing magazine on the property and explosives are delivered to site daily by the supplier. The following sections discuss the opportunity to implement best international practices for: explosives selection, drill and blast designs, and logistics and explosives management at Olympias mine.
16.7.1
Explosives Selection
This section examines the explosives to be utilized for stoping and lateral development at Olympias.
Emulsion is a commonly utilized bulk explosive product that can be used for lateral development and can also be pumped both up and down into stoping blastholes. Emulsion is water resistant, produces low levels of ammonium nitrate in leachate, can be handled in bulk deliveries to reduce costs, has reliable and consistent performance and, depending on the product, can be left in the ground for up to four weeks before firing. Bulk emulsion is proposed as the primary explosive product to be used for both lateral development and stoping (DAF and TLHOS) at Olympias.
ANFO is the most basic bulk explosive product. It is widely available and can be poured or blown into development or DAF blastholes, or used for blind uphole TLHOS. The equipment required to use ANFO is basic and easily maintained. However, the total ammonium nitrate in the leachate from blasting is higher for ANFO than bulk emulsion. If the total ammonium nitrate found in the leachate from explosives detonation is to be minimized, then ANFO should be replaced with emulsion. Further, ANFO is not water resistant, which makes it less versatile than emulsion when wet conditions are encountered. ANFO is also less desirable for blind upholes as it is limited to a blasthole size of 76 mm because of the explosive tending to fall out under gravity. Conversely, bulk emulsion can effectively be used in dry upholes to a maximum diameter of 102 mm. Emulsion is therefore the preferred bulk explosive.
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Packaged explosive products are typically three to five times more expensive on a per unit basis compared to bulk explosive products such as emulsion or ANFO. Although packaged explosives can be used for small narrow blind uphole open stopes (such as in a typical narrow vein mine) they are not commonly used for a bulk mining method such as blind uppers TLHOS, primarily due to inefficiencies in loading large quantities of explosives. Therefore, packaged explosive products have also been eliminated in preference to using bulk emulsion.
It is noted that poor quality / highly fractured ground conditions exist in some DAF areas of the mine and that packaged explosives have been effectively used for blasting in these zones. The switching from primary explosives to bulk emulsion as the main blasting product does not preclude the possibility of maintaining a small inventory of packaged explosive products for use when fractured ground conditions are encountered.
16.7.2
Detonators
Non-electric detonators are the most basic detonation system and are suggested for lateral development and stoping. They are widely available, relatively simple to use, and are less expensive than electronic detonators.
Another viable blast initiation system is electric detonators (or electronic detonators), which have significant advantages over non-electric detonators; these include:
●
Electric detonators are a single product (where the delay is programmable). This reduces the required site stock and increases stock turnover.
●
The detonator delay is programmable. This allows increased blast size, precision blasting and better fragmentation.
●
The delay scatter (prevalent in non-electric detonators) is eliminated.
●
The complexity of the charge process for operators is reduced.
●
Sensitivity of detonators to handling during the charge process is significantly reduced.
●
Multiple security features make them virtually immune to interference, unauthorized use, or unplanned detonation.
Two commonly available electric detonator systems include Orica’s IKON system and Dyno Nobel’s SMARTSHOT system. All electric detonator systems are more complex than non-electric systems and require both operator and engineer training to be used safely and effectively. Training is provided by the system suppliers.
Electric detonators (or electronic detonators) are the highest quality detonating system available but are four to five times more expensive than non-electric systems; they have, therefore, been eliminated as the main detonating system.
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16.7.3
TLHOS Drill and Blast Design
Figure 16-8 and Figure 16-9 show, respectively, a typical longitudinal section and cross section of blind uppers TLHOS of 17 m panel height (floor to stope back), fixed width of 10 m, and length of 26 m. TLHOS stopes account for approximately 31% of the planned stope tonnage for the entire mine. The remaining 69% of planned stope tonnage will be from ore development and DAF.
The overall powder factor for the TLHOS drilling stope design shown in Figure 16-8 is estimated at 0.40 kg/t. Depending on the stope shape and drill pattern, the design powder factor ranges from 0.35 to 0.50 kg/t for all types of longhole open stopes.
Figure 16-8: TLHOS Long Section
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Figure 16-9: TLHOS Cross Section
16.7.4
Lateral Development Blast Design
Figure 16-10 shows the drill layout for a typical lateral development round measuring 5.0 m (H) x 5.0 m (W) x 3.9 m deep.
The overall powder factor for the lateral development design shown is 1.07 kg/t. The powder factor range for all lateral development is from 0.95 to 1.20 kg/t, including DAF stopes and the sill ore drives for the longhole open stopes.
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Figure 16-10: Typical Lateral Development Drill Pattern
16.7.5
Explosives Consumption
At steady-state production of 1,800 tpd (THLOS and DAF) and peak development advance of 12.5 m/day, the monthly mine explosives consumption will be approximately 53 tonnes of bulk emulsion, 12,500 detonators and 12,500 boosters. Table 16-4 shows typical powder factors for THLOS and lateral development (including DAF).
Table 16-4: Powder Factors for Longhole Stoping and Lateral Development/DAF

Description | Ore Tonnes (t) | Explosives (kg) | Powder factor (kg/t)
Longitudinal longhole stope | 13,227 | 5,290 | 0.40
Lateral development (per round) | 340 | 364 | 1.07

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16.8
Underground infrastructure
16.8.1
Service Water
Service water for underground mining is supplied from underground inflows captured in clean-water dams. It is distributed underground via DN100 HDPE lines with pressure regulators installed as required to maintain a pressure below 5.5 bar.
16.8.2
Stockpiles
Ore from the underground mine is transported by truck to the surface crushing plant via the haulage portal, where it is placed into one of four ore sorting bins, depending on its gold grade and point of origin (east versus west ore zones).
16.8.3
Underground Workshop
An area for a new underground workshop has been excavated at the -248 Level. Once equipped, this workshop will be used to maintain any equipment that is captive or rarely leaves the mine. The new shop will be equipped with three heavy equipment maintenance bays as well as a light vehicle service bay, areas for parts storage, office space, hose fitting, machine shop, wash bay, lube bay, tire repair station, first aid station, and restrooms.
Ventilation will be flow-through with a vent raise at the back corner of the shop. As mining progresses, a second workshop or satellite shop or service bay may be required.
16.8.4
Explosives / Cap Magazines
There are currently no magazines at the mine; explosives are delivered to site daily. Capital has been budgeted to develop an underground magazine in 2020.
16.8.5
Fuel Storage
There is currently no underground fuel storage at the mine. Funds have been budgeted to add an underground fuel bay once the new workshop facility is completed in 2020.
16.8.6
Compressed Air
Compressed air demand in the mine is typical for a mine this size. Blasthole loading equipment and shotcrete sprayers have on board compressors; all other mobile and fixed equipment is dependent on the compressed air reticulation system.
Currently an Atlas Copco GA 75 air compressor is set up at the +59 adit. This compressor has a capacity of 12.6 m3/min at 10 bar. For the production increase to 650 ktpa, this will be replaced by an Atlas Copco GA 110 air compressor at the -254 level. The new system will have a capacity of 23.5 m3/min max at 14 bar; however, mine pressure will be maintained below 8 bar. The compressed air is distributed throughout the mine via a network of DN63 HDPE pipes.
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A smaller diesel compressor is also available for emergency situations or power outages.
16.8.7
Electrical Power
There are five underground 20 kV feeds for power underground, three main feeds and two feeds for the pump stations. Of the three main feeds, one runs down the East ramp and two run down the West ramp. All of these feeds are interconnected underground such that the loss of a single cable can be bypassed via a second route. The other two feeds are from the surface pump station to the -95 and -254 pump stations. There is a tie loop between the pump stations.
There is a total of ten main substations located at +46, -85, -173, and -284 in the East ramp and -100, -194, -248, -235, -254 and -260 in the West ramp. The West ramp interconnect is located between the -100 and -194 substations. The west to east interconnect is located between the -284 and -235 substations.
In addition to the main substations, portable 630 kVA substations with 20 kV to 660 V transformers are located where required for mining. All underground power for mobile equipment, fans and pumps is 660 V.
16.8.8
Mine Dewatering
Water inflow has been projected to increase to a maximum of 650 m3/h. Greek regulations require that the pumping capacity must be 2x maximum anticipated flow + standby of rated maximum flow. This means a capacity of 1,300 m3/h and standby pumping capability of 650 m3/h.
The mine dewatering system is rated to meet this requirement and will be expanded to accommodate increased pumping at depth.
There are currently two pumping systems to dewater the mine. The first system is located in the upper East ramp starting at -254 masl and consists of a series of staged pump stations nominally spaced at 30 m vertical intervals. These stations consist of two submersible 37 kW pumps installed in parallel; they discharge into a 100 mm diameter pipe at the lower levels, increasing to 125 mm diameter pipe at higher elevations. The second system consists of two staged systems. These staged systems are located in the West ramp and in the lower East ramp. These systems currently pump water to the old shaft bottom station, where it is stage-pumped to an in-shaft pumping station and finally to surface.
A new pumping station has been excavated and pumps are being installed at the -286 Level. Once complete, it will replace the second system and all water from both the East and West ramps will report to the new pump station, where it will then be pumped via two DN300 steel lines encased in a drill hole (bypassing the existing in-shaft pump station) to the water treatment plant.
In the future, another pump station may be installed similar to the proposed main sump layout at a location near the bottom of the ramps. Any water from below this location would be pumped up the ramp via submersible pumps to the station.
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16.9
Mining Schedule
16.9.1
Stope Cycle Time and Mine Production Rate
Stope cycle times were estimated from an analysis of the various stoping activities. The resulting production rates, recognizing the projected impact of the improvement initiatives, are provided in Table 16-5.
Table 16-5: Olympias Mine, Estimated Stope Productivities

Drift and Fill | Unit | 2019 | 2020 | 2021 | 2022
Number of hours per day | hrs | 12 | 13.5 | 15 | 16.5
Tonnes / day / heading (DAF) | tpd | 109 | 117 | 124 | 140
Time when two faces / heading are available | % | 0 | 25 | 50 | 75
Tonnes / day / heading (DAF) | tpd | 109 | 146 | 186 | 246
Transverse Llonghole (TLHOS) |  |  |  |  |
All-inclusive rate | tpd | 185 | 191 | 197 | 202
Blast / muck / misc. | tpd | 663 | 693 | 721 | 749
Primary / secondary sequence | tpd | 317 | 336 | 353 | 369

The Olympias mine is projected to increase its annual production rate from the current 390,000 tonnes in 2019 up to 650,000 tonnes by 2024. To facilitate this production ramp-up, the mine has budgeted for increased mobile equipment, increased power and an increase in manpower.
The operational improvement plan includes the following elements:
●
Increased equipment availability
●
Upskilling and multiskilling the workforce
●
Increase in manpower utilization
●
Maximizing the use of paste backfill
●
Infrastructural projects to increase underground mining efficiency:
o
Underground workshop
o
Slick line to deliver shotcrete
o
Fuel line to underground
o
Underground slimes filtration system
o
Introduction of remote operations and central control system
These operational improvements will increase the number of hours per day worked at a production face (face time). There are three eight-hour shifts per day and the total face time for all three shifts is indicated as 12 hours per day in 2019, or four hours of face time per shift. Through improved personnel transport, increase in available active headings and time management improvements, it is anticipated that the face time will increase to 5.5 hours per shift or 16.5 hours per day by the year 2022 (Table 16-5). This increase in face time will result in higher stope production rates for both DAF and TLHOS, as indicated in Table 16-5.
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A further key initiative to assist the annual production ramp-up is the optimization of DAF mining. This will be accomplished by initiatives that include the introduction of multi-pass mining as shown in Table 16-4 and Table 16-5. The result will be an increase in the number of concurrently available faces in a DAF stope. The increase in the number of stopes with multiple faces will increase the production rate for individual DAF stopes, in addition to facilitating increased miner face time.
The plan for approximately one-third of total mine production to be provided by the more efficient bulk-mining THLOS system is also a key element of the planned production increase to 650 ktpa.
16.9.2
Development Schedule
The projected LOM development schedule is shown in Table 16-6.
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Table 16-6: Projected Development Schedule over the LOM

Description |  | Total | 2020 | 2021 | 2022 | 2023 | 2024 | 2025 | 2026 | 2027 | 2028 | 2029 | 2030 | 2031 | 2032 | 2033 | 2034 | 2035 | 2036 | 2037 | 2038 | 2039 | 2040
Waste development | km | 4.20 | 0.36 | 0.38 | 0.42 | 0.16 | 0.21 | 0.16 | 0.26 | 0.09 | 0.28 | 0.17 | 0.18 | 0.09 | 0.11 | 0.12 | 0.12 | 0.12 | 0.09 | 0.08 | 0.21 | 0.17 | 0.17
Drift and fill access | km | 6.14 | 1.00 | 1.05 | 0.78 | 0.31 | 0.37 | 0.80 | 0.52 | 0.42 | 0.34 | 0.02 | 0.03 | - | 0.04 | 0.05 | 0.03 | 0.03 | 0.03 | - | - | - | -
West ramp | km | 2.83 | 0.25 | 0.19 | 0.24 | 0.17 | 0.15 | 0.22 | 0.23 | 0.27 | 0.69 | 0.32 | - | - | - | - | - | - | - | 0.01 | - | - | -
Longhole footwall drive | km | 1.64 | 0.05 | - | 0.22 | 0.25 | 0.23 | 0.03 | 0.13 | 0.42 | 0.07 | 0.09 | - | - | - | 0.04 | 0.03 | - | - | 0.04 | 0.03 | - | -
Sump | km | 0.57 | 0.08 | 0.09 | 0.06 | 0.06 | 0.06 | 0.07 | 0.04 | 0.04 | 0.02 | 0.03 | - | - | - | - | 0.01 | - | - | - | - | - | -
Connecting drive | km | 2.77 | 0.23 | 0.27 | 0.11 | 0.46 | 0.21 | 0.42 | 0.36 | 0.48 | 0.06 | - | - | - | 0.07 | 0.05 | 0.01 | - | - | - | - | - | -
Drift and fill footwall drive | km | 4.25 | 0.63 | 0.83 | 0.43 | 0.68 | 0.46 | 0.42 | 0.27 | 0.02 | 0.22 | - | - | - | - | - | 0.06 | - | - | - | - | - | -
Miscellaneous | km | 0.50 | - | - | 0.50 | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | -
Longhole access | km | 5.25 | 0.16 | - | 0.49 | 0.57 | 0.90 | 0.25 | 0.34 | 1.14 | 0.29 | 0.25 | 0.01 | 0.12 | 0.08 | 0.10 | 0.06 | 0.00 | 0.01 | 0.21 | 0.26 | - | -
Ramp access | km | 1.50 | 0.29 | 0.20 | 0.14 | 0.10 | 0.18 | 0.10 | 0.12 | 0.11 | 0.12 | 0.08 | - | - | - | - | - | - | 0.00 | 0.01 | 0.00 | - | -
Remnant ramp | km | 0.32 | 0.13 | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | -
North ramp (not remnants) | km | 0.20 | 0.08 | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | -
East Ramp | km | 2.43 | 0.23 | 0.24 | 0.29 | 0.25 | 0.16 | 0.42 | 0.55 | 0.17 | - | - | - | - | - | - | - | - | - | - | - | - | -
Return air drive | km | 2.81 | 0.30 | 0.31 | 0.30 | 0.08 | 0.25 | 0.36 | 0.35 | 0.36 | 0.25 | - | - | - | - | - | - | - | - | - | - | - | -
Longhole ore drive | km | 9.60 | 0.09 | 0.05 | 0.44 | 1.26 | 0.89 | 0.58 | 0.60 | 0.62 | 0.57 | 0.43 | 0.13 | 0.53 | 0.81 | 0.93 | 0.27 | 0.06 | 0.14 | 0.87 | 0.33 | - | -
Passing bay | km | 0.72 | 0.08 | 0.09 | 0.06 | 0.08 | 0.07 | 0.06 | 0.06 | 0.03 | 0.12 | 0.03 | - | - | - | - | - | - | - | - | - | - | -
Floor bench drives | km | 1.31 | 0.02 | - | 0.03 | - | 0.05 | 0.01 | 0.15 | 0.02 | 0.01 | 0.03 | 0.02 | 0.07 | 0.28 | 0.13 | - | 0.02 | 0.08 | 0.32 | 0.08 | - | -
Recovery drives | km | 0.46 | - | 0.03 | - | 0.06 | - | - | - | 0.03 | - | 0.01 | 0.12 | - | - | 0.02 | - | - | - | 0.02 | - | 0.11 | 0.05
Refuge / lunchroom | km | 0.06 | - | - | - | - | - | 0.02 | 0.03 | 0.02 | - | - | - | - | - | - | - | - | - | - | - | - | -
Remuck | km | 1.05 | 0.12 | 0.11 | 0.09 | 0.13 | 0.11 | 0.06 | 0.06 | 0.12 | 0.09 | 0.06 | - | - | 0.02 | - | 0.02 | - | - | - | - | - | -
Small level sumps | km | 0.07 | 0.01 | 0.01 | 0.00 | - | - | - | 0.02 | 0.01 | - | - | - | - | - | - | - | - | - | - | - | - | -
Electric Bay | km | 0.57 | 0.07 | 0.08 | 0.04 | 0.05 | 0.05 | 0.07 | 0.06 | 0.06 | 0.02 | 0.03 | - | 0.00 | - | - | 0.01 | - | - | - | - | - | -
Service drift (=Exploration drift) | km | 1.07 | 0.20 | 0.67 | 0.09 | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | -
Garage / shops | km | 0.53 | - | - | - | - | - | 0.39 | 0.14 | - | - | - | - | - | - | - | - | - | - | - | - | - | -
Paste fill drift | km | 0.31 | 0.03 | 0.16 | - | - | - | 0.12 | - | - | - | - | - | - | - | - | - | - | - | - | - | - | -
Powder magazine | km | 0.04 | - | 0.00 | 0.04 | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | -
Fuse magazine | km | 0.01 | - | 0.01 | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | -
Fuel bay | km | 0.06 | - | 0.02 | - | - | - | 0.01 | 0.02 | 0.02 | - | - | - | - | - | - | - | - | - | - | - | - | -
Drill drift for flats | km | 0.38 | 0.32 | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | - | -
Main sump | km | 0.76 | - | - | - | - | 0.23 | 0.03 | 0.23 | 0.02 | 0.13 | 0.12 | - | - | - | - | - | - | - | - | - | - | -
Total | km | 52.37 | 4.72 | 4.78 | 4.75 | 4.67 | 4.58 | 4.60 | 4.54 | 4.44 | 3.29 | 1.67 | 0.50 | 0.81 | 1.41 | 1.42 | 0.63 | 0.22 | 0.36 | 1.55 | 0.91 | 0.28 | 0.22

