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Jdsmining.ca

Vancouver Office Kelowna Office

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#860 – 625 Howe Street Vancouver, BC V6C 2T6 #200 – 532 Leon Avenue, Kelowna, BC V1Y 6J6

Revised Preliminary Economic Assessment

Technical Report

Halilağa Project,

Turkey

Effective Date: December 20, 2014

Report Date: February 16, 2015

Prepared for:

Pilot Gold Inc.

1900-1055 West Hastings Street

Vancouver, BC

V6E 2E9

Qualified Persons CompanyGordon Doerksen, P.Eng. JDS Energy & Mining Inc.

Stacy Freudigmann, P.Eng. JDS Energy & Mining Inc.

Dino Pilotto, P.Eng. JDS Energy & Mining Inc.

Maritz Rykaart, P.Eng. SRK Consulting (Canada) Inc.

Greg Abrahams, P.Geo. SRK Consulting (Canada) Inc.

Gary Simmons, MMSA GL Simmons Consulting LLC

Garth Kirkham, P.Geo. Kirkham Geosystems Ltd.

James Gray, P.Geo. Advantage Geoservices Ltd.

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HALILAĞA PROJECT – PEA TECHNICAL REPORT

Effective Date: December 20, 2014 ii

Date and Signature Page

This report entitled Revised Preliminary Economic Assessment Technical Report, Halilağa Project,

Turkey, effective as of December 20, 2014 was prepared and signed by the following authors:

Original document signed by:

Original signed and sealed

Gordon Doerksen, P.Eng. Date Signed


Stacey Freudigmann, P.Eng. Date Signed


Dino Pilotto, P.Eng. Date Signed


Maritz Rykaart, P.Eng. Date Signed


Greg Abrahams, P.Geo. Date Signed


Gary Simmons, MMSA. Date Signed


Garth Kirkham, P.Geo. Date Signed


James Gray, P.Geo. Date Signed

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Effective Date: December 20, 2014 iii

Contents

1

Executive Summary ................................................................................................. 1-15

1.1 Introduction ................................................................................................................. 1-15

1.2 Project Concept .......................................................................................................... 1-16

1.3 Property Description, Location and Ownership .......................................................... 1-16

1.4 Geology and Mineralization ........................................................................................ 1-17

1.5 Exploration and Drilling .............................................................................................. 1-18

1.6 Mineral Processing and Metallurgical Testing ............................................................ 1-19

1.7 Mineral Resource and Mineral Reserve Estimates .................................................... 1-19

1.8 Mine Development and Operations ............................................................................ 1-20

1.9 Waste Management ................................................................................................... 1-24

1.10 Project Infrastructure .................................................................................................. 1-24

1.11

Environmental Considerations ................................................................................... 1-25

1.12 Capital Costs .............................................................................................................. 1-25

1.13 Operating Costs ......................................................................................................... 1-26

1.14 Economics Analysis ................................................................................................... 1-26

1.15 Risks and Opportunities ............................................................................................. 1-28

1.16 Conclusions ................................................................................................................ 1-29

1.17 Recommendations ..................................................................................................... 1-29

2 Introduction ................................................................................................................. 2-1

2.1 Basis of Technical Report ............................................................................................ 2-1

2.2 Terms of Reference ...................................................................................................... 2-1

2.3 Scope of Work .............................................................................................................. 2-2

2.4 Qualified Person Responsibilities and Site Inspections ............................................... 2-3

2.5 QP Site Visits ............................................................................................................... 2-4

2.6 Sources of Information ................................................................................................. 2-4

2.7 Units, Currency and Rounding ..................................................................................... 2-5

3 Reliance on Other Experts ......................................................................................... 3-1

4 Property Description and Location ........................................................................... 4-1

4.1 Property Description ..................................................................................................... 4-1

4.2 Mineral Tenure ............................................................................................................. 4-2

4.3 State Royalties ............................................................................................................. 4-6

4.4 Joint Venture Agreements ............................................................................................ 4-6

4.5 Mining Rights and Title in Turkey ................................................................................. 4-7

4.6 Environmental Studies, Permitting and Social or Community Impact .......................... 4-8

4.7 Community Water Supply Considerations ................................................................. 4-10

5 Accessibility, Climate, Local Resources, Infrastructure, and Physiography ........ 5-1

5.1 Accessibility .................................................................................................................. 5-1

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5.2 Climate ......................................................................................................................... 5-1

5.3 Physiography ................................................................................................................ 5-1

5.4 Vegetation .................................................................................................................... 5-2

5.5 Existing Infrastructure ................................................................................................... 5-3

6 History to 2000 ............................................................................................................ 6-1

6.1 1988-1991 .................................................................................................................... 6-1

6.2 1997 .............................................................................................................................. 6-1

6.3 1998 .............................................................................................................................. 6-2

6.4 2000 .............................................................................................................................. 6-2

7 Geological Setting and Mineralization ...................................................................... 7-1

7.1 Regional Geology ......................................................................................................... 7-1

7.2 Property Geology ......................................................................................................... 7-5

8 Deposit Types ............................................................................................................. 8-1

8.1 General Background .................................................................................................... 8-1

8.2 Halilağa Porphyry Mineralization .................................................................................. 8-4

8.3 Kestane Porphyry Mineralization ................................................................................. 8-5

9 Exploration .................................................................................................................. 9-1

10 Drilling ....................................................................................................................... 10-1

10.1 Halilağa Drilling .......................................................................................................... 10-1

10.2 Pirentepe Drilling ...................................................................................................... 10-10

10.3 West Kestane Drilling ............................................................................................... 10-10

11 Sample Preparation, Analyses, and Security ......................................................... 11-1

11.1 Sample Preparation and Analyses ............................................................................. 11-1

11.2 Sample Preparation .................................................................................................... 11-4

11.3 Assay and Analysis .................................................................................................... 11-5

11.4 Quality Assurance and Quality Control Programs...................................................... 11-8

12 Data Verification ....................................................................................................... 12-1

12.1 Geological Site Visit ................................................................................................... 12-1

13 Mineral Processing and Metallurgical Testing ....................................................... 13-1

13.1 Metallurgical Testing .................................................................................................. 13-1

13.2 Grade and Recovery Predictions ............................................................................. 13-29

13.3 Concentrate Quality .................................................................................................. 13-33

13.4 Further Metallurgical Test Work Required ............................................................... 13-35

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14 Mineral Resource Estimates .................................................................................... 14-1

14.1 Introduction ................................................................................................................. 14-1

14.2

Available Data ............................................................................................................ 14-1

14.3 Geologic Model .......................................................................................................... 14-3

14.4 Assay Compositing ..................................................................................................... 14-4

14.5 Grade Capping ........................................................................................................... 14-5

14.6 Grade Interpolation ..................................................................................................... 14-7

14.7 Density Assignment .................................................................................................. 14-11

14.8 Model Validation ....................................................................................................... 14-12

14.9 Resource Classification and Tabulation ................................................................... 14-17

15 Mining Methods ........................................................................................................ 15-1

15.1 Mining Context ........................................................................................................... 15-1

15.2 Open Pit Optimization ................................................................................................ 15-5

15.3 Open Pit Mine Design .............................................................................................. 15-15

15.4 Mine Sequence/Phasing .......................................................................................... 15-18

15.5 OP Mine Operation ................................................................................................... 15-19

15.6 Mine Schedule .......................................................................................................... 15-21

16 Mineral Recovery Methods ...................................................................................... 16-1

16.1 Process Plant Design Considerations ........................................................................ 16-1

16.2 Processing Plant Configuration .................................................................................. 16-4

17 Project Infrastructure ............................................................................................... 17-1

17.1 General ....................................................................................................................... 17-1

17.2 General Site Arrangement .......................................................................................... 17-1

17.3 Site Access Road ....................................................................................................... 17-2

17.4 Light Vehicle Roads ................................................................................................... 17-2

17.5 Power Supply and Transmission Line ........................................................................ 17-4

17.6 Camp .......................................................................................................................... 17-5

17.7 Process Plant ............................................................................................................. 17-5

17.8 Truck Shop ................................................................................................................. 17-5

17.9 Maintenance Shop, Warehouse, Mine Dry and Administration Building ................... 17-5

17.10 Communications / IT .................................................................................................. 17-5

17.11 First Aid / Emergency Services .................................................................................. 17-6

17.12 Bulk Explosives Storage and Magazines ................................................................... 17-6

17.13 Bulk Fuel Storage and Delivery .................................................................................. 17-6

17.14 Fresh/Firewater Tank and System ............................................................................. 17-6

17.15 Process Water Tank ................................................................................................... 17-6

17.16 Potable Water ............................................................................................................. 17-7

17.17 Sewage Treatment ..................................................................................................... 17-7

17.18 Water Treatment ........................................................................................................ 17-7

17.19 Freight ........................................................................................................................ 17-7

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17.20 Tailings Storage Facilities .......................................................................................... 17-7

17.21 Water Storage .......................................................................................................... 17-11

17.22 Water Storage Design .............................................................................................. 17-12

17.23 Rock Storage Facilities ............................................................................................. 17-15

18 Market Studies and Contracts ................................................................................. 18-1

18.1 Market Studies ........................................................................................................... 18-1

18.2 Royalties ..................................................................................................................... 18-2

18.3 Metal Prices ................................................................................................................ 18-2

19 Environmental Studies, Permitting and Social or Community Impact................. 19-1

19.1 Environmental Baseline Study ................................................................................... 19-1

19.2 Mine Closure Requirements ..................................................................................... 19-19

19.3 Environmental Permitting ......................................................................................... 19-22

20 CAPITAL AND OPERATING COSTS ........................................................................ 20-1

20.1 Capital Cost Estimate ................................................................................................. 20-1

20.2 Operating Costs Estimate ........................................................................................ 20-20

21 Economic Analysis ................................................................................................... 21-1

21.1 Assumptions ............................................................................................................... 21-1

21.2 Revenues & NSR Parameters.................................................................................... 21-3

21.3 Royalties ..................................................................................................................... 21-5

21.4 Taxes .......................................................................................................................... 21-6

21.5 Economic Results ....................................................................................................... 21-7

21.6 Sensitivities................................................................................................................. 21-2

22 Adjacent Properties .................................................................................................. 22-1

22.1 Ağı Dağı and Kirazlı ................................................................................................... 22-1

22.1 TV Tower .................................................................................................................... 22-3

23 Interpretation and Conclusions ............................................................................... 23-1

24 Risks and Opportunities .......................................................................................... 24-1

24.1 Risks ........................................................................................................................... 24-1

24.2 Opportunities .............................................................................................................. 24-4

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25 Recommendations ................................................................................................... 25-6

26 list of Abbreviations ................................................................................................. 26-7

27 References ................................................................................................................ 27-1

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Tables

Table 1.1: Comparison to Initial Estimate at 0.43 g/t AuEq Cut-off ............................................... 1-20

Table 1.2: Mine Planning Optimization Input Parameters* ............................................................ 1-21

Table 1.3: PEA Proposed Mine Plan Summary ............................................................................. 1-22

Table 1.4: LOM Production Schedule ............................................................................................ 1-23

Table 1.5: Capital Cost Estimate ................................................................................................... 1-25

Table 1.6: Operating Cost Estimates ............................................................................................. 1-26