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16.9.3
Underground Mining Production Plan
Table 16-7 shows the scheduled production tonnes and grade for the Olympias mine life. Recovery and dilution factors have been discussed in Section 15. The total projected LOM is 21 years.
Table 16-7: Olympias LOM Ore Production Schedule

Yrs | Total Tonnes (t) | East Tonnes (t) | West Tonnes (t) | Flats Tonnes (t) | Remnants Tonnes (t) | Au (g/t) | Ag (g/t) | Pb (%) | Zn (%)
2020 | 415,000 | 77,146 | 233,606 | 19,069 | 85,179 | 7.22 | 100.28 | 3.3 | 4.0
2021 | 440,000 | 114,098 | 221,126 | 55,143 | 49,633 | 7.72 | 98.23 | 3.1 | 3.5
2022 | 500,000 | 100,000 | 305,888 | 44,112 | 50,000 | 7.65 | 126.95 | 4.0 | 3.9
2023 | 650,000 | 129,893 | 256,993 | 213,113 | 50,000 | 7.80 | 123.91 | 4.0 | 4.3
2024 | 650,000 | 129,815 | 167,110 | 303,076 | 50,000 | 7.31 | 140.78 | 4.7 | 4.9
2025 | 650,000 | 130,000 | 14,378 | 455,622 | 50,000 | 7.70 | 127.46 | 4.2 | 5.1
2026 | 650,000 | 129,611 | 10,223 | 430,291 | 79,874 | 7.35 | 149.86 | 4.9 | 5.3
2027 | 650,000 | 129,812 | 10,286 | 429,979 | 79,924 | 6.49 | 104.79 | 3.5 | 4.8
2028 | 650,000 | 129,992 | 11,608 | 428,433 | 79,967 | 6.77 | 112.67 | 3.5 | 5.1
2029 | 650,000 | 118,568 | 4,696 | 446,736 | 80,000 | 6.63 | 105.50 | 3.7 | 5.2
2030 | 650,000 | 100,144 | - | 469,917 | 79,939 | 7.19 | 125.41 | 4.3 | 6.1
2031 | 650,000 | 129,737 | - | 420,513 | 99,750 | 7.01 | 133.24 | 4.7 | 6.0
2032 | 650,000 | 80,991 | - | 469,373 | 99,636 | 7.29 | 112.50 | 3.8 | 5.4
2033 | 650,000 | 98,025 | - | 451,975 | 100,000 | 7.49 | 113.09 | 4.0 | 5.9
2034 | 650,000 | - | - | 550,000 | 100,000 | 7.26 | 131.22 | 4.5 | 5.5
2035 | 650,000 | - | - | 550,000 | 100,000 | 7.06 | 138.45 | 4.9 | 6.1
2036 | 650,000 | - | - | 500,000 | 150,000 | 6.99 | 126.66 | 4.6 | 5.6
2037 | 650,000 | - | - | 500,000 | 150,000 | 7.13 | 100.13 | 3.8 | 4.9
2038 | 650,000 | - | - | 468,563 | 181,437 | 7.89 | 103.14 | 3.8 | 5.2
2039 | 650,000 | - | - | 650,000 | - | 4.18 | 91.33 | 3.3 | 5.9
2040 | 550,000 | 230,937 | - | 319,063 | - | 5.11 | 103.60 | 3.6 | 5.5
Total | 12,925,000 | 1,844,094 | 1,284,779 | 8,207,886 | 1,715,338 | 7.02 | 119 | 4.1 | 5.3

Note: Totals may not compute exactly due to rounding, 2019 not shown, but contained 2 months of reserve production
Figure 16-11 shows the LOM production profile of ore tonnes from each zone.
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Figure 16-11: LOM Production Profile
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SECTION ● 17
Recovery Methods
17.1
Introduction
A three-phase plan has been adopted for Olympias Project.
Phase I – A new bulk flotation plant was built and commissioned to process the old tailings from 2012 to 2016. A single gold/pyrite/arsenopyrite concentrate was produced.
Phase II – A new selective flotation plant was built and commissioned in May 2017 to process fresh ore from the underground mine at a throughput of around 400 ktpa. Three concentrates are produced, namely, lead/silver concentrate, zinc concentrate, and gold/pyrite/arsenopyrite concentrate. Lead/silver (lead) and gold/pyrite/arsenopyrite (gold) concentrates are sold through Thessaloniki port and zinc concentrate is sold through Stratoni port. A portion of flotation tailing is used for underground backfill. Any surplus of flotation tailings is filtered and then trucked to Kokkinolakkas TMF. The Kokkinolakkas TMF is 8.5 km south-east of Olympias mine.
Phase III – Expansion and upgrade will take place to the underground mine, Olympias process plant and Stratoni port to reach production rate of 650 ktpa of ore. AS indicated previously, a permit is required to increase throughput of the process plant up to 650 ktpa.
17.2
Process Plant Design Criteria
A summary of major process design criteria for Phase II is provided in Table 17-1.
Table 17-1: Major Process Design Criteria for Phase II

Description | Unit | Parameter
1.0 Ore characteristics
Ore Feed Grade - Gold, Silver | g/t | 9.5, 124
Ore Feed Grade - Lead, Zinc, Arsenic, Sulphur | % | 4.2, 5.7, 5.0, 17.5
Ore Specific Gravity |  | 3.39
Crushing Work Index | kW.h/t | 4.2 - 7.7
Rod Mill Work Index - Range, Design | kW.h/t | (7.6 - 12.1), 11.2
Ball Mill Work Index - Range, Design | kW.h/t | (11.4 - 14.3), 13.8
Abrasion Work Index - Range, Design |  | (0.244 - 0.323), 0.28
Ore Moisture Content | % | 5
2.0 Production schedule
Annual Throughput | tpa | 400,000
Operating Days Per Year |  | 350
Lead Concentrate Grade: Pb | % | 63
Lead Concentrate Recoveries: Pb, Ag | % | 85, 78
Zinc Concentrate Zn Grade & Recovery | % | 52, 91
Gold-Arsenic-Pyrite Conc Gold Recovery | % | 86
Gold-Arsenic-Pyrite Conc Sulfur Grade | % | 38
3.0 Crushing & fine ore storage
Crusher Operating Schedule; Shifts/d, h/shift |  | 2, 8
Plant Operating Time | % | 75
Required Daily Production | t | 1,272
Primary Crusher Feeder Type |  | Vibrating Grizzly
Primary Crusher Type |  | Single Toggle Jaw
Crusher Product Size, P80 | mm | 90
Secondary Crusher Feed Screen Aperture – upper / lower deck | mm | 50/20
Secondary Crusher Type |  | Standard Cone
Product Size (P80) | mm | 28
Screen Aperture | mm | 13
Tertiary Crusher Type |  | Short Head Cone
Product Size (P80) | mm | 13
Fine Ore Bin | t | 1,155
Fine Ore Reclaim System | Type | Belt Feeder
Number |  | 5
Maximum Capacity | t/h | 125
4.0 Grinding & classification
Grinding Circuit Type |  | Single stage ball milling
Feed Size F80 | mm | 10
Operating Schedule |  |
Throughput | t/h | 53
Shifts/d, h/shift |  | 3, 8
Plant Availability | % | 90
Plant Operating Time | % | 86
Primary Grinding Mill Type |  | Grate Discharge Ball Mill
Installed Motor Power | kW | 900 (VS)
Classification Type |  | Hydrocyclone
Overflow Density | % solid | 41
Overflow Product Size (P80) | µm | 120
Circulating Load | % | 250
Flash Flotation Cell Type |  | Dual Outlet SkimAir SK240
Proportion of Cyclone U/F to Flash | % | 95
Residence Time | min | 3.2
5.0 Lead flotation, regrind & tails thickening circuits
Lead Flotation Feed Conditioner time | min | 5.5
Lead Rougher Flotation Retention Time | min | 4.6
Rougher Concentrate Mass Pull | % | 5.7
Lead Scavenger Flotation Retention Time | min | 20.4
Scavenger Concentrate Mass Pull | % | 6.3
Lead Regrind Classification Type |  | Hydrocyclone
Overflow Density | % solid | 2.5
Overflow Product Size (P80) | µm | 14
Lead Regrind Mill Type/Size/Model |  | IsaMill, M500
Installed Motor Power | kW | 200
Product Size P80 | µm | 14
Lead 1st Cleaner Feed Conditioning time | min | 12.4
Lead 1st Cleaner Flotation Retention Time | min | 8.9
Concentrate Mass Pull | % | 22
Lead 1st Cleaner Scav Flotation Retention Time | min | 6
Concentrate Mass Pull | % | 5
Lead 2nd Cleaner Flotation Retention Time | min | 12.2
Concentrate Mass Pull | % | 16
Lead 3rd Cleaner Flotation Retention Time | min | 16.4
Concentrate Mass Pull | % | 6
Lead Tailings Thickeners No, Dia | No., m | 2, 8
Underflow Density | % solid | 60
6.0 Zinc flotation & regrind circuits
Zinc Flotation Feed Conditioner No 1 | min | 13.5
Zinc Flotation Feed Conditioner No 2 | min | 13.5
Zinc Rougher Flotation Retention Time | min | 9.4
Concentrate Mass Pull | % | 11
Zinc Scavenger Flotation Retention Time | min | 21.7
Concentrate Mass Pull | % | 9
Zinc Regrind Classification Type |  | Hydrocyclone
Underflow Density | % solid | 53
Overflow Product Size (P80) | µm | 15
Zinc Regrind Mill; Type, Size / Model |  | IsaMill, M500
Installed Motor Power | kW | 200
Product Size P80 | µm | 15
Zinc 1st Cleaner Feed Conditioning time | min | 9.3
Zinc 1st Cleaner Flotation Retention Time | min | 8.1
Concentrate Mass Pull | % | 20
Zinc 1st Cleaner Scav Flotation Retention Time | min | 7.6
Concentrate Mass Pull | % | 7
Zinc 2nd Cleaner Flotation Retention Time | min | 12.9
Concentrate Mass Pull | % | 16
Zinc 3rd Cleaner Flotation Retention Time | min | 15.4
Concentrate Mass Pull | % | 10
7.0 Gold-pyrite flotation
Au/Pyrite Flotation Feed Conditioner No 1 | min | 8.6
Au/Pyrite Flotation Feed Conditioner No 2 | min | 8.6
Au/Pyrite Rougher Flotation Retention Time | min | 6.8
Concentrate Mass Pull | % | 15
Au/Pyrite Scavenger Flotation Retention Time | min | 23.4
Concentrate Mass Pull | % | 49
Au/Pyrite Cleaner Feed Conditioning | min | 7.6
Au/Pyrite 1st Cleaner Flotation Retention Time | min | 17
Concentrate Mass Pull | % | 15
8.0 Concentrate handling
Lead Concentrate |  |
Lead Concentrate Thickener |  | 8 m dia.
Underflow Density | % solid | 65
Lead Concentrate Surge Tank Capacity | h | 23
Lead Concentrate Filtration, Filter Type |  | Pressure - Plate & Frame
Filter Area, Total | m² | 26
Residual Cake Moisture | % | 9.9
Zinc Concentrate |  |
Zinc Concentrate Thickener |  | 8 m dia.
Underflow Density | % solid | 65
Zinc Concentrate Surge Tank Capacity | h | 22
Zinc Concentrate Filtration, Filter Type |  | Pressure Plate & Frame
Filter Area, Total | m² | 49
Residual Cake Moisture | % | 11
Au/Pyrite Concentrate |  |
Au/Pyrite Concentrate Thickener |  | 12 m dia.
Underflow Density | % solid | 65
Au/Pyrite Concentrate Surge Tank | h | 19.7
Au/Pyrite Concentrate Filtration, Filter Type |  | 2 x Pressure Plate / Frame
Filter Area, Total | m² | 620
Residual Cake Moisture | % | 13
9.0 Tailings disposal
Flotation Tailings Thickener Type |  | 10 m dia.
Underflow Density | % solid | 65
Tailings Filter Surge Tank Capacity | h | 5.3
Tailings Filtration Filter Type |  | Plate and Frame
Filter Area, Total | m² | 620
Residual Cake Moisture | % | 13
Tailings Disposal Type (When not backfilling) |  | Dry Stacked Tailings