Table 1.7: Metal Price Assumptions for Economic Analysis .......................................................... 1-26

Table 1.8: Summary of Economic Results ..................................................................................... 1-27

Table 1.9: After-Tax Sensitivity Test Results ................................................................................. 1-28

Table 1.10: Cost Estimate for New Mineral Resource Estimate to Support a PFS ....................... 1-30

Table 2.1: Independent QP Responsibilities .................................................................................... 2-3

Table 4.1: Halilağa Project Licences, Çanakkale Province ............................................................. 4-3

Table 9.1: Details of the Halilağa Project by Year ........................................................................... 9-9

Table 10.1: Halilağa Drill Summary ............................................................................................... 10-1

Table 10.2: Halilağa Drill Hole Collars ........................................................................................... 10-4

Table 10.3: Assay Summary .......................................................................................................... 10-7

Table 10.4: HD -64 Skarn Zone Assays ...................................................................................... 10-10

Table 11.1: Recommended Gold Concentrations of Standards Used at Halilağa ...................... 11-10

Table 11.2: Recommended Copper Concentrations of Standards Used at Halilağa .................. 11-10

Table 11.3: Standards Performance for the Halilağa Project (2004 - 2009) ............................... 11-16

Table 11.4: Standards Performance for the Halilağa Project (2009 – 2012) ............................... 11-17

Table 11.5: Standards Performance for the Halilağa Project (2012) - Au (Pb Fire Assay) ......... 11-18

Table 11.6: Standards Performance for the Halilağa Project (2012) - Cu (4-Acid Digestion) ..... 11-19

Table 11.7: Standards Performance for the Halilağa Project (2012) - Mo (4-Acid Digestion) ..... 11-19

Table 13.1: Rougher Flotation Results for Halilağa Master Composite Samples ......................... 13-3

Table 13.2: Cleaner Flotation results for Halilağa Master Composite Samples ............................ 13-4

Table 13.3: Variability Rougher Flotation Testing Discrete Composites Drill Hole HD-01 ............ 13-7

Table 13.4: Variability Rougher Flotation Testing Discrete Composites Drill Hole HD-04 ............ 13-8

Table 13.5: Mineralogy Examination Results .............................................................................. 13-10

Table 13.6: Summary of Comminution Test Work Data .............................................................. 13-13

Table 13.7: Summary Actual and Modeled BM-Wi Data ............................................................. 13-13

Table 13.8: Cleaner Flotation results for Halilağa Composite Samples ...................................... 13-17

Table 13.9: Locked Cycle Test Results on “Typical Ore” Composite .......................................... 13-20

Table 13.10: Head Analysis ......................................................................................................... 13-21

Table 13.11: Summary Locked Cycle Flotation Testing .............................................................. 13-23

Table 13.12: Summary Cleaner Tails Cyanide Leach (24 Hrs) ................................................... 13-24

Table 13.13: Cyanide and Lime Consumption............................................................................. 13-25

Table 13.14: Twelve Hour Cleaner Tails Leach Analysis ............................................................ 13-28

Table 13.15: Summary Locked Cycle Flotation Data .................................................................. 13-29

Table 13.16: Quality of Concentrate ............................................................................................ 13-34

Table 14.1: Comparison to Initial Estimate at 0.43 g/t AuEq Cut-off ............................................. 14-1

Table 14.2: Resource Block Model Setup ..................................................................................... 14-3

Table 14.3: Grade Capping Levels ................................................................................................ 14-5

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Table 14.4: Composite Statistics ................................................................................................... 14-6

Table 14.5: Variogram Models - Copper ........................................................................................ 14-8

Table 14.6: Variogram Models - Gold ............................................................................................ 14-9

Table 14.7: Variogram Models - Molybdenum ............................................................................. 14-10

Table 14.8: Interpolation Parameters .......................................................................................... 14-11

Table 14.9: Average Rock Type Density ..................................................................................... 14-12

Table 14.10: Metal Equivalence Parameters ............................................................................... 14-13

Table 14.11: Resource Classification Criteria.............................................................................. 14-18

Table 14.12: Halilağa Sulphide Mineral Resource by Gold Equivalent Cut-off* .......................... 14-19

Table 14.13: Oxide Gold Resource ............................................................................................. 14-20

Table 14.14: Halilağa Sulphide Resource: Update vs. Initial at 0.43 g/t AuEq Cut-off ................ 14-20

Table 15.1: SRK Slope Angles per Domain ................................................................................... 15-4

Table 15.2: Input Parameters Used in the LOM Open Pit Optimization ........................................ 15-6

Table 15.3: Cut-off Grade Calculations used in Pit Optimization .................................................. 15-8

Table 15.4: Overall Optimization Results (excluding capital costs) ............................................. 15-10

Table 15.5: LOM Plan Summary ................................................................................................. 15-15

Table 15.6: Halilağa Pit/Phase Tonnages and Grades ............................................................... 15-18

Table 15.7: Major OP Equipment Assumptions (Contractor Estimates may Vary) ..................... 15-20

Table 15.8: Proposed LOM Open Pit Production Schedule ........................................................ 15-22

Table 16.1: Summary of the Process Design Criteria ................................................................... 16-2

Table 17.1: Water Storage Pits Volume Estimates ..................................................................... 17-12

Table 17.2: Scoping Level Cost Estimate Quantities .................................................................. 17-13

Table 18.1: NSR Parameters used in the Economic Analysis ....................................................... 18-1

Table 18.2: Metal Prices used in the Economic Analysis .............................................................. 18-3

Table 19.1: Sub-catchment Areas ................................................................................................. 19-6

Table 19.2: Hydrogeological Parameters for Groundwater Wells ................................................. 19-9

Table 19.3: Analyses Conducted in the Static Testing Program ................................................. 19-16 Table 19.4: Halilağa Operation Licenses ..................................................................................... 19-22

Table 19.5: Studies to be Conducted under the ESIA Framework .............................................. 19-27

Table 20.1: CAPEX Estimate ......................................................................................................... 20-2

Table 20.2: Process Plant Capital Costs ....................................................................................... 20-8

Table 20.3: Summary of Quoted Equipment ................................................................................. 20-9

Table 20.4: Installation Rates ...................................................................................................... 20-11

Table 20.5: On-Site Infrastructure Capital Costs ......................................................................... 20-12

Table 20.6: Plant Support Mobile Equipment CAPEX Estimate .................................................. 20-14

Table 20.7: Scoping Level Cost Estimate Quantities .................................................................. 20-15

Table 20.8: Tailings and Water Storage Facility Costs by Year .................................................. 20-16

Table 20.9: Mine Closure and Rehabilitation Cost Estimate ....................................................... 20-19

Table 20.10: OPEX Estimate Summary ...................................................................................... 20-21

Table 20.11: Mine Average OPEX Estimate by Function ............................................................ 20-24

Table 20.12: Processing Average Unit OPEX Estimate .............................................................. 20-24

Table 20.13: Derivation of Plant OPEX Estimate ........................................................................ 20-26

Table 20.14: Site Labour Cost Summary ..................................................................................... 20-27

Table 20.15: Process Plant Power Cost Summary (M$) ............................................................. 20-28

Table 20.16: Reagent Consumptions and Unit Costs ................................................................. 20-28

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Table 20.17: General and Administration Costs Summary ......................................................... 20-29

Table 20.18: G&A Labour Cost Summary ................................................................................... 20-30

Table 21.1: Life of Mine Plan Summary ......................................................................................... 21-2

Table 21.2: Metal Prices used in the Economic Analysis .............................................................. 21-3

Table 21.3: Summary of Results ................................................................................................... 21-8

Table 21.4: Sensitivity Results for Base Case NPV ...................................................................... 21-3

Table 21.5: Sensitivity Analysis – Foreign Exchange Assumptions by Cost Area ........................ 21-5

Table 21.6: Result of Foreign Exchange Rate Sensitivity Analysis ............................................... 21-6

Table 21.7: Discount Rate Sensitivity ............................................................................................ 21-6

Table 22.1: 2013 Resource for the Ağı Dağı and KirazlıProperties .............................................. 22-3

Table 22.2: Küçükdağ (KCD) Resource by Redox State at 0.5 g/t AuEq Cut-off .......................... 22-5

Table 24.1: Internal Project Risks .................................................................................................. 24-2

Table 24.2: Project Opportunities .................................................................................................. 24-4

Table 25.1: Cost Estimate for Additional Drilling and an Updated Mineral Resource

Estimate to Support a PFS ............................................................................................... 25-6

Table 26.1: Units of Measure & Abbreviations ............................................................................. 26-7

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Figures

Figure 1.1: LOM Production Schedule ........................................................................................... 1-23

Figure 4.1: Map of Turkey ................................................................................................................ 4-1

Figure 4.2 Map of the Biga Peninsula Showing the Halilağa Project .............................................. 4-2

Figure 4.3: License Boarders and General Arrangement of Facilities ............................................. 4-4

Figure 4.4: Halilağa License Map .................................................................................................... 4-5

Figure 5.1 General View of Halilağa (View to the Southeast) ....................................................... 5-2

Figure 5.2 Flooded Coal Pit near the Project.................................................................................. 5-3

Figure 5.3: Layout of the Port of Bandirma ..................................................................................... 5-5

Figure 5.4 Core Logging and Storage Facilities ............................................................................ 5-6

Figure 7.1: Regional Distribution of Metamorphic Rocks in the Biga Peninsula ............................. 7-3

Figure 7.2: Regional Distribution of Volcanic and Igneous Rocks in the Biga Peninsula ................ 7-4

Figure 7.3: Halilağa Local Geology .................................................................................................. 7-7

Figure 7.4: Halilağa Alteration Map ................................................................................................. 7-8 Figure 8.1: Anatomy of a Telescoped Porphyry System ................................................................. 8-2

Figure 8.2: Generalized Alteration-Mineralization Zoning Pattern for Telescoped Porphyry

Copper Deposits ................................................................................................................. 8-3

Figure 8.3: B-Veining in Kestane Porphyry ...................................................................................... 8-6

Figure 9.1: 3-D Magnetic Inversion of Halilağa Data ....................................................................... 9-3

Figure 9.2: Geological Map of Halilağa Property (Sanchez, 2008) ................................................. 9-4

Figure 9.3: Rock Geochemistry after 2008 Sampling ...................................................................... 9-5

Figure 9.4: Compiled IP Chargeability Plan for 100 m Depth .......................................................... 9-7

Figure 9.5: Compiled IP Resistivity Plan for 100 m Depth ............................................................... 9-8

Figure 10.1: Kestane Target with Drill Collars and Traces ............................................................ 10-3

Figure 10.2: Kestane Plan Map with Section Lines for the N‐S Cross Section and the E‐W Long

Section ............................................................................................................................... 10-4

Figure 10.3: Kestane Target Showing the Down Dropping Fault Bound Blocks of QFP

Section 483350 ................................................................................................................. 10-5

Figure 10.4: Kestane Horst at the Center Bounded by North and South Normal Faults ............... 10-2

Figure 11.1: Core Recovery vs. Copper Grade (%) ...................................................................... 11-2

Figure 11.2: Core Recovery vs. Gold Grade (g/t) .......................................................................... 11-3

Figure 11.3: Umpire Check Assays for Drill Hole HD-02; Cu (%)................................................ 11-12