17.3
Process Flowsheet
A simplified process flowsheet is shown in Figure 17-1.
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Source: Eldorado
Figure 17-1: Simplified Process Flowsheet of Olympias Plant
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17.4
Process Plant Description
The current process plant is currently able to treat 400,000 tonnes per annum at 53 t/h. A few key features are as follows.
●
Three-stage crushing
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Single-stage ball milling in closed circuit with hydrocyclones
●
Nearly all hydrocyclones underflow is fed to flash flotation.
●
Lead flotation consists of rougher, scavenger, regrinding, three stages of cleaning
●
Zinc flotation consists of rougher, scavenger, regrinding, three stages of cleaning
●
Gold-Pyrite flotation utilizes rougher, scavenger, and single stage of cleaning
●
Concentrate thickening, filtration, and packaging
●
Tailings thickening and filtration
●
Tailings paste backfill
●
Reagent mixing, storage and distribution
●
Water and air services
The following sections describe in detail the above processes.
17.4.1
Crushing, Ore Storage, and Reclaim
A ROM ore stockpile is situated adjacent to the primary crusher feed bin. The ROM stockpile is divided into a few areas according to ore type and grades to facilitate blending. Ore is reclaimed from the ROM stockpile by a front-end loader (FEL).
The crushing plant is operated 350 days per year, 16 hours per day at a crushing rate of 125 t/h.
The crushing product size is 100% passing 13 mm and 80% passing 10 mm.
A 600 mm x 800 mm inclined static grizzly is fitted to the primary crusher feed hopper for protection from oversize rocks. The oversize rocks are periodically broken to less than 600 mm with a mobile rock breaker. The primary crusher feed hopper is equipped with dust suppression spray.
The primary crusher is fed by a vibrating grizzly feeder that scalps the rocks greater than 75 mm ahead of crushing. The undersize material is directed onto the crusher discharge conveyor. Grizzly feeder oversize feeds a jaw crusher (Sandvik CJ411) at closed side setting (CSS) of 75 mm.
The jaw crusher product (100% passing 180 mm and 80% passing 90 mm) reports to the crusher discharge conveyor and is delivered, along with the fines from the vibrating grizzly feeder, to the primary crushed coarse ore stockpile feed conveyor. The coarse ore stockpile feed conveyor discharges onto the coarse ore stockpile with a live capacity of 79 tonnes. A cross-belt tramp metal magnet is positioned over the coarse ore stockpile feed conveyor. Tramp metal is discharged via a chute into a tramp metal bin located at ground level.
Primary crushed ore is withdrawn from the coarse ore stockpile by a vibrating feeder and transported to the secondary crusher feed screen conveyor. A metal detector is installed on the secondary crusher feed screen conveyor to ensure all tramp metal is removed.
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The primary crushed ore then reports to a double-deck scalping screen with 50 mm aperture for the top deck and 20 mm aperture for the bottom deck. The oversize material from both decks is directed to the secondary crusher (Sandvik CH430 with CSS at 28 mm) whilst the screen undersize and the secondary crusher product are deposited onto the secondary crusher discharge conveyor, which in turn discharges onto the tertiary crusher feed screen conveyor.
Ore on the secondary crusher discharge conveyor transfers onto the tertiary crusher feed screen conveyor, and then to the tertiary product screen. A metal detector is installed on the tertiary feed screen conveyor to ensure all tramp metal is removed for protection of the tertiary crusher. When the presence of metal is detected by the metal detector, the crushing circuit automatically stops to allow the operator to pick the metal from the belt at a pre-determined location. The product screen is a single deck vibrating screen with an aperture of 13 mm. The screen oversize returns to the tertiary crusher.
The tertiary crusher feed hopper is fitted with a belt feeder. The tertiary crusher (Sandvik CH430 at CSS of 13 mm) is a short head cone crusher that operates in closed circuit with the product screen. The crusher discharges directly onto the tertiary crusher discharge conveyor, which in turn discharges onto the secondary crusher discharge conveyor (thus operating in closed circuit with the product screen).
The final crusher product after screening has a P80 of 10 mm and is deposited onto the tertiary screen undersize conveyor, which in turn discharges onto the fine ore bin feed conveyor. The fine ore bin feed conveyor transfers the final crusher product to the fine ore bin (FOB). The crushing rate is monitored by a weightometer located on the fine ore bin feed conveyor. The FOB has a live capacity of 1,155 tonnes to provide 21.8 hours of mill feed.
Mill feed is withdrawn from the FOB via five variable speed belt feeders. A weightometer indicates the instantaneous and totalized mill feed tonnage for control of belt feeder speed. Mill feed from the FOB is transported to the grinding circuit by the ball mill feed conveyor.
17.4.2
Grinding, Classification, and Flash Flotation
The grinding circuit consists of a single-stage overflow ball mill that operates in closed circuit with hydro-cyclones to produce a ground product slurry with P80 of 120 µm and with a flash flotation cell for recovery of galena from the cyclone underflow.
Mill feed enters the grinding circuit through the mill feed chute, where process water is added. The ball mill is 3.65 m diameter by 4.00 m long effective grinding length and powered by a 900 kW variable speed drive. Normal operation requires 700 kW power draw. The ball mill is fitted with a trommel screen. Trommel screen undersize slurry reports to the cyclone feed hopper.
Reagents are added to the cyclone feed hopper for depression of sphalerite, arsenopyrite and pyrite.
Cyclone overflow flows by gravity to the trash screen feed box. A single horizontal vibrating trash screen (0.8 mm x 12 mm slotted aperture) removes trash from the flotation feed. Oversize trash gravitates directly to a trash bin. Screen undersize gravitates to the lead rougher flotation conditioning tank.
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Cyclone underflow gravitates to the underflow distribution box where the flow is split. Up to 100% of cyclone underflow stream can be directed to the flash flotation circuit to recover fast-floating galena. Dilution water and flotation reagents are added. Concentrate flows by gravity to the flash flotation concentrate transfer pump before being pumped to the inline stream analyzer (ISA) for analysis and then regrinding. The flash flotation coarse tailings flows by gravity to the ball mill for further grinding. The fine tailings from the top outlet flows to the cyclone feed hopper or to a surge tank and then to conditioning tank ahead of lead rougher flotation
17.4.3
Lead Flotation and Regrind
The lead flotation circuit consists of rougher, scavenger, and three stages of cleaner. A regrind circuit provides size reduction of the rougher, scavenger, flash flotation and cleaner scavenger concentrates to improve mineral liberation and clean particle surfaces.
Reagents are added to the lead flotation conditioning tank, lead rougher and first lead scavenger. Slurry flows by gravity from the conditioning tank to the first lead rougher flotation cell. Flotation feed (cyclone overflow) is automatically sampled for analysis by the ISA to provide a flotation feed assay and feed slurry density.
The lead rougher/scavenger flotation circuit consists of one 10 m³ rougher tank cell followed by four 10 m³ capacity scavenger flotation tank cells
The rougher concentrate flows by gravity to the lead rougher concentrate pump and the scavenger concentrate is directed to the lead scavenger concentrate pump. The lead scavenger flotation tail discharges via a dart valve to the lead scavenger tail hopper. The hopper is serviced by duty and standby variable speed pumps that pump lead scavenger flotation tail to the lead tails thickeners. An automatic sampler on the lead scavenger tail provides a sample to the ISA.
The flash flotation, lead rougher, and lead scavenger concentrates are pumped separately to the lead regrind cyclone feed hopper, where they are combined with the lead cleaner scavenger concentrate and reagents for sphalerite, arsenopyrite and pyrite depression. The combined concentrate streams are classified in the regrind cyclone cluster consisting of three 150 mm cyclones (two operating and one standby). The regrind cyclones are fed at a pulp density of approximately 18% solids. The cyclone overflow flows by gravity to the lead cleaner conditioning tank.
The cyclone underflow gravitates to the lead regrind mill feed hopper and is diluted to a density of 50% solids before being pumped to the lead regrind mill (IsaMill M500). Media addition is based on mill motor power draw. The regrind mill is configured in open circuit. Pulp discharge from the lead regrind mill joins the cyclone overflow in the lead cleaner conditioning tank. The regrind mill circuit can also be configured to enable a portion of the regrind mill discharge to be recycled to the feed hopper to assist in maintaining a constant volumetric flow rate. The lead cleaner conditioning tank has an effective volume of 12.6 m³. Reagents are added to the conditioning tank and various points in the lead cleaner circuit.
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Four 4.25 m³ Outotec OK-3HG-U flotation cells operate as a single bank for the 1st cleaner. The lead 1st cleaner tail flows by gravity via a dart valve to the lead cleaner scavenger cells (two 4.25 m³). Concentrate from the lead cleaner scavenger cells is pumped to the lead regrind mill cyclone feed hopper. The lead cleaner scavenger tail discharges via a dart valve to the lead cleaner scavenger tail hopper, and then pumped to the lead flotation tailing thickeners.
Concentrate from the lead 1st cleaner cells is pumped to the lead 2nd cleaner cells. Three 4.25 m³ capacity Outotec OK-3HG-U flotation cells are arranged in a single bank. Concentrate from the lead 2nd cleaner cells is pumped to the lead 3rd cleaner cell feed box. Two 4.25 m³ capacity Outotec OK-3HG-U flotation cells operate as a single bank. The lead 3rd cleaner tail is combined with the lead 1st cleaner concentrate to the lead 2nd cleaner circuit. The lead 3rd cleaner concentrate is final concentrate and reports to the lead concentrate thickener.
An ISA is installed in the plant to provide analysis of five key process streams in the plant. These are:
●
Flash flotation concentrate
●
Cyclone overflow
●
Lead scavenger tail
●
Lead cleaner scavenger tail
●
Lead final concentrate
Reagent addition rates are controlled including frother (MIBC), collector (Aerophine 3418A) and depressants (sodium cyanide, zinc sulphate and lime)
17.4.4
Lead Flotation Tails Thickening
The lead scavenger tails and lead cleaner scavenger tails are pumped to a distribution box and then into two 8 m diameter tailings thickeners. The lead tails are thickened to 60% solids, then pumped to zinc flotation conditioning tank 1. Flocculant is added to the lead tailings thickener feed stream to assist in the thickening process. Lead tailings thickener overflow reports to the lead thickener overflow tank and is reused as process water in the grinding and lead flotation circuits.
17.4.5
Zinc Flotation and Regrind
The zinc flotation circuit consists of rougher, scavenger and three stages of cleaner. A regrind circuit provides particle size reduction of the rougher, scavenger and cleaner scavenger concentrates to improve mineral liberation and clean particle surfaces.
The thickened lead tailings are diluted to 35% solids by the addition of process water in the first conditioning tank.
The two agitated zinc flotation conditioning tanks have a total effective volume of 57 m³. Reagents are added to the first and second conditioning tanks, first rougher and scavenger cell feed boxes.
The zinc rougher/scavenger flotation circuit consists of two 10 m³ rougher tank cells followed by four 10 m³ scavenger tank cells.
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The rougher concentrate flows by gravity to the zinc rougher concentrate pump and the scavenger concentrate is directed to the zinc scavenger concentrate pump. The zinc scavenger flotation tail discharges via a dart valve to the zinc scavenger tail hopper and is pumped to the first gold-pyrite flotation feed conditioning tank. A sampler on the zinc scavenger tail stream provides a sample to the ISA.
The zinc rougher and scavenger concentrates are pumped separately to the zinc regrind cyclone feed hopper, where they are combined with the zinc cleaner scavenger concentrate and reagents for sphalerite activation and depression of arsenopyrite and pyrite. The combined concentrate streams are classified in a cyclone cluster consisting of four 150 mm cyclones (three operating and one standby). The regrind cyclones are fed at a pulp density of approximately 20% solids. The cyclone overflow flows by gravity to the zinc cleaner conditioning tank.
The cyclone underflow gravitates to the zinc regrind mill feed hopper and is diluted to a density of 50% solids before being reground. The zinc regrind mill is an IsaMill M500 fitted with a 200 kW motor. Media addition is based on mill motor power draw. The regrind mill is configured in open circuit. Pulp discharge from the regrind mill joins the overflow from the regrind cyclone overflow in the zinc cleaner conditioning tank. The regrind mill circuit can also be configured to enable a portion of the regrind mill discharge to be recycled to the feed hopper to assist in maintaining a constant volumetric flow rate.
The zinc cleaner conditioning tank has an effective volume of 10 m³. Reagents are to the conditioning tank and various points in the zinc cleaner circuit.
Four 4.25 m³ Outotec OK-3HG-U flotation cells operate as a single bank in the 1st cleaner
The 1st cleaner tail gravitates via a dart valve to the zinc cleaner scavenger cells (three 4.25 m³ Outotec OK-3HG-U flotation cells). Concentrate from the cleaner scavenger cells is pumped back to the zinc regrind cyclone feed hopper. The zinc cleaner scavenger tail discharges via a dart valve to the zinc cleaner scavenger tail hopper and then pumped to the first gold-pyrite flotation conditioning tank. A sampler on the tail provides a sample to the ISA.
Concentrate from the 1st cleaner cells is pumped to the 2nd bank of cleaner cells (three 4.25 m³ Outotec OK-3HG-U cells).
Concentrate from the 2nd cleaner cells is pumped to the 3rd cleaner cells (two 4.25 m³ Outotec OK-3HG-U cells). The 3rd cleaner tail stream is combined with the 1st cleaner concentrate as the 2nd cleaner feed. The zinc 3rd cleaner concentrate as final concentrate reports to the zinc concentrate thickener feed hopper.
The ISA provides analysis of three key process streams in the plant.
●
Zinc scavenger tail
●
Zinc cleaner scavenger tail
●
Zinc final concentrate
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Reagent addition rates are controlled, including frother (MIBC), collector (sodium isopropyl xanthate), activator (copper sulphate), and depressant (lime).
17.4.6
Gold-Arsenopyrite-Pyrite Flotation
The gold-arsenopyrite-pyrite flotation circuit consists of rougher, scavenger, and a single stage of cleaner. The cleaner tail is recirculated to the scavenger feed. The scavenger tail reports as the final tail, and the cleaner concentrate joins the rougher concentrate as final concentrate.
The two agitated gold flotation conditioning tanks have a total effective volume of 56 m³. Slurry gravitates from the first conditioning tank to the second conditioning tank. Reagents are added to the first and second conditioning tanks and to the first gold rougher and scavenger cell feed boxes.
The gold rougher/scavenger flotation circuit consists of one 20 m³ rougher tank cell followed by five 20 m³ capacity scavenger tank cells arranged to produce separate concentrates.
The first rougher and the first and second gold scavenger cells are each equipped with a feed box. The gold flotation tail discharges via a dart valve to the gold flotation tail hopper, and then pumped to the tailings thickener. A sampler on the discharge side of the gold flotation tail pump provides a sample to the ISA.
Concentrate from the gold rougher flotation cells is pumped directly to the final gold concentrate thickener or alternatively to the gold cleaner cells. The concentrate produced from the gold scavenger flotation cells is pumped to the gold cleaner conditioning tank.
The gold cleaner flotation conditioning tank has an effective volume of 12.6 m³. Reagents are added to the conditioning tank and to the first gold cleaner cell feed box.
The 1st cleaner flotation circuit consists of a single 10 m³ tank cell followed by a bank of two 10 m³ tank flotation cells.
The 1st cleaner tail discharges via a dart valve to the gold 1st cleaner flotation tail hopper and then pumped to the scavenger bank, where it is combined with the rougher flotation tail.
The 1st cleaner concentrate gravitates to the final gold concentrate hopper, where it combines with the rougher concentrate, and then pumped to the gold concentrate thickener. A sampler on the discharge side of the gold concentrate transfer pump provides a sample to the ISA.
The ISA provides analysis of two key process streams in the plant.
●
Gold scavenger tail
●
Gold final concentrate
Reagent addition rates are controlled, including frother (MIBC), collector (sodium isopropyl xanthate) and activator (sulphuric acid)
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17.4.7
Lead Concentrate Thickening and Filtration
Concentrate from the lead flotation circuit is pumped to the lead concentrate thickener feed hopper, and then to a de-aerator, from where it flows by gravity to the feedwell of an 8 m diameter concentrate thickener. Overflow from the de-aerator gravitates back to the lead concentrate thickener feed hopper with the level allowed to vary between a high and a low limit to maintain the de-aerator operating pressure set-point.
An auto-dilution system on the feedwell of the thickener increases the settling rate. Flocculant is added to the feedwell. The design thickener underflow density is 65% solids.
A bed level measuring device is installed to monitor the thickener bed depth. The addition rate of flocculant is controlled according to the bed depth. The thickener is equipped with two variable speed peristaltic underflow pumps arranged in a duty / standby configuration, which are controlled to maintain thickener bed pressure. A nuclear density gauge measures the density of the thickened slurry to the lead concentrate filter storage tank.
Thickener torque is automatically maintained in pre-set ranges via the thickener local control panel, which raises and lowers the thickener rakes according to torque readings. Torque readings and rake status (i.e. running / stopped / fault) are displayed on the control system. Lead concentrate thickener overflow gravitates to the lead thickener overflow tank. Combined lead concentrate thickener overflow and lead tailings thickener overflow is reused in the grinding and lead flotation circuits.
The lead concentrate filtration section consists of an agitated lead concentrate filter feed storage tank, dual filter feed pumps and a horizontal plate/frame pressure filter. The lead concentrate tank has a working volume of 49 m³, equivalent to 23 hours of concentrate production
The thickened concentrate slurry is pumped from the lead concentrate tank to a pressure filter (26 m2 filtration area) for dewatering. The pressure filter produces a filter cake containing 9 – 10% moisture.
The filter cake discharges onto the lead filter cake conveyor, which in turn transfers it to a bagging station. Bagged lead concentrate is weighed and loaded onto haulage trucks.
A filtrate air separator is used to remove air from the filtrate before it reports to the filtrate hopper. During normal operation, the filtrate is collected in the lead filter filtrate tank. Filtrate is then pumped to the concentrate thickener to remove fine solids prior to re-use in the plant. The filter has a local control panel that controls the operation of the filter and associated valves and pumps, including the feed pumps and the cloth wash water pump.
17.4.8
Zinc Concentrate Thickening and Filtration
Concentrate from the zinc flotation circuit is pumped to the zinc concentrate thickener feed hopper and then to a de-aerator, from where it gravitates to the feedwell of an 8 m diameter concentrate thickener. The de-aerator feed pump speed is controlled automatically to maintain a set pressure in the de-aerator. Overflow from the de-aerator gravitates back to the zinc concentrate thickener feed hopper, with the level allowed to vary between a high and a low limit while maintaining the de-aerator operating pressure set-point.
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An auto-dilution system on the feedwell of the thickener increases the settling rate, and flocculant is added to the feedwell. The design thickener underflow density is 65% solids.
A bed level measuring device is installed to monitor the thickener bed depth. The addition rate of flocculant is controlled according to the bed depth. The thickener is equipped with two variable speed peristaltic underflow pumps arranged in a duty / standby configuration, which are controlled to maintain thickener bed pressure. A nuclear density gauge measures the density of the thickened slurry to the zinc concentrate filter storage tank.
Thickener torque is automatically maintained in pre-set ranges via the thickener local control panel, which raises and lowers the thickener rakes according to torque readings. Torque readings and rake status (i.e. running / stopped / fault) are displayed on the control system. Zinc concentrate thickener overflow gravitates to the water collection sump before being pumped directly to the process water tank.
The zinc concentrate filtration section consists of a zinc concentrate filter storage tank, dual filter feed pumps and a horizontal plate/frame pressure filter. The zinc concentrate tank has a working volume of 91 m³, equivalent to 22 hours of concentrate production. The tank has an ultrasonic level device to monitor levels.
The thickened concentrate slurry is pumped from the zinc concentrate tank to a pressure filter (48 m2 filtration area) for dewatering. The pressure filter produces a filter cake containing 12% moisture.
The filter cake discharges onto the zinc filter cake conveyor, which in turn transfers the filtered concentrate to a bagging station. Bagged zinc concentrate is weighed and loaded onto haulage trucks.
A filtrate air separator is used to remove air from the filtrate produced during the core blow cycle, and the filtrate is collected in the zinc filter filtrate tank before being pumped to the zinc concentrate thickener.
The filter has a local control panel that controls the operation of the filter and associated valves and pumps, including the feed pumps and the cloth wash water pump.
17.4.9
Gold Concentrate Thickening and Filtration
Concentrate from the gold-pyrite flotation circuit is pumped to the gold concentrate thickener stock hopper and then to a de-aerator, from where it gravitates to the feedwell of a 12 m diameter thickener. The de-aerator feed pump speed is controlled automatically to maintain a set pressure in the de-aerator. Overflow from the de-aerator gravitates back to the gold concentrate thickener feed hopper.
An auto-dilution system on the feedwell of the thickener increases the settling rate, and flocculant is added to the feedwell. The design thickener underflow density is 65% solids.
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A bed level measuring device is installed to monitor the thickener bed depth. The addition rate of flocculant is controlled according to the bed depth. The thickener is equipped with two variable speed peristaltic underflow pumps arranged in a duty / standby configuration, which are controlled to maintain thickener bed pressure. A nuclear density gauge measures the density of the thickened slurry.
Thickener torque is automatically maintained in pre-set ranges via the thickener local control panel, which raises and lowers the thickener rakes according to torque readings. Torque readings and rake status (i.e. running / stopped / fault) are displayed on the control system. Gold concentrate thickener overflow flows by gravity to the gold thickener overflow tank before being pumped to the process water tank.
The gold concentrate filtration section consists of two storage tanks, two sets of dual filter feed pumps and two horizontal plate/frame pressure filters. The gold concentrate tanks have a working volume of 119 m³, equivalent to 19.7 hours of concentrate production.
The thickened concentrate slurry is pumped to two pressure filters in parallel. Total filtration area is 1,240 m². The filter cake contains 10% moisture.
The filter cake discharges onto dedicated conveyors, which in turn transfer the filter cake from each filter to a bagging station. Bagged gold concentrate is weighed and loaded onto haulage trucks.
A filtrate air separator is used to remove air from the filtrate before it reports to the filtrate tank. During normal operation, the filtrate from both filters is collected in the gold filter filtrate tank before being pumped to the gold concentrate thickener.
17.4.10
Tailings Thickening and Filtration
Flotation tailings from the gold scavenger flotation are pumped to the 10 m diameter high rate thickener. Flocculant is added to increase the settling rate and underflow density to approximately 65% solids w/w. Tailings thickener overflow gravitates directly to the zinc thickener overflow tank for re-use in the grinding and flotation circuits.
A bed level measuring device is installed to monitor the thickener bed depth. The addition rate of flocculant is controlled according to the bed depth. The thickener is equipped with two variable speed peristaltic underflow pumps arranged in a duty / standby configuration, which are controlled to maintain thickener bed pressure.
Thickener torque is automatically maintained in pre-set ranges via the thickener local control panel, which raises and lowers the thickener rakes according to torque readings. Torque readings and rake status (i.e. running / stopped / fault) are displayed on the control system.
Tailings thickener underflow flows by gravity to the thickened tailings hopper and is then pumped to the tailings filtration or paste backfill plant.
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The tailings filtration section consists of an agitated tailings holding tank, filter feed pumps and a horizontal plate/frame pressure filter. The tailings storage tank has a working volume of 137 m³, equivalent to 5.3 hours of tailings production.
Thickened tailings are pumped from the tailings holding tank to a horizontal plate/frame pressure filter for dewatering. Total filtration area is 620 m². The pressure filter produces a filter cake containing 10% moisture. Filtrate gravitates directly to the tails filter filtrate tank before being pumped back to the tailings thickener. The filter cake discharges onto a stockpile, where it is reclaimed by a FEL. The filtered tails is either transported and stored off site or fed to the paste backfill plant.
17.4.11
Tailings Paste Backfill
There are four Eimco 51 m² (2.7 m diameter x 5 disc) filters, three operating and one standby. Three vacuum pumps are provided, two operating, one standby.
In paste backfill mode, thickened flotation tailings are pumped to two 137 m³ storage tanks which provide approximately 11.8 hours of storage at the design processing rate. Slurry is then pumped to three disc filters by dedicated centrifugal pumps. The level in each filter bath is controlled by a pinch valve.
Each vacuum pump is supplied with filtrate receivers. The vacuum pump sealing water system consists of a closed-circuit system with a chiller to remove heat build-up from the vacuum pump operation. Make-up sealing water is supplied from a filtered water system. The disc filters produces a filter cake containing 20% moisture.
Each filter discharges directly onto the paste backfill filter discharge conveyor, which in turn transfers the filtered tailings to the paste backfill cross conveyor. This conveyor is reversible and can discharge either onto a stockpile or the paste backfill mixer feed conveyor. There is a facility to reclaim stacked tails via the paste backfill feeder conveyor, which discharges onto the mixer feed conveyor.
Filtrate flows by gravity directly to the paste backfill clarifier, where it is mixed with flocculant solution to increase clarifier performance. The clarifier produces an overflow, which is pumped to the paste backfill mixer and paste plant clean water tank via a filter. Underflow from the clarifier is pumped back to the paste backfill storage tanks, with a bleed to the clarifier feedwell to provide seed material. Excess filtrate is sand filtered to provide clean water.
The paste backfill mixer feed conveyor transfers the filtered tailings to a paste plant. Paste components are directed to the feed chute of the paste mixer along with cement dosed at a proportion of the measured dry weight of feed. Cement is transported to site via bulk tankers and pneumatically conveyed via a truck-mounted pneumatic conveyor to the storage silo. A 65 tonne capacity storage silo is installed. The silo discharges via a screw feeder to the paste backfill mixer. The measured components are added to a dual-shaft continuous mixer. Mixed paste is pumped to the underground reticulation system.
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17.5
Reagents Mixing, Storage, and Distribution
The following process reagents are necessary to operate the processing facilities:
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Hydrated lime
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Sodium cyanide
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Zinc Sulphate
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Frother: methyl isobutyl carbinol (MIBC)
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Collector: aerophine 3418A.
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Collector: sodium isopropyl xanthate (SIPX)
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Copper sulphate
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Sulphuric acid
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Flocculant
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Caustic soda
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Antiscalant
Packaged reagents are delivered to site and placed in the reagent compound. A forklift is used to transfer the drums or pallets to the preparation area.
17.5.1
Hydrated Lime
Lime slurry is used as the pH modifier in the flotation process. Hydrated lime is delivered to site as a powder in 20 tonne bulk tankers (containers). Hydrated lime is off-loaded from the tanker with a dedicated blower into a 25 tonne capacity silo. Hydrated lime is then automatically mixed with raw water on a batch basis to generate a 20% w/v slurry. The lime slurry is periodically transferred from the mixing tank to the holding tank. Both the mixing and holding tanks are agitated. Centrifugal slurry pumps distribute the slurry throughout the flotation circuits via a ring main. The pH of the circuit is measured in various places and lime dosage adjusted accordingly. Individual addition points are controlled by an automatic valve and a pH probe
Lime slurry is dosed to the following locations:
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Ball mill feed chute
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Lead regrind mill cyclone feed hopper
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Lead 2nd cleaner feed
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Lead 3rd cleaner feed
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Zinc conditioning tank 1
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Zinc regrind mill cyclone feed hopper
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Zinc cleaner scavenger feed, Zinc 1st cleaner feed and Zinc 2nd cleaner feed
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17.5.2
Sodium Cyanide
Sodium cyanide is used as a sphalerite and arsenopyrite/pyrite depressant during lead flotation. Sodium cyanide is supplied as solid briquettes in 50 kg drums.
Sodium cyanide mixing is completed manually by the operator. Sodium cyanide is added to the mixing tank using a drum tipping device. Raw water is added into the sodium cyanide mixing tank to dissolve crystalline cyanide and the solution is then transferred to the 16 m³ holding tank. A small amount of caustic soda is also added into the mixing tank to provide protective alkalinity. The hood over the mixing tank is ducted to the ventilation scrubber, which removes, scrubs and vents any gases released during the mixing process. A sodium cyanide ring main system with dedicated dosing pumps enables the required additions to the following locations:
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Ball mill feed
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Lead cleaner conditioning tank
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Lead 2nd and 3rd cleaner feed
17.5.3
Zinc Sulphate
Zinc sulphate is used as a sphalerite depressant during lead flotation. Zinc sulphate is supplied in powder form in bags.
Zinc sulphate mixing is completed manually by the operator. The mixing facility comprises a dust enclosure with bag splitter and an agitated 10 m³ capacity mixing tank. The bags are split, and zinc sulphate is then mixed with raw water to generate a 20% w/v solution. After mixing the zinc sulphate solution is pumped to a header tank for distribution to the lead flotation circuit. Zinc sulphate addition is via dedicated dosing pumps running off a ring main.
The zinc sulphate is dosed to the following locations:
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Primary mill cyclone feed hopper
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Lead regrind mill feed pump box
17.5.4
Frother
The frother used in the flotation process is MIBC. The frother is supplied as a 100% concentrated solution in 1,000 litre bulk boxes or 200 litre drums. Frother is pumped undiluted to a head tank for distribution through the flotation circuits. The frother addition is via dedicated dosing pumps running off a ring main.
The frother is dosed to the following locations:
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Flash flotation feed
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Lead rougher feed
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Lead scavenger feed
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Lead 1st cleaner feed
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Lead cleaner scavenger feed
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Zinc rougher feed
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Zinc scavenger feed
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Zinc 1st cleaner feed
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Zinc 2nd cleaner feed.
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Zinc cleaner scavenger feed
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Gold rougher feed
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Gold scavenger feed
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Gold 1st cleaner feed
17.5.5
Collector (Aerophine 3418A)
Aerophine 3418A is the collector in the lead flotation circuit. It is supplied as a 100% concentrated solution in 1,000 litre bulk boxes or 200 litre drums. The collector is pumped to a head tank for distribution through the lead flotation circuit. The lead collector addition is via dedicated dosing pumps running off a ring main.
The lead collector is dosed to the following locations:
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Flash flotation feed
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Lead conditioning tank
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Lead scavenger feed
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Lead cleaner conditioning tank
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Lead cleaner scavenger feed
17.5.6
Collector (SIPX)
SIPX is the collector in the zinc and gold-pyrite flotation circuits. SIPX is supplied as solid pellets in 120 kg drums.
SIPX mixing is completed manually by the operator. SIPX is added to the mixing tank using a drum tipping device and then mixed with raw water to generate a 10% w/v solution. The hood over the mixing tank is ducted to the ventilation scrubber, which removes, scrubs and vents any gases released during the mixing process. After mixing, the SIPX solution is pumped from the mixing tank to a header tank for distribution to the zinc and gold-pyrite flotation circuits. The SIPX solution addition is via dedicated dosing pumps running off a ring main.
The SIPX is dosed to the following locations:
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Zinc conditioning tank 2 launder
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Zinc scavenger feed
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Zinc cleaner conditioning tank
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Zinc cleaner scavenger feed
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Gold flotation conditioning tank 2
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Gold scavenger feed
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17.5.7
Copper Sulphate
Copper sulphate (CuSO4) is the zinc and arsenopyrite/pyrite activator in the flotation process. Copper sulphate is supplied in powder form in 25 kg bags.
Copper sulphate mixing is completed manually by the operator. The bags are loaded manually to the mixing tank via a bag breaker and the powder is then mixed with raw water to generate a 15% w/v solution. The hood over the mixing tank is ducted to the ventilation scrubber, which removes, scrubs and vents any gases released during the mixing process. After mixing, the copper sulphate solution is pumped to a header tank for distribution to the zinc flotation circuit and gold flotation circuit. Copper sulphate addition is via dedicated dosing pumps running off a ring main.
●
Zinc conditioning tank 2
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Zinc regrind cyclone feed hopper
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Gold flotation conditioning tank 2
17.5.8
Sulphuric Acid
Sulphuric acid is the pyrite/arsenopyrite activator in the gold flotation circuit. Sulphuric acid is delivered to site in liquid form at 95% w/v. Acid is stored in a 140 m³ capacity tank and dosing is performed using duty and standby variable speed pumps. The sulphuric acid is dosed to the gold flotation conditioning tank 1.
17.5.9
Flocculant
Flocculants are long chain molecules that aid solids settling by causing individual particles to stick together, thereby forming larger, heavier particles. Flocculant is supplied in powder form in 25 kg bags.
There is a dedicated flocculant mixing system and flocculant holding tank for the concentrate thickeners, tailings thickener and paste backfill clarifier (three systems in total). The flocculant mixing systems consist of a bag breaker and dry flocculant storage hopper, screw feeder and a 2.5 m³ mixing tank with agitator. Mixed solution gravitates to an agitated 5 m³ holding tank. Flocculant is mixed automatically with raw water on a batch basis to generate a 0.2% w/v solution. After a suitable hydration period, the flocculant is discharged into the flocculant holding tank, from where it is pumped by a positive displacement pump to the thickener.
17.5.10
Caustic Soda
Caustic soda is delivered to site in 25 kg bags of pellets. Caustic soda mixing is completed manually by the operator. The bags are loaded manually to the mixing tank via a bag breaker and the pellets are then mixed with raw water to generate a 20% w/v solution. The mixed solution gravitates to a 2.5 m³ holding tank. Caustic addition to the ventilation scrubber and the cyanide mixing tank is via dedicated lines running off a ring main.
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17.5.11
Antiscalant
Antiscalant is added to the discharge of both process water pumps to inhibit the formation of gypsum (CaSO4.H2O) on the pipes in the flotation circuit. The antiscalant is supplied as a solution in 1,000 litre bulk boxes or 200 litre drums.
Antiscalant is distributed to the discharge of the process water pumps via a variable speed dosing pump.
17.6
Services, Water, and Air Distribution
Services and utilities for the plant include process water, well, fresh water, low and high pressure compressed air, and instrument air.
The plant is serviced by a maintenance shop providing mechanical, electrical, and instrumentation facilities. An assay lab provides daily and shift quality monitoring information on solids and water samples for the safe and efficient operation of the plant.
17.6.1
Water Services
Due to the zero-discharge requirement, limited new water enters the circuit. The majority of the requirement for clean water (gland water, cooling water, etc.) is met by filtering recycled thickener and clarifier overflow. Fresh water from the underground aquifer by borehole is utilized for reagent mixing, potable water reticulation and other minor uses.
A seasonally variable quantity of water (up to 650 m3/h) is pumped from the underground mine and treated in a water treatment facility. Total suspended solids (TSS) are reduced in settling ponds followed by two clarifiers. After settling, the water is further treated with lime for pH adjustment and precipitation of metals. Once it meets discharge criteria, the water is released into the nearby Mavrolakkas stream.
Make-up water for the plant is sourced from the water treatment plant as required.
17.6.1.1
Raw Water
Raw water sourced from the mine is directed to settling ponds (three operating and one emergency pond) to remove suspended solids. Raw water supply pumps draw from the final settling pond to feed the raw water tank.
Raw water is stored in a 9 m³ water tank adjacent to the clean water and process water tanks. The raw water tank gravity feeds raw water to the following:
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Process water make-up (minimal)
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Reagent mixing
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Flocculant mixing
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17.6.1.2
Clean Water
Clean water is used for:
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Filtration
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Regrind mill flushing and seal water
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Paste backfill vacuum pump cooling and seal water make-up
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Gland service requirements
Clean water is supplied by a sand-filtering paste plant clarifier, gold tailings, and zinc and gold concentrate thickener overflows.
17.6.1.3
Fire Water
The lower portion of the raw water tank provides a dedicated fire water reservoir for the fire water system. The fire water system includes an electric-driven “jockey” pump and a diesel-driven fire water pump to ensure a continuous supply of water to the fire system in the event of a power failure.
Fire hydrants and hose reels are placed throughout the process plant, fuel storage and plant offices at intervals that ensure coverage in areas where flammable materials are present.
17.6.1.4
Process Water
Process water is stored in a 181 m³ tank. Duty and standby process water pumps provide process water to the following:
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Crushing area dust suppression
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Grinding area dilution water
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Trash screen sprays
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Flotation and regrind areas dilution and knock-down water
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De-aerator dilution water
●
General purpose hose-down points
Process water is recycled from the zinc and gold/pyrite/arsenopyrite concentrate and tailings thickener overflows, concentrate and tailings filtrates and paste backfill clarifier overflow
17.6.1.5
Lead Flotation Process Water
Reclaimed water from the lead concentrate and tailings thickeners is stored in a 100 m3 tank and reused in the grinding and lead flotation circuits. This minimizes the return of zinc flotation reagents to the lead flotation circuit.
17.6.1.6
Potable Water
Potable water is supplied in drinking bottles to the general workforce.
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Well water is produced from a bore located in the regional aquifer and pumped to a water tank for storage. This water flows by gravity from the water tank to service the administration complex, plant offices, control room, site laboratory, washing facilities, and safety shower network. This water is not used as drinking water.
17.6.2
Air Services
A single screw compressor supplies plant and instrument air to the plant air receiver (via a refrigerated air dryer) and to the crushing area air receiver.
Two dedicated air compressors supply high pressure air to separate lead, zinc and gold concentrate filter air receivers for drying purposes.
Two additional air compressors provide air services to the tailings and paste backfill disk filters, as high pressure air, to assist with cake discharge; there are separate receivers at the filter building for each duty.
Low pressure air for flotation cells is provided by five blowers, four operating and one standby.
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SECTION ● 18
Project Infrastructure
18.1
Site Access
The Olympias mine is a producing gold, zinc and lead mine located approximately 2 km west of Olympiada, in the Halkidiki peninsula, Northern Greece. It is 8 km north of the Stratoni mine, also owned by Eldorado.
The Olympias mine lies about 100 km by road from Thessaloniki and is readily accessible by car and bus. Site roads are fully developed as paved roads, with access via Highway 16.
18.2
Local Services
The town of Olympiada can supply services; also, the towns of Stavros and Stratoni are both within 20 km of the mine site. The location of the mine relative to local services is shown in Figure 18-1.
Figure 18-1: Location of Olympias Mine, Port and TMF
Mine infrastructure, including ancillary facilities and services, has been fully developed to support the existing mine production at 440 ktpa and requires minimal capital for expansion to 650 ktpa. Surface ancillary facilities are close to the decline and primary crusher. The ancillary facilities include the production services building, the surface workshop and warehouse, process plant, paste plant and surface fuel storage. The Olympias mine site layout is shown on Figure 18-2.
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Figure 18-2: Olympias Site Layout
The Olympias concentrator is shown in Figure 18-3. Located near the concentrator are the ore storage bins, water treatment plant, and a nursery.
18.3
Port Facilities
A port facility located at Stratoni, 26 km by paved road from the Olympias concentrator, is owned by Hellas Gold. Currently zinc concentrate is shipped via the port facility. Lead and arsenopyrite concentrate are bagged in the facility and shipped by truck to the port at Thessaloniki.
As part of the expansion to 650 ktpa, there is a two-phase plan to improve the port facility so that it can be fully utilized. The first phase is to rehabilitate and cover the conveyors and add dust collection. The second phase is to seismically upgrade the facility, increase the roof height, replace the floor slab, improve site water management, and increase the capacity of the existing material loading facility. These improvements will allow Hellas to ship all the concentrates to the port facility in International Carriage of Dangerous Goods by Road (ADR) certified trucks and eliminate the requirement to bag lead and arseno-pyrite concentrates. This will decrease transport costs. The changes have been budgeted and the two phases are scheduled to be completed in 2021 and 2023. They are part of the overall planned efficiency improvement initiative that will facilitate achievement of the NSR cut-off values used in the development of the mineral reserve.
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Figure 18-3: Olympias Concentrator
18.4
Communication System
Primary means of communication is by radio, using a leaky feeder cable system that advances with ramp and level development. This system provides radio communications throughout the mine.
18.5
Water Supply
Water is supplied by both surface facilities’ contact water run-off and from underground mine dewatering. A series of three settling ponds, with a fourth as a spare, is used to remove suspended solids. This water is pumped to a raw water tank next to the clean water and process water tanks.
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Excess water from mine dewatering is treated in the surface water treatment plant. The water treatment plant consists of two clarification ponds followed by a neutralization unit (if required). Treated water that is not required is then discharged into the Mavrolakkas stream. The system total capacity of 400 m3/hr is currently sufficient under normal circumstances. This system is being upgraded to handle the new anticipated maximum capacity of discharge from the underground of 650 m3/hr as mining progresses to lower elevations. An allowance has also been made for a new secondary treatment plant to be constructed to ensure alignment of water quality with regulatory requirements.
Process water is reclaimed from the tailings thickening and filtration circuit and backfill clarifier overflow. A minimum amount of make-up water is supplied from the raw water tank. The lower portion of the process water tank provides a dedicated fire water reservoir for the fire water system. The fire water system includes an electrically-driven “jockey” pump and a diesel-driven fire water pump to ensure a continuous supply of water to the fire system in the event of a power failure.
Potable water is produced from a borehole located in the regional aquifer and pumped to the potable water tank for storage. Potable water gravitates from the potable water tank to service the administration complex, plant offices, control room, site laboratory, ablution facilities, and the safety shower network.
18.6
Mine dewatering System
Details for the underground dewatering system are provided in Section 16 of this report.
18.7
Mine Waste Management
Waste from the underground mine is brought to surface, where it is recycled underground as backfill or hauled to the waste disposal facility.
Thickened tailings are pumped to the paste plant. If thickened tailings production exceeds the maximum capacity of the paste plant, excess thickened tailings can be pumped to one of two tailings pressure filters located adjacent to the concentrator. Filtered tailings are then trucked to the paste plant for additional feed into the plant when required.
Excess thickened tails are pumped to a tailings filter to produce a cake with a moisture content of 13 wt% solids, which is then trucked to the TMF for dry stacking.
In addition to mine tailings from the Olympias mining operations, tailings from historic mining activities at Olympias are also being hauled to the Kokkinolakkas TMF. This facility is located 8.5 km South of Olympias (23 km by public paved road). It is designed to safely manage approximately 10.5 Mm3 of mine waste at an average dry density of 1.6 t/m3. An overview photograph of the facility under construction is provided in Figure 18-4.
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Figure 18-4: Kokkinolakkas TMF
The TMF consists of upstream and downstream rockfill embankments, which act as tailings dams. The upstream embankment is designed to serve two purposes: as a water-retaining dam to prevent contaminated seepage leaving the facility, and to provide flood containment control. It is designed with a central clay core and associated filter and transition zones located upstream and downstream of the core. The downstream embankment is designed as a water retaining rockfill dam, with a steeper concrete face instead of geosynthetic bedding layers.
A liner system to minimize seepage for the facilities has been engineered that consists of a geosynthetic clay liner (GCL) overlain by a textured HDPE geomembrane and protected by a non-woven heavy-duty geotextile. A sandy gravel layer overlays the membrane to collect any seepage or drainage; it is also protected by a non-woven geotextile.
18.8
New Facilities
Construction of a new geology preparation laboratory, technical services building and expansion to the existing administration building is planned to finish in 2020. This is part of the support for the planned production increase.
18.9
Surface Workshop
There is an existing surface maintenance workshop with two bays located near the East decline portal, with a nearby parts warehouse attached to it. Equipment that regularly exits the mine, such as the haulage fleet and light vehicles, can be serviced at the surface shop, whilst equipment such as production scoop trams (LHDs) jumbos, and production drills, will generally be serviced in the underground workshop, once completed.
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18.9.1
Explosives / Cap Magazines
There is currently no surface or underground magazine. Explosives are brought daily to the mine from the supplier as required. There are plans to install a new set of magazines underground in 2020.
18.9.2
Fuel Storage
There is currently storage capacity for 60,000 litres of diesel fuel on site, with fuelling of equipment taking place on surface.
18.10
Shaft
There is an existing headframe, Koepe hoist and shaft on site located near the mine shop area. Currently the hoist is operational and is used for inspection of the in-shaft pumping station only. The old underground crusher and loading pocket facility is in disrepair. A high-level trade-off study was completed and it was determined that rehabilitation of the shaft for ore haulage was not economically preferable to other material handling options.
18.11
Electrical Power
Currently power is brought to site via a 10 mVA 20 kV overhead line from the local PPC grid. Power is distributed via overhead line to the various facilities throughout the site and to underground at 20 kV. Underground distribution is described further in Section 16.
To supply the required electrical power for the production increase, a new 150 kV 25 mVA Main Substation at +85 Level is being installed. As part of this capacity increase, new fiber optic cable will be installed between the various substations for rapid communications.
Currently there is 3,680 kVA of backup diesel power generation capacity available at site. An additional 2,500 kVA of back-up diesel generation in two generators will be added with the new substation, for a combined total of 6,180 kVA. Backup power is not centrally located, rather it is distributed at the various local substations, which should reduce the risk of power loss due to a failed overhead line.
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SECTION ● 19
Market Studies and Contracts
19.1
Concentrate Marketing
The Olympias plant produces three concentrates: an arsenopyrite/pyrite concentrate containing gold, a lead concentrate which also contains silver, and a zinc concentrate. Hellas Gold has negotiated multiple concentrate sales contracts with commodity traders for concentrates from Olympias. Agreements with several different customers are currently in use.
Maximizing revenue generation from Olympias requires optimization of the plant feed with respect to all three target metals. Optimized mine planning, carefully controlled blending of run-of-mine ore, and targeted operation of the concentrator are required to produce concentrate within the specifications of the current contracts and to maximize revenue at any particular time.
19.2
Concentrate Sales
Eldorado sells gold, lead and zinc concentrates to various customers from its Olympias operation. The contracts in place are valid and represented correctly in the economic analysis.
19.3
Gold Concentrate
Contracts for sales of gold concentrate contain a range of settlement terms, generally based on the grade of gold contained within the concentrate. In summary the terms used in the financial model are as follows:
●
Gold payability ranges from 59-71% and an average of 64% is used over the LOM
●
No treatment or refining changes are assumed
●
No smelter penalties are assumed
19.4
Zinc Concentrate
Contracts for sales of zinc concentrate contain a range of settlement terms. In summary the terms used in the financial model and economic analysis are as follows:
●
Treatment charges ranging from $80 - $275/dmt
●
Zinc payability is fixed at 85%
●
Gold and silver are not payable
●
Smelter penalties applied for As and other metals, and average US$0.8 per year
19.5
Lead Concentrate
Contracts for sales of lead concentrate contain a range of settlement terms. In summary the terms used in the financial and economic analysis are as follows:
●
Treatment charges ranging from of $150- $300/dmt
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Lead payability is fixed at 95%
●
Silver is paid at 95%, gold is paid at an average of 60%
●
Smelter penalties are applied for As, and other metals and average US$1 M per year
●
Gold and silver refining costs are US$20/oz and US$1.00/oz respectively
19.6
Other Contracts
19.6.1
Mining Contract
In 2019 Olympias initiated a mining development contract that is currently in place. The contractor will focus on capital development including the ramps, level access, ventilation drives, and sumps. The contractor supplies machinery, labour, consumables excluding explosives and shotcrete.
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SECTION ● 20
Environmental Studies, Permitting and Social or Community Impact
20.1
Environmental Impact Study
The EIS for the Kassandra mines mineral deposits project includes an area of 26,400 ha, in north-eastern Halkidiki (Macedonia Region). The Complex includes the Skouries, Olympias and Stratoni sites.
The EIS considers the potential impact on the local and regional environment as it relates to:
●
Open pit and underground workings
●
Tailings impoundment
●
Process plant
●
Infrastructure necessary for the Project's operation
ENVECO S.A. (Environmental Protection, Management and Economy consultants) under Hellas Gold's management, has authored the full EIS. The EIS was submitted in August 2010 and approved in July 2011. The EIS covers all environmental issues for the Project.
20.2
Legislative Framework for the preparation of the EIS
For the preparation of the EIS, legislation, standards and directives required by the national and European Community legislation in force were used.
The following assessments have been made on the Olympias Project’s environmental impact.
20.3
Landscape
Impact on the landscape of the property is of low significance, of local character and much will be non‐permanent and partially reversible. All measures to minimize the anticipated impact will be taken.
The Olympias Project will benefit the Olympias area by allowing the rehabilitation of the Olympias valley at the end of the mine life.
20.4
Geological Environment
The DAF method is primarily used for the Olympias Project. The mined areas are backfilled with material comprised of inert tailings and cement. This prevents surface subsidence and ameliorates the acidity of underground water quality due to the alkalinity of the backfill.
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20.5
Ground – Land Use
The impact will be localized to the areas of the disposal facilities but, with careful replanting at closure, will be minimized. No significant impact is anticipated on the rest of the Olympias Project areas. After the completion of operations, the Olympias Project areas will be rehabilitated according to appropriate and approved land uses.
20.6
Water Resources
No significant impact is expected either quantitatively or qualitatively on the water in the immediate mining and processing areas, or those areas within the immediate regional water system. The Olympias Project has been designed such that any leakage can be managed and controlled. The overall impacts have actually been assessed as moderately positive, due to the rehabilitation of the old tailings ponds and permanent restoration of the valley site when operations are complete.
20.7
Atmospheric Environment
Mining activities (mines, surface plants, disposal areas, etc.) are not assessed as causing any significant impact on the atmospheric environment, since all the necessary precautions have been taken for impact minimization. All potential impacts can be considered as moderate, temporary and reversible.
The Olympias Project’s impact on the atmospheric quality of the area will be limited in the direct area of influence and will be mainly related to the local and restricted increase of dust and gas values. There will be no impact at all on the quality of the atmospheric environment of the nearby populated areas.
20.8
Sound Environment
Aggravation of the sound environment is not seen as an issue in the direct and broader area of the operation. Accumulative impact on the sound environment is only anticipated in the area where the waste disposal facility is situated near to the process plant. The rest of the infrastructure elements are located sufficiently apart to avoid any significant accumulating impact on the sound environment of the Property.
20.9
Ecosystems – Flora – Fauna
All the areas used are outside of, but adjoin, Natura 2000 (an EU designated nature reserve); however, this places no restrictions on activities of Hellas Gold providing the terms of the EIS are adhered to.
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Rehabilitation will prove positive for the return of certain species of mammals. The impacts are therefore estimated to be minor, permanent and partially reversible. In some occasions, impacts are positive.
20.10
Social and Financial Environment
The positive consequences of the Olympias Project are very important for the Halkidiki Prefecture and for the National Economy. The general financial and social consequences deriving from the Olympias Project are the following:
●
A significant contribution is made to the National Economy
●
Significant infrastructure is constructed and equipped by local companies
●
Service industries in the local economy expand
●
New jobs are progressively created
20.11
Closure Plan
The fundamental criteria for Project closure and environmental rehabilitation include the following:
●
The Project site must be handed back in a state that will not give rise to risks to the health and safety of people, the flora and fauna in the area and to environmental safety in general.
●
All remaining structures, including interventions, in the natural terrain of the Project site, must not generate any risk to public health, safety, or the environment in terms of geotechnical stability.
●
All remaining materials must not generate a risk to public health or the environment for future users of the area.
●
Environmental rehabilitation must lead towards a self-sustaining ecosystem typical of the area. The purpose of the rehabilitation program must be to meet future land needs in the area, and rehabilitation must seek to re-create safe and stable biological conditions that encourage natural regeneration and the development of biodiversity.
Eldorado has undertaken to ensure that rehabilitation is in line with goals for safety, health and the environment. Surface facilities at the Olympias site will be decommissioned, and subsequently rehabilitated, once operations are completed.
Voids created from the extraction process will be hydraulically filled with a mixture of tailings and cement. When exploitation of the mineral reserves is complete, the only pending arrangements for full underground closure are the removal of mining and mechanical equipment from the workings, backfilling access works where appropriate, and the rehabilitation of the area around the entrances of the main access tunnels.
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20.12
Environmental Costs and Guarantees
Hellas Gold has provided a €50.0M ($57.5M) Letter of Guarantee to the MOE as security for the due and proper performance of rehabilitation works in relation to the mining and metallurgical facilities of the Kassandra Mines project and the removal, cleaning and rehabilitation of the old disturbed areas from the historic mining activity in the wider area of the project. Additionally, a Letter of Guarantee to the MOE, in the amount of €7.5M ($8.6M), has been provided as security for the due and proper performance of the Kokkinolakkas TMF (total €57.5M).
A ten-year monitoring period with estimated costs of US$1 Mpa per site (Stratoni, Olympias, and Skouries) is allowed for (US$30M) as a cost.
Five years after completion of the rehabilitation works and the change in land use of each individual facility in line with the master plan, the effectiveness of the guarantee will be examined by the External Auditing Committee, which will make a recommendation on any adjustment to the letter of guarantee that is needed.
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SECTION ● 21
Capital and Operating Costs
All costs are presented in US Dollars (US$) based on the exchange rates shown in Table 21-1.
Table 21-1: Exchange Rates

Currency code | Currency name | Exchange rate
US$ | United States Dollar | US$1.00 = US$1.00
C$ | Canadian Dollar | C$1.00 = US$0.75
€ | Euro | €1.00 = US$1.15

21.1
Capital Costs
The total capital cost estimate includes the expansion to 650 ktpa in the process plant, as well as capital development of the underground mine and sustaining capital, which is primarily for development, mine, processing infrastructure, and maintenance. Expansion capital specifically includes process plant upgrades, water management facilities, Phase 1 and Phase 2 of the Stratoni port facility upgrade as allocated amongst the Kassandra assets, related indirect costs, EPCM, owner’s and contingency costs. The majority of the expansion capital is spent in the first three years (2020 – 2022) Sustaining capital costs are spread out over the LOM (to 2040). Capital cost estimates are summarized in Table 21-2. Sunk costs to end of 2019 are excluded from the capital cost estimate.
Table 21-2: Summary of Capital Costs

Area | Growth ($ x 1,000) | Sustaining ($ x 1,000) | Total($ x 1,000)
Mining Dev't and Equipment | 9,616 | 89,471 | 99,087
Mine Infrastructure | 3,876 | 49,161 | 53,037
Process Plant | 18,892 | 14,356 | 33,248
Tailings + Water | 9,457 | 17,630 | 27,087
Surface Supports | 5,739 | 2,056 | 7,795
Stratoni Port Facility | 15,478 | 5,700 | 21,178
Others | 1,566 | 2,829 | 4,395
Total Direct Cost | 64,624 | 181,203 | 245,827
Indirect Cost | 6,894 | 10,990 | 17,884
Spares and Fills | 3,153 | - | 3,153
EPCM | 6,306 | 3,060 | 9,366
Owners Costs | 3,153 | - | 3,153
Contingency | 15,765 | - | 15,765
Total Installed Costs | 99,894 | 195,253 | 295,147

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Direct costs were developed from a combination of firm or budget quotes, material take-offs, existing contracts, project specific references, and historical benchmarks.
Contingency was applied to each cost item in the estimate, based on the level of engineering definition and reliability of unit rates. The overall Project contingency is 19% for the growth capital and no contingency is applied to the sustaining capital.
Labour rates were derived from existing contracts for the Olympias Phase 2 Project. The all-in crew labour rates include all direct and indirect costs associated with contractors.
Quantities were generally based on detailed material take-offs and equipment lists, with some allowances for minor items.
Indirect costs were factored from direct costs and based on historical experience.
Owner’s costs include labour and general and administrative costs for the owner’s team during the period of active construction and were applied over the three years from 2020 to 2022.
21.2
Operating Costs
Operating costs include allocations for the following.
●
Underground mining
●
Processing
●
Concentrate transport
●
Tailings filtration, materials handling, and stacking
●
Water management
●
General and administration
Operating costs are summarized in Table 21-3 and were calculated for each year of operation. It should be noted that operating costs calculated here are exclusive of refining and transport charges; these charges are excluded from the cut-off grade operating costs estimate as they are included as part of the calculation in developing the NSR model.
The underground mine operating costs were estimated based on actual operating costs to date and budget estimates that allow for the production ramp-up. Costs for THLOS which is currently not used, are estimated from factors to account for the difference in consumables and manpower related to the new methods.
The underground operating costs include all consumables (ground support, explosives, services, cement, aggregates, and fuel) and equipment required to meet the development and production schedule objectives. The operating unit costs for mobile equipment and fuel consumption rates were largely obtained from manufacturers and validated with the current operations costs. Labour requirements were developed to support the operation and maintenance of the fleet and for the general operation of the underground mine.
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Table 21-3: Summary of Operating Costs