Figure 11.4: Umpire Check Assays for Drill Hole ID HD-02; Au (ppm) ....................................... 11-13

Figure 11.5: Umpire Check Assays for Drill Hole HD-02; Mo (%) ............................................... 11-14

Figure 13.1: Location of Drill Hole ................................................................................................. 13-1

Figure 13.2: Master Composite HD01 Copper Rougher Recovery vs. Copper Grade ................. 13-5

Figure 13.3: Master Composite HD04 Cleaner Copper Recovery vs. Cleaner GoldRecovery ........................................................................................................................... 13-6

Figure 13.4: Copper and Gold Rougher Recovery Variability HD-01 ............................................ 13-8

Figure 13.5: Copper and Gold Rougher Recovery Variability HD-04 ............................................ 13-9

Figure 13.6: Location of Drill Holes HD-40 and HD-49 ................................................................ 13-12

Figure 13.7: Summary Real and Modeled BM-Wi Data .............................................................. 13-14

Figure 13.8: Summary Depth Down hole (meters) vs. BM-Wi (kWh/t) ........................................ 13-15

Figure 13.9: Cleaner Copper and Gold Recovery for Composite Samples ................................. 13-18

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Figure 13.10: Re-grind P80 vs. Cu Concentrate Grade .............................................................. 13-19

Figure 13.11: Location of Drill Holes HD-109, HD-115 and HD-124 ........................................... 13-22

Figure 13.12: Au Extraction vs. Leach Time ................................................................................ 13-26

Figure 13.13: Cyanide Consumption vs. Leach Time .................................................................. 13-27

Figure 13.14: Copper Feed Grade vs. Copper Recovery ............................................................ 13-30

Figure 13.15: Copper Concentrate Cu Grade Model ................................................................... 13-31

Figure 13.16: Gold Feed Grade vs. Gold Recovery .................................................................... 13-32

Figure 13.17: Copper Concentrate Au Grade Model ................................................................... 13-33

Figure 14.1: Halilağa Drilling and Resource Model Outline (New Holes in Red) .......................... 14-2

Figure 14.2: Sectional Interpretation 483,550E (view to west) ...................................................... 14-4

Figure 14.3: Copper Grade Swath Plots Comparing OK, ID and NN Estimates ......................... 14-14

Figure 14.4: Gold Grade Swath Plots Comparing OK, ID and NN Estimates ............................. 14-16

Figure 15.1: View of RQD Data Plotted Down Drill Hole behind Northern and Southern

Walls (25kt/d Pit Shown in Grey) ...................................................................................... 15-2

Figure 15.2: Section View Showing 3D Surface Defining the Upper and Lower Domain byHonouring RQD Values Down Drill hole ........................................................................... 15-3

Figure 15.3: Typical Example of Rock Mass Quality in Drill Core from the Upper and Lower

Domains ............................................................................................................................ 15-4

Figure 15.4: Geotechnical Domain Clipped to the 25 kt/d pit ....................................................... 15-5

Figure 15.5 Pit Optimization Results Graph .............................................................................. 15-12

Figure 15.6: Open Pit Optimization Incremental Value Results (excluding capital costs) ........... 15-13

Figure 15.7: Open Pit Optimization Incremental Tonnage Results ............................................. 15-14

Figure 15.8: Plan View Halilağa Pit Shell and Rock Storage Facilities ....................................... 15-16

Figure 15.9: Section View of Halilağa Pit Shell Showing Phase (Stage) Pushbacks .................. 15-17

Figure 15.10: Halilağa Project - Phase Summary ....................................................................... 15-19

Figure 15.11: Process Tonnes, Waste Tonnes and Strip Ratio by Period .................................. 15-23

Figure 15.12: Mineralized Tonnes and Grade by Phase and Period .......................................... 15-24 Figure 15.13: Annual Benches and Total Material Mined by Phase and Period ......................... 15-25

Figure 15.14: Period Plot Year 1 ................................................................................................. 15-27




Figure 15.18: Period Plot Year 14 ............................................................................................... 15-31

Figure 16.1: Simplified Process Flowsheet.................................................................................... 16-1

Figure 16.2: Overall Plant Site Plan ............................................................................................... 16-3

Figure 17.1: Proposed Site Layout ................................................................................................ 17-3

Figure 17.2: Proposed TSF General Arrangement ........................................................................ 17-9

Figure 17.3:Proposed TSF Plan View ......................................................................................... 17-10

Figure 17.4: Typical Section: Tailings Embankment ................................................................... 17-14

Figure 18.1: Average Copper Price as at November 2014 ............................................................ 18-2

Figure 18.2: Average Gold Price as at November 2014 ................................................................ 18-3

Figure 19.1: Protected and Environmentally Sensitive Areas ....................................................... 19-3

Figure 19.2: Land Use Types ........................................................................................................ 19-4

Figure 19.3: Land Use Capability Classes ..................................................................................... 19-5

Figure 19.4: Watershed Catchments in the Region ....................................................................... 19-7

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Figure 19.5: Watershed Sub-catchments and Surface Water Monitoring Points .......................... 19-8

Figure 19.6: Groundwater Monitoring Locations ......................................................................... 19-10

Figure 19.7: Depth to Groundwater in the Open Pit Area ............................................................ 19-11

Figure 19.8: Groundwater Elevations in the Open Pit Area ......................................................... 19-12

Figure 19.9: Community Water Supply Infrastructure ................................................................. 19-14

Figure 19.10: Halilağa Mineral Licenses ...................................................................................... 19-23

Figure 19.11: EIA Permitting Process Flowchart ......................................................................... 19-25

Figure 20.1: Pre-Production Capital Costs .................................................................................... 20-5

Figure 20.2: Sustaining & Closure Capital Costs .......................................................................... 20-5

Figure 20.3: Life of Mine Operating Costs ................................................................................... 20-22

Figure 20.4: Chart of Processing OPEX Estimate by Component .............................................. 20-25

Figure 21.1: Net Revenues by Metal ............................................................................................. 21-3

Figure 21.2: Flotation Circuit Revenues ........................................................................................ 21-4

Figure 21.3: Au Recovered by Process ......................................................................................... 21-4

Figure 21.4: Au Recovered by Production Year ............................................................................ 21-5

Figure 21.5: Royalty Payments by Year ........................................................................................ 21-6

Figure 21.6: Annual After-Tax Cash Flows .................................................................................... 21-9

Figure 21.7: Cash Flow Model ....................................................................................................... 21-1

Figure 21.8: After-Tax Sensitivity Graph for Base Case Results................................................... 21-3

Figure 22.1: Location of Halilağa and Adjacent Properties ........................................................... 22-1

Appendices

Appendix A: QP Certificates

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Effective Date: December 20, 2014 xiv

Important Notice

This report was prepared as a National Instrument 43-101 Technical Report for Pilot Gold Inc.(“Pilot Gold”) by JDS Energy & Mining Inc. (“JDS”). The quality of information, conclusions, andestimates contained herein are consistent with the level of effort involved in JDS’s services, basedon: i) information available at the time of preparation, ii) data supplied by outside sources, and iii)the assumptions, conditions, and qualifications set forth in this report. This report is intended foruse by Pilot Gold subject to the terms and conditions of its contract with JDS and relevant securitieslegislation. The contract permits Pilot Gold to file this report as a Technical Report with Canadiansecurities regulatory authorities pursuant to National Instrument 43-101, Standards of Disclosurefor Mineral Projects. Except for the purposes legislated under provincial securities law, any otheruse of this report by any third party is at that party’s sole risk. The responsibility for this disclosure

remains with Pilot Gold. The user of this document should ensure that this is the most recentTechnical Report for the property as it is not valid if a new Technical Report has been issued.

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Effective Date: December 20, 2014 1-15

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1 EXECUTIVE SUMMARY

1.1 Introduction

JDS Energy & Mining Inc. (JDS) was commissioned by Pilot Gold Inc. (Pilot Gold), Joint Venture

partner with Teck Madencilik Sanayi Ticaret A.Ş., (Teck Madencilik) ., to carry out a revised

preliminary economic assessment (PEA) pursuant to Canadian Securities Administrators’ National

Instrument 43-101 and Form 43-101F1 standards (collectively, NI 43-101) of the Halilağa copper-

gold deposit (the Halilağa Project).

This engineering study is an update from the SRK Consulting (Canada) Inc. (SRK) 2012 PEA

Technical Report for the Halilağa Project (SRK 2012). Several sections from the SRK 2012 report

have been used in this report with updates included as appropriate.

The purpose of producing this PEA (JDS 2014) is to update the SRK 2012 study with the following

changes:

Updated mineral resource estimate;

Mill throughput of 25,000 tonnes per day (t/d). SRK 2012 assumed 50,000 t/d;

Revised tailings management plan;

Addition of a gold leaching circuit for cleaner tails to improve the overall recovery of gold;

Revised capital and operating costs;

Updated closure strategy;

Revised tax calculations as per advice from Ernst & Young;

Modification of royalty calculations to current Turkish law; and

Use of a mining contractor.

This PEA includes the use of inferred mineral resources that are considered too speculative

geologically to have the economic considerations applied to them that would enable them to be

categorized as mineral reserves. This PEA, by definition is preliminary in nature and because of

the use of inferred resources in the mine plan there is no certainty that the economic results shown

will be realized.

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1.2 Project Concept

The proposed Halilağa Project concept is to develop the copper-gold deposit with open pit (OP)mining and conventional milling, flotation concentration and carbon-in-leach (CIL) methods

producing both a copper-gold concentrate and gold doré. The production rate was selected to be

25,000 t/d with approximately 124 million tonnes (Mt) of mineralized material planned to be mined

and processed during the project life. The overall strip ratio (the ratio of waste rock to mineralized

rock) of the mine is approximately 1.3:1 and the average life-of-mine (LOM) grade of the plant feed

is estimated to be 0.34% copper and 0.34 grams/tonne (g/t) of gold.

It is proposed that the copper-gold concentrate would be transported by trucks to the Port of

Bandırma and then shipped to an existing smelter in Europe or the Far East.

The project site is undeveloped but has road access and electrical grid power supply. Labour

supply and industrial service providers are available in the immediate region.The proposed mine production life is 14 years. In addition, it is anticipated that there would be a

4-year to 5-year feasibility, permitting and pre-production construction period as well as a

reclamation period at the end of the mine life.

1.3 Property Description, Location and Ownership

The Halilağa Project is located about 40 km southeast of Çanakkale between the villages of

Halilağa and Muratlar on the Biga Peninsula in Northwestern Turkey. The Biga Peninsula has

fertile soil and a Mediterranean climate with mild, wet winters and hot, dry summers. Maximum

daily temperatures average 30° Celsius (C) in July and August while January, the coldest month,

averages highs of 9o C and lows of 1o C. The annual rainfall for the Çanakkale region is

approximately 63 cm, generally falling as mixed rain and snow in late fall and winter.

In 2007, Teck Resources Ltd. (Teck) (60%) and a subsidiary of Fronteer Gold Inc. (Fronteer) (40%)

formed a Joint Venture company called Truva Bakır Maden İşletmeleri A.Ş. (Truva Bakır) which

has ownership or beneficial interest in 100% of the licenses that comprise the Halilağa Project.