Category | LOM ($M) | LOM average ($/t ore) | Steady State: 2024 to 2038 ($/t ore)
Mining costs | 908 | 69.6 | 48.9
Tailings and water management costs | 56 | 4.4 | 4.4
Processing costs | 652 | 50.4 | 50.4
G&A costs | 310 | 23.9 | 21.3
Total operating cost | 1,925 | 148.3 | 125.0

Process operating costs were based on estimated annual consumption of process reagents, major wear parts, and utilities. Budget quotations were obtained for supply of all significant consumables and utilities. Power consumption was calculated based on electrical load lists.
Operating costs for the water management facility are based on operation and maintenance requirements of pumps, pipelines, and mobile equipment. Labour requirements for operation of mobile equipment and supervision were calculated based on fleet sizing.
General and administrative costs were estimated based on a projected personnel list with salaries indicative of local standards, and annual allowances for general supplies.
Underground mine operating costs were based on the production schedule, equipment operating costs from manufacturers, quotations on bulk items from primarily Greek suppliers and the staffing plan and consumables as per the 2020 budgetary quotes.
The process operating costs were based on consumption inputs from the design criteria, load lists, updated quotations of consumables and the staffing plan.
General and Administrative (G&A) costs were estimated based on actuals from the 2018-2019 years and include allocations of costs from a general Kassandra administrative centre and from the Athens office which overseas permitting and other related tasks.
Salaries and wages were provided as full-burden rates by Hellas Gold, and based on detailed personnel lists. Energy costs for diesel and power supply, consumable and maintenance costs have been obtained from supplier quotations, relevant in-house data, or were factored.
Current labour productivities are considered very low when compared to other international operations. Eldorado believes that the envisaged improvement initiatives are reasonable and achievable and has allowed for effective work hours per shift to gradually increase from 4 hrs to 5.5 hrs. This will be assisted by the increase in number of available faces for DAF mining though design and sequence changes. The introduction of more cost-effective THLOS mining for approximately one third of the steady-state tonnage is also a major factor in the future operating costs schedules. All of these initiatives should be in place as steady state by 2024. To illustrate the positive impact of these initiatives on LOM operating costs, a separate column was created in Table 21-3 showing estimated costs from 2024 to 2038.
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SECTION ● 22
Economic Analysis
22.1
Methods, Assumptions and Basis
The economic analysis is based on the mineral reserves as defined in Section 15, the mining methods and production schedule as shown in Section 16, the recovery and processing methods as described in Section 17, and the capital and operating costs as outlined in Section 21.
The metal prices used in the economic model are $1,400/oz Au, $18/oz Ag, $2,100/t Pb and $2,400/t Zn. The economic model was also evaluated for sensitivity on metal prices, capital and operating costs at ±15%. All results reported in this section are in US$.
The economic model has been prepared on an annual life of mine basis. The effective date of the estimate is assumed to be December 31, 2019. The LOM for the current mineral reserves is projected to be 21 years.
22.2
Cash Flows
The annual cash flow forecast is built from the production schedule and capital and operating cost models. The model results are shown in Table 22-1 and Table 22-2. The build details of the production schedule, the operating and capital costs, the mill production and offsite costs, and the net cashflows over the LOM are shown.
The after-tax cash flow analysis shows that Olympias is a robust project once the additional capital is applied and the project reaches a steady state level of production. Beyond 2022 the project cashflows turn positive. The net present value (NPV) of the Project is estimated to be $1,067 M, using a discount rate of 5%, with positive after-tax net cash flows projected in all years other than 2019, the initial year of expansion capital expenditure.
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Table 22-1: Key Economic Results

Olympias | Unit | Value
Total ore milled | kt | 12,925
Gold grade | g/t | 7.02
Silver grade | g/t | 119
Lead grade | % | 4.1
Zinc grade | % | 5.3
Gold contained | koz | 2,920
Silver contained | koz | 49,300
Lead contained | kt | 525
Zinc contained | kt | 680
Payable gold | koz | 1,637
Payable silver | koz | 35,637
Payable lead | kt | 408
Payable zinc | kt | 494
Net revenue | $M | 4,501
Total Capital costs | $M | 295
Operating costs (total) | $M | 1,925
Transport, treatment and refining costs | $M | 408
Royalties | $M | 67
Corporate tax | $M | 423
Net post-tax cash flow | $M | 1,866
Post-tax NPV @ 5% discount rate | $M | 1,061
Post-tax NPV @ 8% discount rate | $M | 783

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Table 22-2: Olympias Mine Cashflow Analysis

Description |  | LOM | 2020 | 2021 | 2022 | 2023 | 2024 | 2025 | 2026 | 2027 | 2028 | 2029 | 2030 | 2031 | 2032 | 2033 | 2034 | 2035 | 2036 | 2037 | 2038 | 2039 | 2040
Ore and Waste Production Schedule
East Zone Ore | t | 6,890,063 | 389,241 | 376,469 | 392,632 | 357,649 | 251,146 | 350,584 | 299,611 | 295,480 | 325,598 | 352,132 | 413,562 | 313,222 | 215,021 | 221,602 | 298,568 | 307,410 | 300,431 | 316,216 | 247,376 | 320,912 | 309,895
West Zone Ore | t | 4,135,387 | - | 24,294 | 34,820 | 130,419 | 257,457 | 189,242 | 235,044 | 227,937 | 219,091 | 216,740 | 142,285 | 219,693 | 270,108 | 269,370 | 269,370 | 269,369 | 267,830 | 180,041 | 293,527 | 266,161 | 152,587
Flats Ore | t | 1,105,322 | 16,169 | 16,169 | 16,169 | 60,352 | 60,352 | 60,352 | 60,352 | 60,352 | 60,352 | 60,352 | 60,352 | 60,352 | 60,352 | 60,352 | 60,352 | 60,352 | 60,352 | 50,510 | 50,510 | 50,510 | 44,184
Remnants Ore | t | 1,075,747 | 34,661 | 23,070 | 56,390 | 101,585 | 81,048 | 49,870 | 54,996 | 66,191 | 45,120 | 20,734 | 33,801 | 56,735 | 104,516 | 98,686 | 21,720 | 12,881 | 21,387 | 103,233 | 58,613 | 12,413 | 10,636
Lateral waste | t | 2,682,377 | 285,316 | 301,379 | 270,703 | 229,363 | 243,957 | 270,231 | 256,898 | 249,397 | 190,815 | 96,966 | 7,075 | 14,302 | 17,230 | 21,569 | 24,487 | 4,900 | 6,442 | 31,192 | 17,181 | 9,167 | 6,058
Vertical waste | t | 24,851 | 1,863 | 3,617 | 2,840 | 615 | 2,334 | 1,798 | 2,933 | 4,619 | 3,255 | - | - | - | - | - | - | - | - | - | - | - | -
Total waste | t | 2,707,228 | 287,179 | 304,996 | 273,543 | 229,978 | 246,291 | 272,029 | 259,831 | 254,016 | 194,071 | 96,966 | 7,075 | 14,302 | 17,230 | 21,569 | 24,487 | 4,900 | 6,442 | 31,192 | 17,181 | 9,167 | 6,058
Mill Feed Production Schedule
Ore tonnage | t | 12,925,000 | 415,000 | 440,000 | 500,000 | 650,000 | 650,000 | 650,000 | 650,000 | 650,000 | 650,000 | 650,000 | 650,000 | 650,000 | 650,000 | 650,000 | 650,000 | 650,000 | 650,000 | 650,000 | 650,000 | 650,000 | 550,000
Gold | g/t | 7.02 | 7.22 | 7.72 | 7.65 | 7.8 | 7.31 | 7.7 | 7.35 | 6.49 | 6.77 | 6.63 | 7.19 | 7.01 | 7.29 | 7.49 | 7.26 | 7.06 | 6.99 | 7.13 | 7.89 | 4.18 | 5.11
Silver | g/t | 119 | 100.28 | 98.23 | 126.95 | 123.91 | 140.78 | 127.46 | 149.86 | 104.79 | 112.67 | 105.5 | 125.41 | 133.24 | 112.5 | 113.09 | 131.22 | 138.45 | 126.66 | 100.13 | 103.14 | 91.33 | 103.6
Lead | % | 4.1 | 3.3 | 3.1 | 4 | 4 | 4.7 | 4.2 | 4.9 | 3.5 | 3.5 | 3.7 | 4.3 | 4.7 | 3.8 | 4 | 4.5 | 4.9 | 4.6 | 3.8 | 3.8 | 3.3 | 3.6
Zinc | % | 5.3 | 4 | 3.5 | 3.9 | 4.3 | 4.9 | 5.1 | 5.3 | 4.8 | 5.1 | 5.2 | 6.1 | 6 | 5.4 | 5.9 | 5.5 | 6.1 | 5.6 | 4.9 | 5.2 | 5.9 | 5.5
Float Concentrates Produced
Lead concentrate | dmt | 704,263 | 18,734 | 19,997 | 22,514 | 32,499 | 33,602 | 31,746 | 31,649 | 33,284 | 32,475 | 33,331 | 41,968 | 41,035 | 39,109 | 38,690 | 35,997 | 36,068 | 38,740 | 35,256 | 32,307 | 35,840 | 28,582
Zinc concentrate | dmt | 1,257,253 | 29,855 | 28,367 | 37,119 | 55,484 | 61,620 | 54,622 | 55,581 | 53,999 | 53,922 | 57,263 | 67,156 | 74,624 | 72,167 | 73,121 | 66,902 | 71,726 | 69,344 | 68,452 | 69,985 | 66,806 | 51,171
As/Py concentrated | dmt | 3,825,477 | 134,110 | 122,183 | 145,905 | 193,203 | 184,498 | 177,421 | 180,313 | 171,302 | 185,049 | 172,521 | 169,659 | 178,316 | 206,839 | 192,745 | 196,019 | 202,112 | 180,346 | 202,485 | 196,067 | 199,625 | 163,765
Flotation Tailings Produced
Floatation tailings tonnes | dmt | 7,419,527 | 257,373 | 269,455 | 294,473 | 368,819 | 370,283 | 386,259 | 382,461 | 391,376 | 378,716 | 386,844 | 371,218 | 356,028 | 331,883 | 345,456 | 351,093 | 340,106 | 361,569 | 343,807 | 351,666 | 347,725 | 273,784
Float tailings as paste to ug | dmt | 6,603,260 | 220,036 | 220,001 | 250,005 | 325,003 | 325,002 | 325,024 | 325,002 | 324,980 | 325,081 | 324,979 | 325,000 | 325,002 | 324,999 | 325,005 | 325,005 | 325,006 | 325,000 | 325,000 | 325,013 | 324,998 | 258,651
Float tailings to Kokinolakkas | dmt | 816,267 | 37,337 | 49,454 | 44,468 | 43,817 | 45,282 | 61,235 | 57,459 | 66,396 | 53,635 | 61,865 | 46,218 | 31,026 | 6,884 | 20,450 | 26,088 | 15,100 | 36,569 | 18,807 | 26,653 | 22,727 | 15,133
Lead Concentrate
Recovery of Pb | % | 83.52 | 70.00 | 70.00 | 70.00 | 85.00 | 85.00 | 85.00 | 85.00 | 85.00 | 85.00 | 85.00 | 85.00 | 85.00 | 85.00 | 85.00 | 85.00 | 85.00 | 85.00 | 85.00 | 85.00 | 85.00 | 85.00
Recovery of Au | % | 4.45 | 2.78 | 3.27 | 3.23 | 3.52 | 3.73 | 3.51 | 3.52 | 3.88 | 3.77 | 3.89 | 4.80 | 5.26 | 6.24 | 5.42 | 5.35 | 4.54 | 5.14 | 4.39 | 4.93 | 5.36 | 4.91
Recovery of Ag | % | 77.57 | 73.00 | 73.00 | 76.00 | 78.00 | 78.00 | 78.00 | 78.00 | 78.00 | 78.00 | 78.00 | 78.00 | 78.00 | 78.00 | 78.00 | 78.00 | 78.00 | 78.00 | 78.00 | 78.00 | 78.00 | 78.00
Grade of Pb | % | 62.01 | 53.00 | 53.00 | 53.00 | 63.00 | 63.00 | 63.00 | 63.00 | 63.00 | 63.00 | 63.00 | 63.00 | 63.00 | 63.00 | 63.00 | 63.00 | 63.00 | 63.00 | 63.00 | 63.00 | 63.00 | 63.00
Zinc Concentrate
Recovery of Zn | % | 90.54 | 85.00 | 85.00 | 88.00 | 91.00 | 91.00 | 91.00 | 91.00 | 91.00 | 91.00 | 91.00 | 91.00 | 91.00 | 91.00 | 91.00 | 91.00 | 91.00 | 91.00 | 91.00 | 91.00 | 91.00 | 91.00
Grade of Zn | % | 52.00 | 52.00 | 52.00 | 52.00 | 52.00 | 52.00 | 52.00 | 52.00 | 52.00 | 52.00 | 52.00 | 52.00 | 52.00 | 52.00 | 52.00 | 52.00 | 52.00 | 52.00 | 52.00 | 52.00 | 52.00 | 52.00
Gold Concentrate
Recovery of Au | % | 86.00 | 84.5 | 85.0 | 85.0 | 86.0 | 86.0 | 86.0 | 86.0 | 86.0 | 86.0 | 86.0 | 86.0 | 86.0 | 86.0 | 86.0 | 86.0 | 86.0 | 86.0 | 86.0 | 86.0 | 86.0 | 86.0
Payability of gold | % | 64% | 58.9 | 60.8 | 63.2 | 63.2 | 63.2 | 70.8 | 63.2 | 63.2 | 63.2 | 63.2 | 70.8 | 62.1 | 62.1 | 62.1 | 62.1 | 62.1 | 70.8 | 62.1 | 62.1 | 62.1 | 62.1
Payable Metals Production
Payable lead | t | 408,183 | 10,944 | 10,406 | 15,506 | 20,417 | 23,687 | 21,489 | 24,988 | 17,512 | 17,529 | 18,922 | 21,526 | 24,001 | 19,400 | 20,033 | 22,839 | 24,883 | 23,154 | 19,175 | 19,384 | 16,657 | 15,730
Payable silver | oz | 35,637,630 | 1,015,696 | 996,998 | 1,473,480 | 1,869,614 | 2,124,130 | 1,923,195 | 2,261,093 | 1,581,117 | 1,700,043 | 1,591,831 | 1,892,251 | 2,010,402 | 1,697,465 | 1,706,307 | 1,979,833 | 2,089,008 | 1,911,115 | 1,510,840 | 1,556,233 | 1,377,962 | 1,369,018
Payable zinc | t | 494,297 | 12,668 | 11,115 | 14,150 | 20,519 | 23,241 | 23,989 | 25,268 | 23,032 | 24,354 | 24,934 | 29,122 | 28,699 | 25,497 | 28,231 | 26,299 | 29,078 | 26,493 | 23,221 | 24,739 | 28,141 | 21,507
Payable gold | oz | 1,637,572 | 52,829 | 58,033 | 67,681 | 90,713 | 84,735 | 99,952 | 85,172 | 75,504 | 78,779 | 76,983 | 93,279 | 79,796 | 83,296 | 85,610 | 82,760 | 80,383 | 90,617 | 81,443 | 90,337 | 47,454 | 52,215
Revenue from Concentrate Sales
Gross lead revenue | $M | 857 | 23 | 22 | 33 | 43 | 50 | 45 | 52 | 37 | 37 | 40 | 45 | 50 | 41 | 42 | 48 | 52 | 49 | 40 | 41 | 35 | 33
Gross zinc revenue | $M | 1,186 | 30 | 27 | 34 | 49 | 56 | 58 | 61 | 55 | 58 | 60 | 70 | 69 | 61 | 68 | 63 | 70 | 64 | 56 | 59 | 68 | 52
Gross silver revenue | $M | 641 | 18 | 18 | 27 | 34 | 38 | 35 | 41 | 28 | 31 | 29 | 34 | 36 | 31 | 31 | 36 | 38 | 34 | 27 | 28 | 25 | 25
Gross gold revenue | $M | 2,293 | 74 | 81 | 95 | 127 | 119 | 140 | 119 | 106 | 110 | 108 | 131 | 112 | 117 | 120 | 116 | 113 | 127 | 114 | 126 | 66 | 73
Total gross revenue | $M | 4,977 | 146 | 148 | 188 | 253 | 262 | 277 | 273 | 226 | 236 | 236 | 280 | 267 | 249 | 260 | 263 | 272 | 273 | 237 | 255 | 194 | 182
Treatment Transport and Refining and Royalties Costs
Total transport, treatment and refining cost | $M | 408.4 | 28.6 | 23.1 | 13.6 | 18.3 | 19.8 | 19.0 | 20.7 | 17.0 | 17.8 | 17.9 | 19.8 | 20.7 | 19.1 | 19.6 | 20.3 | 21.9 | 20.0 | 18.4 | 18.6 | 18.4 | 15.7
Total royalties | $M | 66.8 | 1.78 | 1.95 | 2.63 | 3.54 | 3.53 | 3.90 | 3.63 | 3.08 | 3.22 | 3.18 | 3.83 | 3.49 | 3.38 | 3.52 | 3.49 | 3.53 | 3.71 | 3.23 | 3.53 | 2.33 | 2.33
Net Revenue
Net revenue total | $M | 4,501 | 115 | 123 | 172 | 231 | 239 | 254 | 249 | 206 | 215 | 215 | 256 | 243 | 227 | 237 | 239 | 247 | 250 | 216 | 232 | 173 | 164
Net rev per ore tonne | $/t | 348 | 259 | 279 | 343 | 355 | 368 | 391 | 383 | 317 | 331 | 331 | 394 | 374 | 349 | 365 | 367 | 380 | 384 | 332 | 358 | 266 | 299
Operating Costs
Mining | $M | 908.1 | 47.26 | 49.51 | 47.45 | 49.46 | 47.48 | 47.49 | 45.97 | 48.04 | 44.85 | 43.70 | 42.12 | 40.91 | 38.88 | 39.46 | 41.20 | 40.36 | 40.48 | 41.27 | 39.04 | 40.14 | 33.02
Processing | $M | 651.7 | 21.84 | 21.84 | 25.29 | 32.74 | 32.74 | 32.74 | 32.74 | 32.74 | 32.75 | 32.74 | 32.74 | 32.74 | 32.74 | 32.74 | 32.74 | 32.74 | 32.74 | 32.74 | 32.74 | 32.74 | 26.15
Tailings and Water | $M | 56.2 | 1.00 | 1.12 | 1.71 | 2.35 | 3.01 | 3.16 | 3.12 | 3.22 | 3.09 | 3.17 | 3.04 | 2.91 | 2.72 | 2.82 | 2.87 | 2.74 | 2.96 | 2.77 | 2.88 | 2.84 | 2.74
G+A | $M | 309.8 | 18.11 | 16.11 | 14.25 | 14.60 | 14.60 | 14.60 | 14.60 | 14.60 | 14.60 | 14.60 | 14.60 | 14.60 | 14.60 | 14.60 | 14.60 | 14.60 | 14.60 | 14.60 | 14.60 | 14.60 | 13.13
Total operating costs | $M | 1,925.8 | 88.22 | 88.57 | 88.70 | 99.15 | 97.83 | 97.99 | 96.43 | 98.59 | 95.29 | 94.21 | 92.50 | 91.16 | 88.94 | 89.62 | 91.42 | 90.45 | 90.78 | 91.38 | 89.26 | 90.32 | 75.03
Capital Costs
Development / expansion capital | $M | 99.8 | 12.3 | 52.6 | 29.3 | 5.1 | 0.3 | 0.0 | 0.0 | 0.0 | 0.0 | 0.0 | 0.0 | 0.0 | 0.0 | 0.0 | 0.0 | 0.0 | 0.0 | 0.0 | 0.0 | 0.0 | 0.0
Sustaining, mining capital | $M | 195.2 | 35.0 | 29.0 | 13.7 | 14.1 | 8.4 | 21.6 | 10.2 | 9.2 | 10.0 | 4.0 | 2.7 | 6.0 | 2.8 | 2.6 | 6.3 | 2.4 | 2.3 | 5.9 | 2.8 | 2.6 | 2.6
Total capital | $M | 295.1 | 47.4 | 81.6 | 43.0 | 19.2 | 8.7 | 21.6 | 10.2 | 9.2 | 10.0 | 4.0 | 2.7 | 6.0 | 2.8 | 2.6 | 6.3 | 2.4 | 2.3 | 5.9 | 2.8 | 2.6 | 2.6
Taxation
Tax rate |  |  | 27% | 26% | 25% | 25% | 25% | 25% | 25% | 25% | 25% | 25% | 25% | 25% | 25% | 25% | 25% | 25% | 25% | 25% | 25% | 25% | 25%
Corporate taxes paid | $M | 423.3 | 0 | 0 | 0 | 0 | 16.6 | 19.8 | 18.5 | 7.6 | 10.3 | 10.4 | 34.2 | 33.2 | 30.8 | 33.6 | 33.5 | 36.2 | 37.0 | 28.7 | 33.5 | 19.0 | 20.6
Cash Flow
Undiscounted cashflow after tax | $M | 1,851 | (-35.8) | (-68.4) | 31.8 | 113.7 | 121.9 | 117.1 | 125.4 | 91.5 | 100.1 | 107.8 | 127.9 | 113.1 | 105.0 | 111.9 | 106.9 | 118.0 | 119.8 | 89.1 | 107.3 | 61.2 | 77.6
Cash cost | $/oz | (-173) | 889 | 812 | 176 | (-52) | (-266) | (-164) | (-388) | (-24) | (-120) | (-167) | (-354) | (-502) | (-253) | (-325) | (-380) | (-544) | (-354) | (-125) | (-185) | (-342) | (-310)
All-in sustaining cost | $/oz | (-54) | 1,552 | 1,312 | 379 | 103 | (-167) | 52 | (-268) | 99 | 6 | (-115) | (-325) | (-427) | (-220) | (-294) | (-304) | (-514) | (-329) | (-52) | (-154) | (-287) | (-260)
Olympias Project Financial Results
After tax | Discount rate | Net present value |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
NPV0% | 0% | $1,866 |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
NPV5% | 5% | $1,061 |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
NPV8% | 8% | $783 |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |

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22.3
Royalties and Other Fees
Olympias is subject to a mineral production royalty, which is a sliding scale NSR type of royalty payable to the Greek government.
The relevant royalties for the relevant economic metals are shown in Table 22-3 to Table 22-6. The royalties are calculated on the payable metals that are produced by the site, with deductions allowable for the downstream smelter costs including refining and treatment. As such, the royalty can be considered as a net smelter return royalty.
For the Project case economics reported in this section, the gold royalty is assumed to be 2.0%, silver royalty is 1.0%, lead and zinc royalty are both 1.0%. The royalty legislation is prepared in Euros and the exchange rate used to convert to US$ is 1.2 US$: 1 Euro.
Table 22-3: Gold Royalty

Gold price ($/oz) | Royalty
From | To
691 | 1035 | 1.0%
1035 | 1265 | 1.5%
1265 | 1495 | 2.0%
1495 | 1725 | 3.0%
1725 | 1898 | 4.0%
1898 | 2070 | 5.0%
2070 | Above | 6.0%

Table 22-4: Silver Royalty

Silver price ($/oz) | NSR royalty
From | To
12 | 20 | 1.0%
20 | 29 | 1.5%
29 | 37 | 2.0%
37 | 46 | 3.0%
46 | Above | 4.0%

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Table 22-5: Lead Royalty

Lead price ($/t) | NSR royalty
From | To
1610 | 2070 | 0.5%
2070 | 2530 | 1.0%
2530 | 2990 | 1.5%
2990 | 3450 | 2.0%
3450 | Above | 2.5%

Table 22-6: Zinc Royalty

Zinc price ($/t) | NSR royalty
From | To
1610 | 2070 | 0.5%
2070 | 2530 | 1.0%
2530 | 2990 | 1.5%
2990 | 3450 | 2.0%
3450 | Above | 2.5%

22.4
Closure and Salvage Value
Closure costs are captured by the economic model in the form of an asset retirement obligation (ARO) of €57.5M in favour of the Greek State that is offset by progressive rehabilitation and anticipated salvage value to be realized at the end of the mine life.
A ten-year monitoring period with estimated costs of $1 Mpa per site (Stratoni, Olympias, and Skouries) is allowed for ($30M) as a cost to the Kassandra mines as a whole.
22.5
Taxation
Corporate income tax rates in Greece are 28% of net earnings for 2019. Favourable changes in the Greek Corporate Tax rate indicate a steady reduction in the rate by 1% each year until stable rates of 25% are applied from 2022 onwards. Corporate Tax has been applied in the financial model as follows:
●
2019, 28%
●
2020, 27%
●
2021, 26%
●
2022 and onward, 25%
Income from operations can be offset by operating costs and by depreciation of capitalized items according to a schedule of depreciation. For Olympias an average depreciation rate for all assets of 7.2% is used.
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22.6
Financing Costs
Cost of financing such as interest on loans are not included in the economic model. Olympias is assumed to be funded by Hellas Gold, and any costs or charges relating to Eldorado’s funding of the Hellas Gold subsidiary are beyond the scope of the analysis.
22.7
Third Party Interests
Hellas Gold is the 100% owner of the Kassandra mines, which includes Olympias. Eldorado owns a 95% interest in Hellas Gold, with the remaining 5% held by Aktor Enterprises Ltd., a Greek construction firm.
All data provided in this report is shown at the 100% ownership level.
22.8
Sensitivity Analysis
The economic model was subjected to a sensitivity analysis to determine the effects of changing metal prices, and of operating and capital cost assumptions on the projected financial returns. The results of the sensitivity analysis are provided in Table 22-7 and Table 22-8.
The test of economic extraction for Olympias is demonstrated by means of this sensitivity analysis. At the mineral reserve metals prices stated in Section 15, Olympias shows strong positive economics.
The sensitivity analysis also shows that the Olympias economics are robust when evaluated using lower metal price assumptions, or higher operating and capital costs.
Expansion to 650,000 tpa is required to obtain the economic performance outlined here. While there is limited capital required the process plant and no change in the overall footprint, the 650,000 tpa case still requires a permit from the Greek Government. If the project cannot expand, then the economics and the reserve that it is based on, would need further review.
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Table 22-7: Metal Price Sensitivity Analysis

Sensitivity Ranges
Parameters | Units | -15% | -7.5% | Project case | +7.5% | +15%
Gold price | $/oz | 1,200 | 1,300 | 1,400 | 1,500 | 1,600
Silver price | $/oz | 15.0 | 17.0 | 18.0 | 19.0 | 21.0
Lead price | $/t | 1,800 | 1,900 | 2,100 | 2,300 | 2,400
Zinc Price | $/t | 2,000 | 2,200 | 2,400 | 2,600 | 2,800
Results (after tax)
NPV 0% | $M | 1,327 | 1,604 | 1,866 | 2,164 | 2,415
NPV 5% | $M | 724 | 897 | 1,062 | 1,241 | 1,395
NPV 8% | $M | 519 | 655 | 783 | 925 | 1,046
Taxation | $M | 248 | 336 | 423 | 512 | 591
Royalties | $M | 42 | 52 | 66 | 76 | 120

Table 22-8: Capital and Operating Cost Sensitivity Analysis

Sensitivity Ranges
Parameter | Units | -15% | Project case | 15%
LOM Capex | $M | 256 | 295 | 347
NPV 0% (after tax) | $M | 1,927 | 1,866 | 1,836
NPV 5% (after tax) | $M | 1,105 | 1,062 | 1,030
NPV 8% (after tax) | $M | 823 | 783 | 755
LOM operating costs | $/t ore | 126 | 148 | 170
NPV 0% (after tax) | $M | 2,100 | 1,866 | 1,663
NPV 5% (after tax) | $M | 1,205 | 1,062 | 929
NPV 8% (after tax) | $M | 899 | 783 | 678

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SECTION ● 23
Adjacent Properties
The Property is located within the Kassandra mines complex, which is comprised of a group of Hellas Gold mining and exploration concessions, covering 317 km2. The other properties within the complex include Stratoni, currently in production, and the Skouries copper-gold porphyry deposit that is under development and on hold, pending permits.
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SECTION ● 24
Other Relevant Data and Information
There is no additional information or explanation required to make the technical report understandable and not misleading.
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SECTION ● 25
Interpretation and Conclusions
25.1
Introduction
Olympias is a mine with a long history that has only been recently put back into production by Eldorado. Commercial production was achieved in December 2017. Olympias is a gold-rich polymetallic carbonate replacement deposit hosted in an interlayered sequence of feldspar-biotite gneiss and marble. The orebody has a complicated geometry, with mining operations planned in five areas: East, West, Flats, Remnant and Tops. There are sufficient Mineral Reserves for a mine life of 20 years at a steady-state production rate of 650 ktpa. The mine has embarked on a series of initiatives to increase both productivity and production rates.
25.2
Geology and Mineral Resources
The geological understanding of the deposit is good and, with the aid of an extensive core relogging program, the ore types have been well defined.
The data collection, sampling, sample preparation, security, and analytical procedures adopted by Eldorado for its exploration programs meet accepted industry standards, and the QA/QC results confirm that the assay results may be relied upon for Mineral Resource estimation purposes.
An important measure of performance at any producing mine is reconciliation of the block model to the final mill production figures. The reconciliation is detailed and thorough. It is currently providing a quarterly snapshot and demonstrating that the block model, and thus the mineral resources, are valid and robust. This validates the data underpinning the model and is, by association, a good verification of the work done.
The modelling and estimation are carried out in a diligent fashion and the reconciliation mentioned above has provided corroboration of the reasonableness of the mineral resource estimates. The mineral resource was classified as measured, indicated and inferred using logic consistent with the CIM Definition Standards (2014) referred to in NI 43-101.
25.3
Mining and Mineral Reserves
Approximately 12.92 Mt of ore is planned to be mined at Olympias using a combination of DAF and TLHOS mining methods. DAF stope shapes were generated by Eldorado using MSO software to produce stope shapes based on the NSR COG of $133/t, minimum stope width of 4 m, a fixed stope height of 5 m, and a maximum FW / HW angle of 45°. The TLHOS stope shapes were created by Eldorado using a COG of $116/t to generate grade shells.
In the evaluation of underground Mineral Reserves, modifying factors were applied to the tonnages and grades of all in situ mining shapes to account for dilution and ore losses that are common to all mining operations.
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Approximately 52% of the planned production is from DAF from all mining areas, whereas 30% is TLHOS and is planned to be mined only in the Flats zone. The remainder is from development ore and remnant extraction. The final selection of the most effective and economical mining method in any area of the mine should also consider the geometric configuration of the orebody, dilution, recovery, productivity and operating cost.
The achievement of increased productivity and the steady-state production rate of 650 ktpa is dependent on the successful implementation of a series of optimization initiatives, including:
●
Introduction of the more cost effective TLHOS mining in the Flats area where ground conditions permit.
●
A ramp-up period of three years to steady-state, with increasing realization of economies of scale.
●
Improved worker and equipment productivity and reduced downtime.
●
Modifying the mining method for larger DAF stopes such that there are multiple faces available in each stope.
The new PF system has been designed to meet future backfill requirements up to 650 ktpa ore production rate and operating at up to 70% utilization. PF will be used in DAF and TLHOS stopes. Backfill optimization study work is recommended. Successful implementation of best practice backfill design would reduce the regulated cement requirements and considerably reduce operating costs.
25.4
Metallurgy and Processing
EGC embarked on a three-phase execution of the Olympias Project:
●
Phase I involved the recommissioning of the flotation circuit and processing of the existing tailings.
●
Phase II (currently in operation) was commissioned in December 2017 and included upgrading of the process facilities to process 430 ktpa of ore from the underground mine. Lead / silver (lead) and arsenopyrite / pyrite gold (gold) concentrates are shipped through the Port of Thessaloniki and zinc concentrates are shipped through the Stratoni port. Tailings are used for underground backfill or filtered and trucked from the Olympias processing facility to the Kokkinolakkas TMF.
●
Phase III involves upgrading of the existing Olympias process plant and port facilities at Stratoni to a rate of 650 ktpa.
The testwork completed to date demonstrated that the ore response to flotation was well understood, and the processing route is deemed appropriate. Actual results for the recoveries of all metals exceeded budget targets for first half of 2019. This would indicate that the processing plant is performing well above initial projections and this is anticipated to continue.
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Other conclusions include:
●
The main minerals of significance of the Olympias ores are galena, sphalerite, pyrite, and arsenopyrite. Other than quartz, all other minerals can be considered minor.
●
The liberation of all sulphide minerals is excellent at a primary grind size P80 of 120 µm.
●
Flotation of galena, sphalerite and pyrite / arsenopyrite in a sequential flowsheet is proven effective at producing lead, zinc and gold concentrates.
25.5
Infrastructure
Mine infrastructure including ancillary facilities and services have been fully developed to support the existing mine production at 440 ktpa. Surface ancillary facilities are close to the decline and primary crusher.
As part of the expansion to 650 ktpa, there is a two-phase plan to improve the port facility such that it can be fully utilized. The first phase is to rehabilitate the conveyors, cover the conveyors and add dust collection. The second phase is to seismically upgrade the facility and increase the capacity of the existing material loading facility. These improvements will allow Hellas Gold to ship all concentrates from the port facility and eliminate the requirement to bag lead and arseno-pyrite concentrates, thereby decreasing transport costs. These phased changes have been budgeted and are scheduled to be completed in 2021 and 2023, respectively.
Currently power is brought to site via a 10 mVA 20 kV overhead line from the local PPC grid and is distributed via overhead to the various site facilities and to underground. For the production increase to 650 ktpa, a new 150 kV 25 mVA Main Substation at +85 Level is being installed, along with a new fibre optic cable for rapid communications.
Excess water from mine dewatering is treated in the surface water treatment plant (total capacity 400 m3/hr). This system is being upgraded to handle the new anticipated maximum capacity of discharge from the underground of 650 m3/hr.
25.6
Financial Performance
Sufficient provision ($295M) has been made for the development of the Olympias project with respect to mine development, processing and infrastructure. Total LOM operating costs are estimated at $1.917B, for an average of $148/t ore milled. This reflects the assumed proportional use of the selected mining methods. The actual proportional use will determine actual mining costs at the time of extraction. The ability to increase productivity from the current labour force will be a key factor in realizing operating cost projections.
The after-tax cash flow analysis shows that Olympias is a robust project financially. On a post-tax basis, assuming the metal prices and payabilities used in the economic model (1400 $/oz Au, 18 $/oz Ag, 2100 $/t Pb, 2400 $/t Zn) the net present value (NPV) of the Project is estimated to be $1,061M at a discount rate of 5%, with positive cash flows in all years after 2022
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25.7
Closure
In order to ensure that rehabilitation of the Olympias Mine is in line with goals for safety, health and environment, surface facilities at Olympias will be decommissioned and the site rehabilitated once operations are completed. Voids created from the extraction process will be hydraulically filled. When mining ceases, equipment will be removed from the workings and the area around the portals rehabilitated. Hellas Gold has provided the Mining & Industrial Minerals Directorate of the Ministry of the Environment, with a letter of guarantee for €50M in favour of the Greek State as an assurance that the funds necessary for rehabilitation projects will be available. Additionally, a Letter of Guarantee to the MOE, in the amount of €7.5M, has been provided as security for the due and proper performance of the Kokkinolakkas TMF.
25.8
Concluding Statement
Olympias mine economics are projected to be very robust. Continued efforts to bring the mine into full production as soon as possible are recommended.
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SECTION ● 26
Recommendations
26.1
Mining and Mineral Reserves
●
It is recommended that additional drilling be carried out in the Inferred Mineral Resource and further along strike to the south and down plunge where the deposit is open in order to potentially extend mine life.
●
Eldorado should continue to evaluate opportunities to apply TLHOS methods in good ground conditions.
●
Olympias should initiate a backfill geotechnical study to optimize the strength requirements of backfill.
●
Olympias should prepare a technical brief for submission to the mining authorities on an optimized backfill strength design using current industry best practice for stable openings of cured fill.
●
Current labour productivities are considered low compared to other international operations and industry best practice. A detailed and monitored action plan should be in place to support the improvement initiative programs that include increasing productivity from the labour force over the next few years to support increased production rates.
26.2
Metallurgy and processing
●
Vibratory screening as a method of classification and the replacing the current cycloning in the primary grinding circuit should be considered for improved grinding, reduced slimes, and reduced operating costs through reduced recirculating loads.
●
Olympias ores should be tested, for their amenability to dense media separation (DMS). This could be an important factor, particularly for Phase III plant expansion consideration. DMS, as utilized on many lead-zinc ores, would pre-concentrate ROM ores by separating low grade waste after crushing and ahead of grinding. Potential benefits of DMS include increased flotation feed grades, reduced cut-off grades (thereby increasing resource tonnes), reduced capital costs for mill expansion, reduced operating costs, and the possibility of producing NAG waste ahead of grinding.
●
Use of ore sorting as a pre-concentration stage to remove waste material from the ROM stream should be investigated. Commercially available systems are available that separate particulate material into two streams based on particle bulk density. A crushed and sized fraction of the ROM stream is generally treated, with the lighter waste fraction being discarded as waste. Significant upgrading of the existing plant feed, and consequent significant savings in energy required for comminution, are possible if waste is rejected prior to grinding.
●
Rejection of arsenic and antimony in the lead concentrate is challenging. Entrainment is believed to be a factor and thus column flotation, which involves a high degree of froth washing, is recommended to aid in the rejection of arsenic in the lead concentrate.
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●
Gold, sulphur, and arsenic grades of the gold-pyrite concentrate should be optimized in order to maximize smelter returns.
●
Studies should be continued into options for extracting more value from the gold concentrate currently being sold directly to smelters and traders worldwide. Technical solutions do exist for extraction of much of the contained gold but these need to be able to demonstrate acceptable rates of return and to comply with environmental and safety requirements. If a viable extraction method is developed, then construction of the new metallurgical facility would follow.
26.3
Infrastructure and other
●
Eldorado should maintain a clear focus and progress on the expansion to 650 ktpa, inclusive of the two-phase plan to improve the port facility, which will decrease transport costs. These phased changes have been budgeted and are scheduled to be completed in 2021 and 2023 respectively.
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SECTION ● 27
References
Behre Dolbear 1999, A Review of the Data and Methodology Used in a Resource Estimate, Olympias Project, Stavros, Chalkidiki, Greece, May 1999.
ENVECO S.A. 2010, Environmental Impact Assessment of the Mining‐Metallurgical Facilities of Company Hellas Gold in Halkidiki for Hellas Gold S.A.
Forward, P., Francis, A., and Lidell N. 2011, Technical Report on the Olympias Project, Pb Zn Ag Deposit, Northern Greece, effective date 14 July 2011.
Himmerkus, F., Reischmann, T., and Kostopoulos, D. 2006, Geophysical Research Abstracts Permo‐ Carboniferous and Upper Jurassic Basement ages in the Kerdilion Unit, eastern Serbo‐Macedonian Massif, Northern Greece, Vol 8, 05758.
Kalogeropoulos, S., Kilias, S., Bitzios, D., Nicolaou, M., and Both, R. 1989, Genesis of the Olympias Carbonate Hosted Pb‐Zn Deposit, Eastern Chalkidiki Peninsula, Northern Greece, Economic Geology Vol 84, pp 1,210‐1,234.
Kalogeropoulos, S. and Theodoroudis, A. 1988, K/Ar and Rb/Sr of the Minerals Biotite, Muscovite and Amphibolite from Rocks of the Servo‐Macedonian Massif, N Greece, Unpublished report IGME Athens Greece.
Kilias, A., Falalakis, G., and Mountrakis, D. 1999, Cretaceous‐Tertiary Structures and Kinematics of the Serbo‐ Macedonian Metamorphic Rocks and Their Relation to the Exhumation of the Hellenic Hinterland (Macedonia Greece), International Journal of Earth Sciences 88, pp 513‐531.
Papadakis, A. 1971, On the Age of Granitic Intrusion, Chalkidiki (Greece, Annales Geol Pays Hellenic Vol 23, pp 297‐300.
Siron, C.R. 2018, Revised Geological Controls on Ore Formation of the Olympias Deposit, report for Eldorado Gold Corporation, 16 June 2108.
Siron, C.R., Rhys, D., Thompson, J.F.H., Baker, T., Veligrakis, T., Camacho, A., and Dalampiras, L. 2018, Structural controls on porphyry Au-Cu and Au-rich polymetallic carbonate-hosted replacement deposits of the Kassandra Mining District, northern Greece: Economic Geology, Vol. 13, p. 309-345.
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SECTION ● 28
Date and Signature Page
Date and Signature Page
The effective date of this report entitled “Technical Report, Olympias Mine, Greece” is December 31, 2019. It has been prepared for Eldorado Gold Corporation by David Sutherland, P. Eng., Colm Keogh, P.Eng, Paul Skayman, FAusIMM, Ertan Uludag, P.Geo. and Sean McKinley P.Geo. , each of whom are qualified persons as defined by NI 43-101.
Signed the 28th day of February 2020.