Fronteer’s interest in the Halilağa Project was subsequently transferred to Pilot Gold in April 2011.

BakırBakırThe main area of interest is the Kestane porphyry copper-gold zone located at 483200E,

4419200N UTM Central meridian 27 (ED50 datum). The Halilağa Project consists of 14 licenses

covering 8,866.18 ha. Thirteen licenses are directly held by Truva Bak ır. One license is held by a

subsidiary of Teck, Teck Madencilik Sanayi Ticaret A.Ş. (Teck Madencilik) for the benefit of Truva

Bakır. Five licenses are “Operation-Type”, seven licenses are pending conversion from

“Exploration-Type” to “Operation-Type” and two licenses remain as “Exploration-Type”. The mainlicense hosting the Central Zone at the Kestane porphyry has been converted to an operation

license that is valid until May 2019 and readily renewable thereafter.

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1.4 Geology and Mineralization

The geology of Halilağa is characterized by various lithological associations made up of: (1)Paleozoic and early Mesozoic basement metamorphic rocks; (2) Permian and Mesozoic

sedimentary and ophiolitic rocks; (3) Tertiary volcanic and intrusive rocks; and (4) Neogene

sedimentary rocks. Older rocks are affected by several collisional orogenic events. Tertiary rocks

record mainly brittle extensional and transtensional deformation.

The property is located in a district with significant potential for both porphyry copper-gold and

high-sulphidation-style gold deposits. Investigation of geological relationships in the field along with

geochemical and geochronological data suggest that magmatic activity began with the intrusion of

granitoids, coeval with an early phase of volcanic activity, and was superseded by a second phase

of volcanism. The mineralized intrusives are thought to be part of a suite of 26 to 38 million year old

(Ma) quartz monzonites and granodiorites.

The Halilağa district is mainly underlain by Oligo-Miocene volcanic and-sedimentary rocks

overlying a basement of schist and carbonate rocks which outcrop to the southeast of the Bak ırlik

area. Granodioritic stocks intrude the basement rocks, giving rise to metasomatism and

skarnification. The Kestane porphyry stock, the main host rock to the Halilağa porphyry system,

has a pronounced hornfelsed halo.

The Halilağa Project is centered on a large mineralized system containing porphyry copper-gold,

skarn and high-sulphidation gold mineralization with related alteration assemblages extending over

an area of more than 4 km by 2 km. The Kestane porphyry stock is characterized by potassic

alteration (K-feldspar-biotite-magnetite) overprinted by phyllic (quartz-sericite-pyrite) alteration.

Most quartz veins are B-type that average 5% of the rock by volume. A-type veinlets are rare but

locally present. Chalcopyrite and pyrite are the dominant sulphide minerals. The highest gold and

copper grades in drill core are associated with early biotite + magnetite + chalcopyrite associated

with dense array of A- and B-type quartz veins, overprinted by phyllic alteration. In places the

mineralization has been subject to near surface oxidation and leaching to form supergene

chalcocite.

At higher elevations to the south of Kestane, the Kunk-Kumlugedik hilltops are characterized by

advanced argillic alteration (quartz-alunite-dickite) surrounded by zones of argillic and propylitic

alteration. Skarn-related alteration (magnetite-epidote) is located to the southeast of Kestane in

the Bakırlik and Bostanlikbasi areas.

The geology of the Halilağa area is affected by post mineral faults of the North Anatolian Fault

System characterized by ENE-WSW strike-slip faults, with subsidiary WNW-ESE striking faults.

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1.5 Exploration and Drilling

In 2006-2007, a total of 23 holes (including five abandoned holes) totaling 6,346 m werecompleted. Most of the holes targeted the Kestane porphyry and intersected porphyry-style copper-

gold mineralization with economic grades, as shown by discovery drill hole HD-01, which

intersected 1.03 g/t gold and 1.03% copper over 105.4 m. A 25 m thick chalcocite blanket

averaging approximately 2% copper was also intersected close to the surface in holes HD-01, HD-

02, HD-04, and HD-14.

In 2008, the Bakırlik skarn zone (4 km ESE of Kestane) was the major focus of the drilling program

due to the similarity of its ground magnetic signature to the Kestane area. A total of 20 diamond

holes totaling 4,051 m were completed during that period. Holes HD-21 and HD-25 intersected

narrow zones of skarn mineralization with high-grade copper + gold ± silver values. Detailed

prospecting and systematic rock chip sampling of altered volcanic rocks at Kumlugedik Hill, the

northern limb of Kunk Hill, and Madenderesi targets returned with anomalous gold values thatranged from 0.2 g/t to 2.0 g/t.

In 2009, a total of 18 holes (including four abandoned holes), totaling 5,670 m (excluding the 247m

of hole HD-42D, a deviated hole) were completed at Kestane.

In 2010, the program was designed to continue grid-drilling the Central Zone (the main area of

porphyry copper-gold mineralization). A total of 25 holes (20 diamond and five RC) totaling 9,076.6

m (including 14 abandoned holes) were completed.

In addition to drilling, Induced Polarization (“IP”) geophysical surveys were carried out in 2009 and

2010 that highlighted deep chargeability targets 1 km west of Kestane, and also a chargeability

target at Madeneresi.

The 2011 program focused on extending the mineralization and acquiring data sufficient to produce

an NI 43-101 resource estimate, which was initially presented in a technical report on March 23,

2012 (Gray and Kirkham 2012). Significant intersections were encountered, including a zone of

significant grade corresponding to the near-surface chalcocite blanket encountered in 2007 drilling.

A total of 44 holes (including four abandoned holes) totaling 19,599 m were completed.

A series of north-south geological sections were constructed every 100 m through the deposit.

Sectional interpretations now indicate two E-W-trending normal faults bounding the porphyry

mineralization to the north and south, creating a mineralized horst at the centre.

In 2012, a total of 7,483.5 m (including 563 m of abandoned drill metres) of drilling was carried out in

25 diamond holes (including six of abandoned) in order to 1) convert Inferred mineral resources to

indicated mineral resources; and 2) to define the southern and northern limits of the mineralized body.

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1.6 Mineral Processing and Metallurgical Testing

The Halilağa deposit is classified as a copper-gold porphyry system. The Halilağa process plantand associated service facilities are proposed to process 25,000 t/d of run of mine (ROM) material,

producing copper concentrate for sale into the commercial smelter market and doré bullion for

processing at a precious metal refinery. The proposed process plant design includes crushing and

grinding, rougher and cleaner flotation, regrinding, and dewatering of copper concentrate, cleaner

tails CIL and cyanide detoxification ahead of tailings disposal. It was assumed that copper

concentrate would be trucked from site to the Port of Bandirma and doré bars would be shipped to

a precious metals refinery for processing and sale.

The existing test work as referenced in Section 13.2.1 considered suitable for this level of study

and the comminution data are considered adequate for a conceptual milling circuit design. The

design of the processing circuits is based on test work data in conjunction with reasonable

assumptions based on typical industry values when insufficient data were available.

The Halilağa mineralized material is of moderate competency and hardness, and amenable to

grinding in a conventional semi-autogenous grinding (SAG) and ball milling circuit. The copper

mineralogy is fine grained and test work indicates a requirement to re-grind to a fine particle size to

achieve adequate liberation for cleaner flotation. Grind sizes of 80% passing (P80) 150 microns (µ)

and 20 µ were assumed for the primary grind and re-grind respectively.

Flotation test work recoveries from lock-cycle and open circuit tests were modeled to produce

recovery curves for copper and gold. The LOM recoveries were modeled at 88.2% and 58.4% for

copper and gold respectively. An additional 15% of the gold head grade would be recovered

through the cleaner tail leach circuit increasing overall gold recovery to 73.4%.

1.7 Mineral Resource and Mineral Reserve Estimates

This resource estimate is an update of the initial Halilağa resource completed, and documented in

an NI 43-101 Technical Report, in March 2012 (the “2012 Resource”). This revised mineral

resource estimate is based on assay data available as of July 4, 2013. The geologic model used

for this resource was prepared by Teck staff and is conceptually the same as that used for the

initial resource with the addition of an altered porphyry unit recognized during the 2012 fieldwork.

Geologic control for estimation was based on rock type as well as structural zonation on the flanks

of the porphyry unit, as it was for the 2012 Resource.

The initial resource was estimated by inverse distance interpolation; this update was estimated

geostatistically by ordinary kriging. Copper, gold and molybdenum grades were estimated using

2.0 m composited drill data. The revised resource is tabulated within the same optimized pit shellas was generated and used for the 2012 Resource, as the optimization assumptions are still valid.

The impact of drilling since the 2012 resource has been to increase confidence as reflected by the

increase in the Indicated Mineral Resource as a portion of the total resource. Table 1.1 compares

the 2014 updated sulphide resource with the initially reported numbers; the 0.43 g/t AuEq cut-off

approximately corresponds to the 0.2% CuEq cut-off used in the 2012 disclosure.

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Table 1.1: Comparison to Initial Estimate at 0.43 g/t AuEq Cut-off

ResourceModel

Indicated InferredTonnes

(Millions)Cu(%)

Au(g/t)

Mo(%)

AuEq(g/t)*

Tonnes(Millions)

Cu(%)

Au(g/t)

Mo(%)

AuEq(g/t)*

Updated (2014) 182.7 0.27 0.30 0.0057 0.90 178.7 0.23 0.24 0.0087 0.77

Initial (2012) 168.8 0.30 0.31 0.0054 0.97 199.6 0.23 0.26 0.0067 0.78

Difference +8% -10% -4% +6% -8% -10% -1% -7% +30% -2%

*Gold equivalent (AuEq) grades were calculated using the following parameters: Cu price and recovery of $2.90/lb and 90% Au price and recovery of $1200/oz and 70% Mo price and recovery of $12.50/lb and 50%

Source: Gray 2014

Inferred resources were used in the LOM plan with Inferred resources representing 31% of the

material planned for processing.

The reader is cautioned that mineral resources that are not mineral reserves do not have

demonstrated economic viability. There is no certainty that all or any part of the mineral

resources would be converted into mineral reserves. Mineral reserves can only be

estimated as a result of an economic evaluation as part of a preliminary feasibility study

(PFS) or a feasibility study (FS) of a mineral project. Inferred resources are considered to be

too speculative geologically to have the economic considerations applied to them that

would enable them to be characterized as mineral reserves. Accordingly, at the present

level of development, there are no mineral reserves at the Halilağa Project.

1.8 Mine Development and Operations

The Halilağa deposit is amenable for development as an open pit (OP) mine. Mining of the deposit

is planned to produce a total of 124 Mt of processing plant feed and 158 Mt of waste (1.3:1 overall

strip ratio) over a 14 year mine life. The current LOM plan focuses on achieving consistent plant

feed production rates, and early mining of higher grade material, as well as balancing grade and

strip ratios.

The mine design process for the deposit commenced with the development of CAE Mining NPV

Scheduler (NPVS) OP optimization input parameters. These parameters included estimates of

metal price, mining dilution, process recovery, offsite costs, geotechnical constraints (slope angles)and royalties (see Table 1.2). The resource model was based on a 20 m by 20 m by 10 m block

size.

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Table 1.2: Mine Planning Optimization Input Parameters*

Item Unit Values

Metal Prices

Gold US$/oz 1,250

Copper US$/lb 3.00

Recovery to Cu Concentrate

Gold % var.w/ Au grade

Copper % var.w/ Cu grade

Recovery CIL

Gold (cleaner tails CIL) % 15

Cu Concentrate Grade (“conc.”)

Gold g/t var.w/ Au and Cu grade

Copper % 30

Moisture content 8

TCRC and Smelter PayablesGold in Dore % 99

Gold in Cu conc % 96

Gold deduction in Cu conc. g/t in conc 1

Copper in Cu conc % 96

Cu conc treatment $/dmt conc 75.00

Cu refining charge $/lb pay Cu 0.075

Au refining charge $/oz pay Au 7.00

Transport, marketing, ins, etc. $/dmt conc 62.7

Other Parameters

Grade factor (variable) % 95

Royalties % 4

Pit Slope Angles overall degrees 36 to 48

Dilution % 5Mining recovery % 100

Strip ratio (est.) t:t 1.3

Internal NSR cut-off $/t 8.97

External NSR cut-off (est.) $/t 13.8

Processing rate tpd milled 25,000

Operating Costs

O/P Waste mining Cost $/waste tonne 2.00

OP Mineralized material Mining Cost $/mill feed tonne 2.00

OP Processing and G&A Cost $/milled tonne 8.54

The OP mineable resources are reported at an internal cut-off value of $8.97/t based on input parameters above.

*The values in this table vary slightly from those used in the economic model as parameters were further refined in

the economic model as the project progressed. The differences are not considered material to pit shape definition.

Source: JDS 2014

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CAE Mining’s NPV Scheduler software was used to determine the optimal mining shells with the

assumed overall slope angles shown in the previous table. Preliminary mining phases were

selected and preliminary mine planning and scheduling was then conducted on these selectedoptimal shells. The mineable resources for the Halilağa deposit are presented in Table 1.3.

Both Indicated and Inferred resources were used in the LOM plan of which, Indicated resources

represent 69% (86 Mt) of the material planned to be processed..

Table 1.3: PEA Proposed Mine Plan Summary

Description Unit Value

Mine Production Life yr 14

Process Feed Material Mt 124

Diluted Copper Grade (mill head grade) % 0.34

Contained copper Mlbs 920

Diluted Gold grade (mill head grade) g/t 0.34

Contained gold koz 1,357

Waste Mt 158

Total material Mt 282

Strip ratio t:t 1.3

Source: JDS 2014

The mining sequence was divided into a number of stages designed to maximize grade, reduce

pre-stripping requirements in the early years and, maintain the plant at full production capacity. The

LOM production schedule is shown in Table 1.4.

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HALILA

Effective Da

.

Table 1.4

Item

Mineralize

Gold Feed

Contained

Copper Fe

Contained

Waste Mat

Total Mate

Strip Ratio

Total Mate

Source: J

Figure 1.

Source: J

A PRO JECT

te: December 2

: LOM Prod

Material

Grade

Gold

ed Grade

Copper

erial

rial

rial Mined

S 2014

1: LOM Prod

S 2014

– PEA TEC

, 2014

ction Sche

uction Sche

HNICAL REP

ule

Unit

Mt

g/t

koz

%

Mlbs

Mt

Mt

t:t

tpd

dule

OR T

Total

124.3

0.34

1,357

0.34

920

157.6

281.9

1.3

-1

1

0.64

21

0.95

21

8.7

9.7

8.7

26,486

1

8.1

0.46

121

0.73

131

9.3

17.4

1.1

47,712

2

9.1

0.46

134

0.57

116

13.8

23

1.5

62,918

3

9.1

0.41

120

0.38

76

14.3

23.5

1.6

64,321

4

9.1

0.36

107

0.31

63

14.7

23.8

1.6

65,317

5

9.1

0.33

96

0.28

56

14.7

23.8

1.6

5,210 6

Y

6

9.1

.32 0

95

.28 0

57

14.7 1

3.9 2

1.6

5,371 65

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1.9 Waste Management

Waste rock from mining operations is planned to be deposited in an engineered rock storagefacility (RSF) immediately adjacent to the proposed open pit. The RSF is designed to hold a total of

160 Mt of material.

Tailings from the process plant are proposed to be deposited in a Tailings Storage Facility (TSF).

The TSF design consists of two rock fill embankments with a fully lined containment area. The

starter embankment is designed to be contained in the initial larger valley, while an additional

smaller valley would be required to contain the ultimate facility. Seepage collection ponds would be

constructed downstream of the impoundments.

The upstream side of the embankment is designed to be lined with an HDPE liner. The liners within

the facility would be placed on a protective bedding layer.

1.10 Project Infrastructure

The site has been configured for optimum construction access and operational efficiency. Primary

buildings are designed to allow easy access from the existing mine access road and to utilize

existing topography to minimize bulk earthworks. The process plant is planned to be located in

close but safe proximity to the pit and at an elevation that facilitates the conveying of mineralized

rock. Existing roads would be upgraded and reused wherever possible.

The TSF would be accessed using a new all-weather gravel road. A short side road off the tailings

access road is planned to be used as an entrance to the explosives storage facility.

There is an existing 154 kV power transmission line owned by the Turkish Electricity Transmission

Corporation (TEIAS) that crosses the project site and specifically the planned open pit. This highvoltage line transmits electrical power from the Ҫan power plant to the towns of Ezine, Ayvacıkand

Bayramiç, which are 30 km to 60 km to the west-southwest of the project site. This existing line

would have to be re-routed around the project site prior to open pit mining but could potentially be

used during construction. A preliminary re-routing of the line to the south of the project was

designed and would require 9 km of new line to avoid the open pit and RSF. It is estimated that the

site’s power demands would be approximately 30 MW and the project concept includes a new

approximately 25 km long, 154 kV transmission line from the Ҫan area.

Water supply to the processing facilities are planned to come from the collection of surface water

and use of the abandoned, flooded coal pits (water storage pits) in the immediate vicinity.

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1.11 Environmental Considerations

An Environmental Baseline Study (EBS) was conducted by SRK Dan ışmanlık ve Mühendislik A.Ş.(SRK Turkey) in 2011-2012 in the Halilağa exploration program surroundings (SRK, 2012). Due to

the early stages of the program, project information was not sufficient to conduct a comprehensive

EBS. Therefore, a limited review of environmental resources was conducted in order to obtain a

preliminary insight about the existing conditions and covered the protected areas, land use and

soils, hydrology, hydrogeology, community water resources, water quality, geochemical

assessments, and biodiversity. There does not appear to be any environmental or permitting fatal

flaws and the project site has no residents.

1.12 Capital Costs

The estimated capital costs (CAPEX) include mine pre-stripping, mine development, site

preparation, process plant, first fills, buildings, utilities, road works, etc. as summarized in Table

1.5. The estimates are considered to have an overall accuracy of ±30%.

Table 1.5: Capital Cost Estimate

Capital CostPre-Production

US$M

Sustaining/

Closure

US$M

LOM Total

US$M

Capitalized Mining Costs 17.9 0.0 17.9

Contractor Mobilization/Demobilization 1.0 1.0 2.0

Mining 0.6 0.0 0.6

Site Development 5.5 0.0 5.5

Process Plant 131.6 0.0 131.6

On-Site Infrastructure 29.6 0.0 29.6

Tailings Management 25.0 103.3 128.3

Indirects 37.6 0.0 37.6

EPCM 25.3 0.0 25.3

Owner's Costs 6.4 0.0 6.4

Sustaining 0.0 15.8 15.8

Closure 0.0 50.2 50.2

Subtotal 280.6 170.3 450.8

Contingency 65.4 42.3 107.7

Total Capital Cost 346.0 212.6 558.5 Source: JDS 2014

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1.13 Operating Costs

The operating expenditure (OPEX) estimate has been developed from numerous sources withdeterminations being based on fixed and variable components relating to plant throughput and

other plant feed material characteristics. A summary of the operating cost estimates is shown as

unit and total costs in Table 1.6.

Table 1.6: Operating Cost Estimates

Operating Cost US$/t processed LOM US$M

Mining‡ 4.05 503.7

Re-Handle* 0.01 1.0

Processing (incl. Tails) 8.35 1,038.2

G&A 0.70 86.6Total OPEX 13.11 1,629.4

(‡): Excludes capitalized pre-stripping costs(⁰): Based on $1.85/t mined (assuming average LOM 1.3 strip ratio) (*) Re-handle cost amounts to $1/t re-handled. Total material re-handled amounts to 1M tonnes.Source: JDS 2014

1.14 Economics Analysis

An engineering economic model was developed to estimate annual cash flows and project

sensitivities. Pre-tax estimates of project values were prepared for comparative purposes, while

after-tax estimates were developed to approximate the true investment value. It must be noted thatthe tax estimates involve many complex variables that can only be accurately calculated during

operations and, as such, the after-tax results are approximations to represent an indicative value

of the after-tax cash flows of the project.

One metal price scenario was used for the economic analysis and prices were held constant

throughout the mine life, demonstrated in Table 1.7. Metal prices were based on LME closing spot

prices during December 2014.

Table 1.7: Metal Price Assumptions for Economic Analysis

Metal Unit Value

Copper Price US$/lb 2.90

Gold Price US$/oz 1,200

Source: JDS 2014

The results of the economic analysis are presented in Table 1.8.

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The reader is cautioned that this PEA is preliminary in nature and includes the use of inferredmineral resources that are considered too speculative geologically to have the economic

considerations applied to them that would enable them to be categorized as mineral reserves and,as such, there is no certainty that the PEA economics will be realized. The PEA uses 31% inferredmineralized material.

Table 1.8: Summary of Economic Results

Summary of Results Unit Value

Cu Payable LOM M lbs 779.4

Au Payable LOM k oz 924.2

Operating Costs US$/t processed 13.11

Total Capital Costs Incl. Contingency US$M 558.5

Discount Rate % 7.0

Pre-Tax NPV US$M 510.9

Pre-Tax IRR % 45.8

Pre-Tax Payback Years 1.2

After-Tax NPV US$M 473.8

After-Tax IRR % 43.1

After-Tax Payback Years 1.3

Cu Cash Cost‡ US$/Cu lb 2.50

Cu Cash Cost (Net of By-Products)* US$/Cu lb 1.08

Cu Cash Cost (incl. Sustaining Capital)** US$/Cu lb 2.78

Cu Cash Cost (Incl. Sustaining Capital) Net of By-

Products⁰ US$/Cu lb 1.35

‡ Cash Cost = (Treatment Charge + Refining Charges + Royalties + Operating Costs) / Payable Cu lbs* Cash Cost (Net of By Products) = ((Treatment Charge + Refining Charges + Royalties + Operating Costs)-(Payable Au * Au Price))/Payable Cu lbs** Cash Cost (incl. Sustaining Capital) = (Treatment Charge + Refining Charges + Royalties + Operating Costs +Sustaining Capital Costs) / Payable Cu lbs⁰ Cash Cost (incl. Sustaining Capital) Net of By-Products = ((Treatment Charge + Refining Charges + Royalties +Operating Costs + Sustaining Capital)- (Payable Au * Au Price))/Payable Cu lbs

Source: JDS 2014

A sensitivity analysis was performed on the Base Case metal pricing scenarios to determine which

factors most affect the project economics. The sensitivity analysis completed revealed that the

Halilağa Project is most sensitive to changes in F/X rate (based on specific assumptions made inthe sensitivity analysis with respect to the project’s costs and exposure to foreign exchange

fluctuations). The Halilağa Project also showed it was highly sensitive to changes in metal prices

and head grades, followed by changes in operating costs. The project showed least sensitive to

changes in capital costs. Table 1.9 outlines the results of the sensitivity test performed on the

after-tax NPV7% for the Base Case evaluated.