“Signed and Sealed” David Sutherland David Sutherland, P. Eng. | “Signed” Colm Keogh Colm Keogh, P. Eng.
“Signed” Paul J. Skayman Paul J. Skayman, FAusIMM | “Signed and Sealed” Ertan Uludag Ertan Uludag P. Geo.
“Signed and Sealed” Sean McKinley Sean McKinley, P. Geo. |

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CERTIFICATE OF QUALIFIED PERSON
David Sutherland, P. Eng.
1188 Bentall 5, 550 Burrard St.
Vancouver, BC
Tel: (604) 601-6658
Fax: (604) 687-4026
Email: david.sutherland@eldoradogold.com
I, David Sutherland, am a Professional Engineer, employed as Project Manager, of Eldorado Gold Corporation located at 1188 Bentall 5, 550 Burrard St., Vancouver in the Province of British Columbia.
This certificate applies to the technical report entitled Technical Report, Olympias Mine, Greece, with an effective date of December 31, 2019.
I am a member of the Engineers & Geoscientists of British Columbia. I graduated from the Lakehead University with a Bachelor of Science (Physics) in 2003 and a Bachelor of Engineering (Mechanical) in 2005.
I have practiced my profession continuously since 2005. Since receiving my profession designation, I have worked exclusively on the design of mineral processing plants, assisted on numerous National Instrument 43-101 studies and have directed engineering and procurement on three mineral processing projects through construction. For 30 years I have been working in heavy industry including operations, maintenance and construction. During this time, I have lead the design and construction of major greenfield and brownfield construction projects in Canada, Turkey, and Greece.
As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43-101.
I have visited the Olympias Mine on numerous occasions with my most recent visit occurring on October 2, 2019 (1 day).
I was responsible for coordinating the preparation of the technical report. I am responsible for the preparation or supervising the preparation of items 1, 2, 3, 4, 5, 6, 18, 20, 24, 25, 26, and 27 in the technical report.
I have had continual prior involvement with the property that is the subject of this technical report.
I am not independent of Eldorado Gold Corporation in accordance with the application of Section 1.5 of National Instrument 43-101.
I have read National Instrument 43-101 and Form 43-101FI and the items for which I am responsible in this report entitled, Technical Report, Olympias Mine, Greece, with an effective date of December 31, 2019, has been prepared in compliance with same.
As of the effective date of the technical report, to the best of my knowledge, information and belief, the items of the technical report that I was responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading
Dated at Vancouver, British Columbia, this 28th day of February 2020.
“Signed and Sealed”
David Sutherland
________________________
David Sutherland, P. Eng.

Certej Project, Romania Technical Report |

CERTIFICATE OF QUALIFIED PERSON
Colm Keogh, P. Eng.
1188 Bentall 5, 550 Burrard St.
Vancouver, BC
Tel: (604) 687-4018
Fax: (604) 687-4026
Email: colmk@eldoradogold.com
I, Colm Keogh, am a Professional Engineer, employed as Operations Manager, Olympias Mine of Eldorado Gold Corporation and residing at 1107 Miller Road, Bowen Island in the Province of British Columbia.
This certificate applies to the technical report entitled Technical Report, Olympias Mine, Greece, with an effective date of December 31, 2019.
I am a member of the Engineers & Geoscientists British Columbia (formerly the Association of Professional Engineers and Geoscientists of British Columbia) and also a member of l’Ordre des ingénieurs du Québec. I graduated with a Bachelor of Applied Science degree (Mining) from the University of British Columbia in Vancouver, British Columbia in 1988.
I have practiced my profession continuously from 1988 until 2001 and again from 2007 until present. I have been involved in mining engineering at underground base metal and precious mining operations in Canada, Ireland, Greece and Turkey. I have further provided mining engineering consultant services to underground base metal and precious metal properties in the United States, Peru, Brazil, Mexico, Mali, Canada and Ireland.
As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43-101.
I have been working at the Olympias Mine since July 2019.
I was responsible for reviewing matters related to the estimation of mineral reserves, mining methods, project infrastructure, capital and operating costs, and economic analysis. I am responsible for sections 15, 16, 21 and 22.
I have had continual prior involvement with the property that is the subject of this technical report.
I am not independent of Eldorado Gold Corporation as defined in Section 1.5 of National Instrument 43-101.
I have read National Instrument 43-101 and Form 43-101FI and the items for which I am responsible in this report entitled, Technical Report, Olympias Mine, Greece, with an effective date of December 31, 2019, has been prepared in compliance with same.
As of the effective date of the technical report, to the best of my knowledge, information and belief, the items of the technical report for which I am responsible contain all scientific and technical information that is required for disclosure.
Dated at Vancouver, British Columbia, this 28th day of February 2020.
“Signed”
Colm Keogh
__________________________
Colm Keogh, P.Eng.

Certej Project, Romania Technical Report |

CERTIFICATE OF QUALIFIED PERSON
Paul Skayman, FAusIMM
1188 Bentall 5, 550 Burrard St.
Vancouver, BC
Tel: (604) 601-6658
Fax: (604) 687-4026
Email: paul.skayman@eldoradogold.com
I, Paul Skayman, am a Professional Extractive Metallurgist, employed as Special Advisor to the COO, of Eldorado Gold Corporation and reside at 3749 West 39th Avenue in Dunbar, Vancouver, in the Province of British Columbia.
This certificate applies to the technical report entitled Technical Report, Olympias Mine, Greece, with an effective date of December 31, 2019.
I am a fellow of the Australian Institute of Mining and Metallurgy. I graduated from the Murdoch University with a Bachelor of Science (Extractive Metallurgy) degree in 1987.
I have practiced my profession continuously since 1987 and have been involved in operation and management of gold and base metal extraction operations in Australia, Ghana, Tanzania, Guinea, China Turkey and Greece. This work has also included Feasibility Studies, Project Acquisition, Development / Construction and closure of said projects.
As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43-101.
I have visited the Olympias Mine on numerous occasions with my most recent visit occurring on November 11, 2019 (1 day).
I was responsible for the preparation of the sections in this report that dealt with metallurgy and process operations and related costs and payability of the technical report. I am responsible for the preparation or supervising the preparation of items 13, 17 and 19 in the technical report.
I have had continual prior involvement with the property that is the subject of this technical report.
I am not independent of Eldorado Gold Corporation in accordance with the application of Section 1.5 of National Instrument 43-101.
I have read National Instrument 43-101 and Form 43-101FI and the items for which I am responsible in this report entitled, Technical Report, Olympias Mine, Greece, with an effective date of December 31, 2019, has been prepared in compliance with same.
As of the effective date of the technical report, to the best of my knowledge, information and belief, the items of the technical report that I was responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading
Dated at Vancouver, British Columbia, this 28th day of February 2020.
“Signed”
Paul J. Skayman
________________________
Paul Skayman, FAusIMM

Certej Project, Romania Technical Report |

CERTIFICATE OF QUALIFIED PERSON
Ertan Uludag, P.Geo.
1188 Bentall 5, 550 Burrard St.
Vancouver, BC
Tel: (604) 601-6658
Fax: (604) 687-4026
Email: ertan.uludag@eldoradogold.com
I, Ertan Uludag, am a Professional Geoscientist, employed as Resource Geologist, of Eldorado Gold Corporation and reside at 6779 Kitchener Street in the City of Burnaby, in the Province of British Columbia.
This certificate applies to the technical report entitled Technical Report, Olympias Mine, Greece, with an effective date of December 31, 2019.
I am a member of the Engineers & Geoscientists British Columbia (formerly the Association of Professional Engineers and Geoscientists of British Columbia). I graduated from Middle East Technical University in Ankara Turkey with Bachelor of Science in Geological Engineering in July 1994.
I have practiced my profession continuously since 1996. I have been involved in ore control, mine geology and resource modelling work on gold, copper, zinc, lead and silver underground and open pit properties in Turkey, China, Greece, Canada and Romania, and rock mechanics work in South Africa
As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43-101.
I have visited the Olympias Mine on numerous occasions with my most recent visit occurring on August 21 to 27, 2019.
I was responsible for the mineral resources and the preparation of related sections on sample preparation and analyses, data verification and mineral resource estimation for this technical report. I am responsible for the preparation or supervising the preparation of items 11, 12 and 14 in the technical report.
I have had continual prior involvement with the property that is the subject of this technical report.
I am not independent of Eldorado Gold Corporation in accordance with the application of Section 1.5 of National Instrument 43-101.
I have read National Instrument 43-101 and Form 43-101FI and the items for which I am responsible in this report entitled, Technical Report, Olympias Mine, Greece, with an effective date of December 31, 2019, has been prepared in compliance with same.
As of the effective date of the technical report, to the best of my knowledge, information and belief, the items of the technical report that I was responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading
Dated at Vancouver, British Columbia, this 28th day of February 2020.
“Signed and Sealed”
Ertan Uludag
________________________
Ertan Uludag, P.Geo.

Certej Project, Romania Technical Report |

CERTIFICATE OF QUALIFIED PERSON
Sean McKinley, P.Geo.
1188 Bentall 5, 550 Burrard St.
Vancouver, BC
Tel: (604) 601-6658
Fax: (604) 687-4026
Email: Sean.McKinley@eldoradogold.com
I, Sean McKinley, am a Professional Geoscientist, employed as Senior Geologist- Resource Development, of Eldorado Gold Corporation and reside at 2231 Bellevue Ave, in the City of Coquitlam in the Province of British Columbia.
This certificate applies to the technical report entitled Technical Report, Olympias Mine, Greece, with an effective date of December 31, 2019.
I am a member of the Engineers & Geoscientists British Columbia (formerly the Association of Professional Engineers and Geoscientists of British Columbia). I graduated from Queen’s University in Kingston, Ontario with a Bachelor of Science (Honours) degree in geology in 1992 and subsequently obtained a Master of Science degree in geology from the University of British Columbia.
I have practiced my profession continuously since 1996 and have been involved in: mineral exploration (both greenfields and brownfields), mine geology (underground and open pit settings) and geological modelling on gold, copper, lead, zinc and silver projects in Canada, Ireland, Sweden, China, Mexico, Romania, Greece and Turkey.
As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43-101.
I have visited the Olympias Mine on numerous occasions with my most recent visit occurring on July 2 to 5, 2019.
I was responsible for the preparation of the sections in this report concerned with geological information, exploration and drilling for this technical report. I am responsible for the preparation or supervising the preparation of items 7, 8, 9, 10 and 23 in the technical report.
I have had continual prior involvement with the property that is the subject of this technical report.
I am not independent of Eldorado Gold Corporation in accordance with the application of Section 1.5 of National Instrument 43-101.
I have read National Instrument 43-101 and Form 43-101FI and the items for which I am responsible in this report entitled, Technical Report, Olympias Mine, Greece, with an effective date of December 31, 2019, has been prepared in compliance with same.
As of the effective date of the technical report, to the best of my knowledge, information and belief, the items of the technical report that I was responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading
Dated at Vancouver, British Columbia, this 28th day of February 2020.
“Signed and Sealed”
Sean McKinley
________________________
Sean McKinley, P.Geo.
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