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Table 1.9: After-Tax Sensitivity Test Results

After-Tax NPV7% (US$M)Variable -15% 0% 15%

Metal Price (Combined) 193.2 473.8 754.4

Cu Price 281.4 473.8 666.3

Au Price 385.7 473.8 562.0

Head Grade 201.9 473.8 748.5

Operating Costs 602.4 473.8 345.3

Capital Costs 523.7 473.8 424.0

Source: JDS 2014

1.15 Risks and Opportunities

As with almost all mining ventures, there are a large number of risks and opportunities that can

affect the outcome of the project. Most of these risks and opportunities are based on uncertainty,

such as lack of scientific information (test results, drill results, etc.) or the lack of control over

external factors (metal price, exchange rates, etc.).

Subsequent higher-level engineering studies would be required to further refine these risks and

opportunities, identify new risks and opportunities, and define strategies for risk mitigation or

opportunity implementation.

The main preliminary risks identified for the Halilağa Project are, summarized as follows:

Permit acquisitions;

Stakeholder support;

Reduced metal prices;

Geological interpretation and mineral resource classification (31% of the mineral resources

used in the mine plan are classified as Inferred) and there is no guarantee these resources

can be upgraded to Indicated or Measured;

Increased OPEX and/or CAPEX;

Geotechnical and hydrogeological considerations;

Metal recovery and mineral processing assumptions, including deleterious elements; and

Water supply and the right to use it.

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The following opportunities may improve the project economics:

Exploration potential from under-explored near-by anomalous zones;

Further optimization of the mine plan and production schedule;

Regional exploration and the potential to increase mineral resources;

Processing of the oxide material (currently treated as waste);

Further metallurgical optimization;

Improved metal prices;

Additional tax and investment incentives potentially available to the project;

Possible synergies and economies of scale related to the proximity of other properties

such as TV Tower; and

Reduction in CAPEX and/or OPEX from value engineering.

1.16 Conclusions

Industry standard mining, processing, construction methods and economic evaluation practices

were used to assess the Halilağa Project. There was adequate geological and other pertinent data

available to generate a PEA.

Based on current knowledge and assumptions, the results of this study show that the project has

positive economics (within the very preliminary parameters of a PEA) and could be advanced to

the next level of study by conducting the work indicated in the recommendations section of this

report.While a significant amount of information is still required for a complete assessment of the project,

at this point there do not appear to be any fatal flaws.

The study has achieved its original objective of providing a preliminary review of the potential

economic viability of the Halilağa Project.

1.17 Recommendations

JDS recommends that the project be advanced to the next level of study, a preliminary feasibility

study (PFS). Prior to undertaking the PFS, the potentially mineable resource will have to be drilled

more extensively in an attempt to convert Inferred resources to Indicated resources although there

can be no assurances that this will be successful. After drilling, sampling and assaying, a newresource model will be required. A high-level estimate of the resource drilling and re-estimation

cost us as follows:

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Table 1.10: Cost Estimate for New Mineral Resource Estimate to Support a PFS

Item and DescriptionNew Mineral Resource

Estimate Cost (M$)

Resource Definition Drilling (8,000 m x $160/m) 1.28

Assaying ($40/m average) 0.32

Camp Operations, Trucks, Fuel, Supplies 2.3

Resource Estimation 0.16

Salaries and staff costs 3.2

Condemnation Drilling under Surface Facilities (2,000 m x $160) 0.32

Mineral Resource Estimate 7.58

Source: JDS 2014

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2 INTRODUCTION

2.1 Basis of Technical Report

This Technical Report was compiled by JDS for Pilot Gold Inc. and its subsidiary Pilot Investments Inc,

to summarize the results of the PEA study. Pilot Investments Inc. (40%) and Teck Madencilik (60%) are

the sole shareholders in a Joint Venture company called Truva Bak ır Maden İşletmeleri A.Ş. (Truva

Bakır) which has 100% ownership of the licenses that comprise the Halilağa Project. This report was

prepared following the guidelines of the Canadian Securities Administrators NI 43-101.

The purpose of producing this PEA (JDS 2014) is to update the previous SRK study with the following

changes:

Updated mineral resource estimate;

Mill throughput of 25,000 tonnes per day (t/d). SRK 2012 assumed 50,000 t/d;

Revised operating and capital costs;

Addition of a gold leaching circuit for cleaner tails to improve the overall recovery of gold;

Revised tailings management plan;

Updated closure strategy;

Revised tax calculations as per advice from Ernst & Young;

Modification of royalty calculations to current Turkish law; and

Use of a mining contractor.

The reader is cautioned that the PEA summarized in this technical report is only intended to provide an

initial, high-level review of the project. The PEA mine plan and economic model include the use of a

significant portion of inferred resources (31% of mine plan mineralized tonnage) which are considered

to be too speculative to be used in an economic analysis except as permitted by NI 43-101 for use in

PEA’s. There is no guarantee that Inferred resources can be converted to Indicated or Measured

resources and, as such, there is no guarantee that the project economics described herein will be

achieved.

2.2 Terms of Reference

A previous technical report was prepared for the Halilağa project by SRK in 2012 titled “Resource

Estimate for the Halilağa Copper-Gold Property NI 43-101 Technical Report (SRK 2012)” based on the

mineral resource estimate presented by James Gray, P.Geo and Garth Kirkham, P.Geo. (“Gray and

Kirkham 2012”).

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This PEA technical report summarizes the various technical and economic conditions at the time of

writing. Given the nature of the mining business, these conditions can change significantly and actualresults may be significantly different.

All drill hole and geological information in this report are current to the effective date of the Report.

2.3 Scope of Work

This report summarizes the work carried out by the following Consultants. The scope of work for each

company is listed below.

JDS’s scope of work included:

Overall technical report compilation;

Project setting and history; Mine planning;

Production schedule and potentially mineable resources;

Process design based on metallurgical results;

Site infrastructure, identify proper sites, plant facilities and other ancillary facilities;

Environment, Social Impact and Permitting;

OPEX and CAPEX;

Economic model;

Conclusions, risks and opportunities; and

Recommendations.

SRK’s scope of work included:

PEA-level geotechnical assessment and estimate of appropriate overall pit slope angles;

Environmental, social and permitting considerations; and

Tailings Storage Facility design, associated construction cost.

Gary Simmons’ scope of work included:

Assessment of metallurgical test results.

Garth Kirkham’s scope of work included:

Review of sections 7 to 12, which include geology, mineralization, sampling drilling and QA/QC;

James Gray’s scope of work included:

Geology description; and

Mineral resource estimate.

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2.4 Qualified Person Responsibilities and Site Inspections

The Qualified Persons (QPs) preparing this technical report are specialists in their fields and meet NI

43-101 regulations for QPs.

The results of this PEA are not dependent upon any prior agreements concerning the conclusions to be

reached, nor are there any undisclosed understandings concerning any future business dealings

between Pilot Gold and the QPs. The QPs are being paid a fee for their work in accordance with normal

professional consulting practices.

The following individuals, by virtue of their education, experience and professional association, are

considered QPs as defined in the NI 43-101, and are members in good standing of appropriate

professional associations. The QPs are responsible for the specific sections as follows:

Table 2.1: Independent QP Responsibilities

QP Company RoleReport

Section(s)

Gordon Doerksen, P.Eng. JDS Energy & Mining Inc.Project Sponsor, Costs and

Economics

1-6, 17-27, except17.20, 17.21 and

17.22

Stacy Freudigmann,P.Eng.

JDS Energy & Mining Inc. Project Manager and

Processing16

Dino Pilotto, P.Eng. JDS Energy & Mining Inc. Mining15 (excluding

15.1)

Greg Abrahams, P.Geo. SRK Consulting (Canada) Inc. Open Pit Geotech 15.1

Maritz Rykaart, P.Eng. SRK Consulting (Canada) Inc. Waste Facilities17.20, 17.21 and

17.22

Gary Simmons, MMSA GL Simmons Consulting LLC Metallurgy 13

Garth Kirkham, P.Geo. Kirkham Geosystems Ltd. Geology, Sampling, Drilling 7-12

James Gray, P.Geo. Advantage Geoservices Ltd. Mineral Resource Estimate 14

Source: JDS 2014

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2.5 QP Site Visits

Table 2.2: QP Site Visits

QP Company Site Visit Dates

Gordon Doerksen, P.Eng. JDS Energy & Mining Inc.February 29 – March 1, 2012

April 26 – May 2, 2013April 7 – 14, 2014

Stacy Freudigmann, P.Eng. JDS Energy & Mining Inc. No site visit

Dino Pilotto, P.Eng. JDS Energy & Mining Inc. February 29 – March 1, 2012

Greg Abrahams, P.Geo. SRK Consulting (Canada) Inc. No site visit

Maritz Rykaart, P.Eng. SRK Consulting (Canada) Inc. February 29 – March 1, 2012

Gary Simmons, MMSA GL Simmons Consulting LLC April 26 – May 2, 2013

Garth Kirkham, P.Geo. Kirkham Geosystems Ltd. August 13 – 16, 2011

James Gray, P.Geo. Advantage Geoservices Ltd. No site visit

Gordon Doerksen, P.Eng., Maritz Rykaart, P.Eng., and Dino Pilotto, P.Eng. visited the project site from

February 29 to March 1, 2012. The site visit encompassed an inspection of the core storage area, a

review of selected core and a tour of major centres, local villages and potential open pit, plant site,

water reservoir and rock and tailings storage facility locations.

Gordon Doerksen, P.Eng. and Gary Simmons visited the project site from April 26 to May 2, 2013.

Gordon participated in a follow-up site visit between April 7-14, 2014 to evaluate port facilities, potential

mining contractor options and permitting, power supply and engineering consultants.

Garth Kirkham, P. Geo., visited the property between August 13 and August 16, 2011. The site visit

entailed inspection of the camp, accommodations, core logging facilities, offices, active drill sites,

outcrops, historic drill collars, core storage facilities, core receiving area, core sawing stations, and tours

of major centres and surrounding villages most likely to be affected by any potential mining operation.

On all site visits, QPs were accompanied by senior Pilot Gold and Truva Bakır representatives.

The QPs that did not visit the site relied upon information from the other QPs that did visit the site. The

QPs that did not visit the site, were not required to as there was no information on site pertinent to their

areas of responsibility.

2.6 Sources of Information

Sources of information include data and reports supplied by Pilot Gold as well as external documents

such as SRK’s 2012 PEA, cited throughout the report and referenced in Section 27.

All tables and figures are sourced from JDS in 2014 unless otherwise indicated.

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2.7 Units, Currency and Rounding

Unless otherwise specified or noted, the units used in this technical report are “metric” as per the

International System of Units (SI). Every effort has been made to clearly display the appropriate units

being used throughout this technical report. Currency is in United States dollars (US$ or $) unless

otherwise noted.

This report includes technical information that required subsequent calculations to derive subtotals,

totals and weighted averages. Such calculations inherently involve a degree of rounding and

consequently introduce a margin of error. Where these occur, the QPs do not consider them to be

material.

Frequently used abbreviations and acronyms can be found in Section 26.

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3 RELIANCE ON OTHER EXPERTSTo prepare this report, the QPs used their experience to determine if the information from previous

reports provided by Pilot Gold were suitable and updated information where appropriate. P These

reports, as referenced in Section 27, were exploration reports and data from previous exploration

programs, internal desktop studies, and consultants’ reports, including the NI 43-101 Technical Report

authored by SRK, on behalf of Truva Bakır and issued October 10, 2012.

It is believed that the information, conclusions, and recommendations are accurate and reliable. It is

also believed that the drilling, geological, geochemical, and geophysical data reported by other

companies and government agencies regarding the prospect and its environment is accurate and

reliable and has been performed by competent professionals operating to industry standards and best

practices.

This report was prepared using public and private information provided by Pilot Gold and information

from previous technical reports listed in Section 27 of this report. The current report also relies on the

work and opinions of non-QP experts and non-independent QPs. The QPs believe that the information

provided and relied on for preparation of this report was accurate at the time of reporting, and that the

interpretations and opinions expressed by these individuals are reasonable based on their current

understanding of the deposits. Each contributing QP has made a reasonable effort to verify the

accuracy of the data used to develop this report and takes full responsibility for the information

contained in this report.

Specific non-QP input for this report included assistance from Ernst and Young for Turkish taxation

information and calculations, and from Bora Arpacıoğlu, Principal Environmental Engineer, SRKDanışmanlık ve Mühendislik A.Ş. (SRK Turkey) for environmental, social impact and permitting

information.

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4 PROPERTY DESCRIPTION AND LOCATION

4.1 Property Description

The Halilağa Project is located about 40 km southeast of Çanakkale between the villages of Halilağa

and Muratlar on the Biga Peninsula in Northwestern Turkey (Figure 4.1). The main area of interest is

the Kestane porphyry copper-gold zone located at 483200E, 4419200N UTM Central Meridian 27

(ED50 datum) in the central part of the tenement group.

Figure 4.1: Map of Turkey

Source: Pilot Gold Inc. 2014

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Figure 4.3: License Boarders and General Arrangement of Facilities

Source: Pilot Gold 2014

As shown in Figure 4.3, there are some facilities that currently are planned to extend beyond the

tenement boundaries, for which surface rights would need to be acquired. The goal of further studies

will be to maintain all project facilities within the existing tenements.

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Figure 4.4: Halilağa License Map


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4.3 State Royalties

At time of writing, the Government of Turkey is currently reviewing the State’s Mining Royalties. For this

report the previous 2012 SRK report’s royalties was also used, such that the Government of Turkey

would receive a 2% of Net Smelter Royalty (known as the State’s rights) for base metals and 4% for

precious metals. Because the project uses copper flotation, the copper net smelter royalty is reduced to

1%.

The State’s rights, paid by the license holder, would be distributed to; the Special provincial

Administration of Çanakkale (25%), Turkish Treasury (50%), and Sub-provincial Administration (in this

case Bayramiç and Can as it is in between those sub-provinces) to be used for infrastructure (25%).

The Council of Ministers can apply a maximum 25% discount in the State’s rights rates depending on

the type of mineral, the region of production, and other criteria.

The project is located on State-owned land; therefore an additional 30% of the royalty payment is

required to be paid, increasing the gold royalty to 2.6% and the copper royalty to 1.3%.

Each year the license holder pays the royalty on the last day of June.

4.4 Joint Venture Agreements

In 2002, the Halilağa Project was acquired at auction by Teck Cominco Arama ve Madencilik Sanayi

Ticaret A. Ş. (now Teck Madencilik ). In 2004, Teck Madencilik and Fronteer entered into an option

agreement that covered several properties in the Biga area (including the Halilağa Project) that enabled

Fronteer to acquire 100% interest in the properties subject to certain earn-back rights by Teck

Madencilik which was exercised on November 30, 2006. Subsequently, Teck Madencilik earned a 60%interest in the property by investing $2.5 million during 2007. On December 31, 2009, Teck Madencilik

declined to earn an additional 10% interest in the Halilağa Project.

Teck Madencilik (60%) and Fronteer (40%) formed a Joint Venture company called Truva Bak ır Maden

İşletmeleri A.Ş. which owns, or has beneficial interest in, the licenses that comprise the Halilağa

Project. Fronteer’s interest in the Halilağa Project was subsequently transferred to Pilot Gold Inc. in

April 2011.

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4.5 Mining Rights and Title in Turkey

Truva Bakır’s project activities are required to follow the mining codes as set out within Turkey’s state

and local environmental regulations. Truva Bakır must protect the environment from spills, capture and

dispose of hazardous material including aviation fuel, reclaim disturbed ground, cap drill holes, and

remove all refuse. All of the necessary forest and environmental permits were obtained for the 2011 site

work, including permission for timbering, road construction, drill site construction, and drilling for

exploration.

The Law Regarding Amendments on Mining Law, and other Certain Laws, (such as No. 5995), was

approved by the Grand National Assembly of Republic of Turkey on June 10, 2010, and regulations to

regulate the law released on December 6th, 2010.

On May 25, 2009, Truva Bakır received a renewed Group IV Operation-type minerals license, renewingthe exploration license at the Halilağa Project. In 2012, the operation permit for an open pit clay

operation was acquired for the main Kestane licence before the third anniversary of the operation

license. This permit is valid until 2019 and can be readily renewed. At the same time, a small

underground copper-gold mining project application was submitted, for which approval is pending.

An approved Environmental Impact Assessment (EIA) report and the GSM permit are the final permits

required to acquire the copper-gold Operations Permit, allowing the filed copper-gold related operation

at Halilağa to continue. Mining rights and minerals are exclusively owned by the State. The ownership

of the minerals in Turkey is not subject to the ownership of the relevant land. The State, under the

mining legislation, delegates its rights to explore and operate to individuals or legal entities by issuing

licences for a determined period of time in return for a royalty payment. The licenses for mining rights

are granted to Turkish citizens, legal entities established under Turkish laws, and some authorized

public bodies. Companies established under Turkish law, according to the provisions of the Turkish

Commercial Code, are Turkish companies even if they are established by foreign persons with 100%

foreign capital. Consequently, there is no distinction between the mining rights that may be acquired by

local investors and those that may be acquired by foreign investors, provided that the foreign investors

establish a company in Turkey under Turkish law (Ozkan, 2007).

Under the Turkish Mining Law, mines have been divided into six groups under which the Halilağa

Project falls in Group IV. These groups are subject to different terms and conditions on licensing

principals and procedures. These groups are;

Sand and gravel (Group I); Marble and other similar decorative stones (Group II); Salts in solution that can be obtained from aqueous solutions (Group III); Metal and industrial minerals (Group IV), (the group Halilağa would be classified under); Precious metals and gem stones (Group V); and Radioactive minerals and substances (Group VI).

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The two types of licenses granted for prospecting and operating mines are as follows; (i) exploration

licenses, enabling a holder to carry out prospecting activities in a specific area; (ii) exploitation/operationlicenses, enabling a holder to carry out operational activities (including exploration) within the same

area as stated in the prospecting license. For production (extractive activity) to occur, an operations

permit must also be obtained. An operations permit enables the holder to operate a specific mine as

specified in the Exploitation/Operation license, and as contemplated by an approved EIA report.

Applications to convert from an exploration to an operation-type license must be submitted before the

end of the term of an exploration-type license, and must demonstrate the presence of an economic

resource on the license.

The conversion application includes providing a resource estimate, a conceptual mine plan, a positive

conceptual economic analysis and an initial description of likely environmental impacts. The pre-

requisite to conversion application is the EIA permit, business opening and work permit, and

governmental land use (e.g. forestry, pasture lands etc.) permits. When a license conversion happens,the exploration drilling permits are cancelled and the holder must apply for a new forestry permit to drill

on the project. Each license type is valid for a predetermined period of time and must meet a variety of

requirements in order to remain in good standing, including a requirement to receive a number of

permits from the Government of Turkey’s Mining Affairs General Directorate of the Ministry of Energy

and Natural Resources (the "General Directorate-Mining Affairs").

Applications to renew an exploration-type license, as well as applications to receive, or renew an

operation-type license, are made to the General Directorate-Mining Affairs, and are subject to an

extensive review.

4.6 Environmental Studies, Permitting and Social or Community Impact

In December 2011, Truva Bakır submitted an EIA report for a small-scale copper-gold test-miningunderground scenario (an adit) to the Turkish Ministry of Environment and Urban Planning (the

“Ministry”) in connection with an application to meet the requirements of an operation stage permit for

the principal licenses that comprise the Halilağa Project. in August 2012, Truva Bakır was informed that

the Ministry had been served a legal petition by certain claimants in Turkey to annul the Ministry’s

approval of the EIA report. The petition filed with the Çanakkale Administrative Court (the “Court”)

names the Ministry as the respondent and does not name Truva Bak ır or its shareholders. The petition

also requested suspension of mining (exploitation) activities contemplated within the EIA area by way of

an interim decision to be granted by the Court.

Following discovery and the consequential administrative hearing, on November 20, 2013, the Court

found that the EIA report for Halilağa had been appropriately approved by the Ministry, and concurredthat the report was valid. The Court however, awarded interim injunctions suspending any activities

contemplated in the EIA relating to the designated area contemplated (Licence Number 81802). There

is no impact or restriction on Truva Bakır for planned activities at Halilağa outside of the designated

areas.

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The Court also concluded initially, that notwithstanding the validity of the EIA report, certain additional

analyses should be included in an amended report, including an analysis of the cumulative impactassessment of the disturbances considered in the Halilağa EIA when examined along with all other

contemplated EIA reports submitted in the greater Çanakkale area.

In December 2013, the Ministry appealed the interim injunction, and the Court’s inclusion of a

cumulative impact assessment requirement. The District Administrative Court at Edirne, Turkey,

rejected the Ministry’s appeal on December 30, 2013. An administrative hearing convened on March 7,

2014, to determine if a revised and amended EIA is required. Rulings from the Court led to the

annulment of the existing EIA and required that the EIA’s be resubmitted with a cumulative impact

assessment. The Ministry has appealed the decision.

Truva Bakır is building long-term relationships within surrounding communities by:

Providing information and regular updates to the public regarding exploration activities in the

region;

Providing a safe environment for exploration;

Providing a regular flow of information between project leaders and community groups;

Ensuring all concerns, questions, and feedback from community groups are efficiently and

effectively addressed;

Assessing the environmental and social risks associated with each phase of the project;

Implementing mitigation measures when necessary;

Continuing with the Grievance/Complaint Resolution procedure to give all stakeholders an

opportunity to present a dispute or concern related to exploration activities;

Supporting transparent and fair employment strategies at the local level;

Allocating resources to the study and procurement of social, economic, and educational

infrastructure within those settlements directly affected by the project; and

Supporting community investment focused on education and health to promote local

development, decrease poverty and enhance quality of life.

The goal for all community investment (“CI”) projects is to have the initiative benefit the community as a

whole, rather than benefitting an individual or specific group within the community. In addition to the

immediately-required infrastructure and educational support, CI activities would include long-term,sustainable goals for local communities that generate employment opportunities, improve long-term

health services, and increase the capacity for growth as defined within the business objectives of the

company.

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4.7 Community Water Supply Considerations

Drinking water in the region is supplied from springs flowing out of the hills at higher elevations by using

gravity to transport water. The quantity and the quality of the water obtained from the springs would

potentially be affected once any proposed pit development is initiated. An alternative resource for the

drinking and utilization water for local consumption needs to be identified.

As part of this study, it is comtemplated that a potable water treatment system will be installed

supplemental to the process plant facility. It will be capable of producing potable water at a capacity of

138 litres per minute. As part of the sustainable mine plan, 20 km of buried potable water lines may be

installed in an effort to supply potable water to nearby villages. Further investigation pertaining to

surrounding right-of-ways and ground conditions will need to be completed to determine the validity of

such a network.

4.8 Other Considerations

To the extent known at this time, there are no other risks that may affect access, title or the right or

ability to perform work on the property.

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5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES,

INFRASTRUCTURE, AND PHYSIOGRAPHY

5.1 Accessibility

The Halilağa Project is located 40 km southeast of the city of Çanakkale between the villages ofHalilağa and Muratlar on the Biga Peninsula, Northwestern Turkey. The nearest town of significance isÇan which is 30 road-km north-east of the project. (See Figure 4.2)

Well-maintained, paved roads provide excellent access from the main highways to the project site. The

deposit area is accessed by a series of good forestry roads. Year-round access to the properties for

field exploration is possible, although snow during winter may temporarily limit access by vehicle.

5.2 Climate

The Biga Peninsula has fertile soil and a Mediterranean climate with mild, wet winters and hot, dry

summers. Maximum daily temperatures average 30° C in July and August, while January, the coldest

month, averages highs of 9o C and lows of 1o C. The annual rainfall for the Çanakkale region is

approximately 63 cm, generally falling as mixed rain and snow in late fall and winter. The project

construction and operation would be unencumbered by weather.

5.3 Physiography

The Halilağa Project contains a 4-km-long, E-W-trending topographic high, with the Kestane porphyry

located on the northern flank of the hill (see Figure 5.2). The highest elevations on the property are

approximately 600 m with the Kestane Zone occurring at an elevation of approximately 350 m.

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Figure 5.1 General View of Halilağa (View to the Southeast)

Note: Kestane area and drill roads are in the middle ground crossed by a power line.

Source: Gray and Kirkham 2012

5.4 Vegetation

Vegetation in the area is dominated by scrub oak and various low-lying shrubs as well as pine trees

planted by the Forestry Department. Various grasses also grow in the area and provide grazing for

livestock. Vegetation in higher elevation is predominantly coniferous trees while various crops and

grasses predominate in areas developed for farming.

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5.5 Existing Infrastructure

The Biga Peninsula has excellent infrastructure with power, road, rail and port facilities.

Electric Power

The project site itself has access roads and a 154 kV power transmission line that traverses the

planned open pit. The transmission line is fed from a coal-fired power generation plant in Çan. The

transmission line would have to be relocated but may serve as a source of power during construction.

Water

There are several abandoned, flooded coal pits within 6 km of the proposed plant site and these are

planned to be used as water reservoirs for the project. Bathymetry has not been done on the pits butconservative estimates indicate that they would store sufficient water for the project needs (see Figure

5.3).

Figure 5.2 Flooded Coal Pit near the Project

Source: JDS 2012

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.

Port Facilities

For concentrate shipping, several port options exist in the region. Based on a high-level review of all the

regional ports, this study assumes that concentrate would be trucked to the Port of Bandirma, 140 km

by road from Halilağa. An overview of the Bandirma port facility is shown in Figure 5.3.

The Bandirma port facility was privatized in 2010 and is run by Çelebi Bandirma International Port

Management Inc. (Çelebi). The facility can accommodate Panamax-class ships and has 12 m of water

depth. There are conceptual plans to deepen the port waters to a depth of 14 m. The port has the

ability to load 24 hour/day. The port handled 4 Mt of freight last year. Port management have

expressed a desire to bring the flow up to 11 Mt. Discussions with Çelebi’s Port Manager and CEO

indicated a strong interest in handling potential Halilağa concentrate, with future port upgrade capital

borne by Çelebi.

The Bandirma Port has excellent space, multiple-jetty availability, and a willing owner. Bulk material

storage handling facilities are needed. Bandirma appears to be the best option for the shipment of

Halilağa concentrate. The port facility was toured by Pilot Gold personnel and QP Gord Doerksen of

JDS. The port has adequate space for a concentrate storage and loading facility. Port management

showed interest in providing this service.

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HALILAĞA

Effective Date

.

Figure 5.3:

Source: JDS

Railway

Türkiye C

the Port o

(about 100

it stands n

Other Fac

Teck Mad12 km aw

PRO J ECT

: December 20,

Layout of t

2014

mhuriyeti D

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km distanc

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ilities

ncilik utilizy from the p

PEA TECH

2014

he Port of B

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NICAL REPO

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5-5

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about

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.

Figure 5.4 Core Logging and Storage Facilities

Source: JDS 2014

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.

6 HISTORY TO 2000

This section is taken from SRK 2012 and remains valid.

The following details an outline of the exploration activities on the Halilağa Project that occurred prior to

involvement of Truva Bakır or related parties. Historic exploration activities were focused primarily on

the Halilağa and Pirentepe targets. These targets, along with the Halilağa North target, are now all

within the Halilağa Project and are owned by Truva Bakır.

6.1 1988-1991

The government’s General Directorate of Mineral Research and Exploration of Turkey (MTA) conducted

a regional scale exploration program over the Biga Peninsula between 1988 and 1991. In the vicinity of

the Halilağa village, MTA located zones of silicification and argillic alteration at Halilağa North coveringapproximately 700 m by 150 m and trending NE-SW at Sağluk Tepe. Samples with significant gold

values (> 0.5 g/t Au in rock) originated from the western part of Sağluk Tepe in silica-argillic altered

zones. Southeast of Halilağa at Taskesilen-Kocatas Tepe, a NE-trending silicification zone, measuring

1,000 m long and 300 m wide, was also mapped.

MTA drilled two diamond drill holes totaling 302 m to test a geochemical anomaly identified by rock chip

sampling at Halilağa North. MJTC-16 intersected narrow intervals of gold mineralization and returned

0.58 g/t Au over 13.85 m. MJTC-17 did not intersect any significant mineralization. Core samples were

selectively analyzed for gold, silver, copper, lead, zinc, antimony, mercury, and molybdenum.

6.2 1997

Cominco collected several rock chip samples from silicified outcrops at Halilağa North and at

Kumlugedik Hill area, where numerous gold anomalies have been detected. The highest-grade sample

from Halilağa North contained 1.17 g/t Au and the highest grade sample from Kumlugedik contained

2.2 g/t Au.

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6.3 1998

A total of 293 soil samples were collected from the Kunk-Kumlugedik lithocap by Cominco. The most

anomalous gold in these soil samples highlights the area east of Kumlugedik and Guvemtasi Hills.

6.4 2000

Cominco conducted reconnaissance soil sampling and rock chip sampling. A total of 107 samples were

collected over five N-S soil lines. The assay results returned anomalous gold (> 10 ppb) from the East

and Southwest Limbs of Celdiren Hill.

There are historical adits and a small pit on the property; however, the background and production

history on these workings are undocumented and unknown.

The authors are not aware of any previous mineral resource estimates, reserve estimates, or mineral

production from the property.

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7 GEOLOGICAL SETTING AND MINERALIZATION

7.1 Regional Geology

The Halilağa property is located in the central part of the Biga Peninsula in Western Turkey. The

geology of the peninsula is complex and characterized by various lithological associations made up of:

(1) Paleozoic basement metamorphic rocks; (2) Permian and Mesozoic sedimentary and volcanic rock

units; (3) Tertiary volcanic and intrusive rocks; and (4) Neogene sedimentary rocks. The regional

geology is shown in Figures 7.1 and 7.2.

Paleozoic and early Mesozoic basement metamorphic rocks occur in three distinct lithological

associations, as summarized by Yiğit (2012). These include the Çamlıca metamorphic complex,

Kazdağ Massif, and Permo-Triassic Karakaya complex. The latter comprises two distinct lithological

associations, namely: (a) a strongly deformed greenschist facies metamorphic sequence of metabasitesintercalated with phyllite and marble accompanied by minor amounts of metachert, meta-gabbro and

serpentinite; and (b) a thick series of low grade metamorphic rocks. Metamorphic rocks variably record

Carboniferous, Late Triassic and Oligo-Miocene metamorphic events.

Pre-Cenozoic sedimentary rocks in the Biga Peninsula include (1) Triassic terrigenous to shallow

marine clastic sedimentary rocks; (2) Middle to Upper Jurassic platform-type neritic limestones; (3)

Lower Cretaceous pelagic limestones; and (4) Upper Cretaceous through Paleocene volcanic and

sedimentary rocks comprising accretionary melange and ophiolitic rocks.

Cenozoic sedimentary rocks in the Biga Peninsula can be evaluated in four time-intervals, separated by

disconformities: Maastrichtian–Early Eocene, Middle Eocene–Oligocene, Miocene and Pliocene–

Holocene. Early-Middle Miocene times are characterized by coeval volcanism and sedimentation.

Lacustrine sediments like shale, siltstone and tuffs were deposited in small basins including economic

coal deposits, such as the Çan lignite.

Cenozoic volcano-plutonic rocks dominate the geology of the Biga Peninsula and therefore disguise

older rocks (Figure 7.2). Cenozoic volcanism in the Biga Peninsula started in the Eocene in extensive

areas with mainly andesitic and dacitic, calc-alkaline character and continued to basaltic alkaline

volcanism through Late Miocene. Broadly, volcanism in the Biga Peninsula initiates with Middle Eocene

medium-K calc-alkaline and continues through Oligocene with high-K calc-alkaline character. Early

Miocene volcanism is characterized by high-K to shoshonitic lavas. In the Middle Miocene to Late

Miocene, volcanism shifted to mildly-alkaline and alkaline characters respectively. Geochemistry of the

volcanic rocks suggests increasing amounts of crustal contamination with decreasing subductionsignature during the evolution of magmas from the Eocene through the Early Miocene. Middle to Late

Miocene volcanism gives geochemical signatures indicating decreasing crustal component with an

enriched asthenospheric mantle- derived melt. Cenozoic calc-alkaline volcanism hosts many important

economic deposits of metallic and industrial minerals.

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Small to medium-sized intrusive bodies are exposed throughout the Biga Peninsula. Most of these

intrusions trend either northeast, following the major tectonic grain of the peninsula, or east-northeast,cutting the major tectonic grain. The main Cenozoic intrusions in the Biga Peninsula show calc-alkaline

character with compositions ranging from granite to quartz diorite. Young granitoids in the Biga

Peninsula generally are the products of Eocene to Oligo-Miocene plutonism. The radiometric ages from

Cenozoic intrusions range from 52.7±1.9Ma (Karabiga Pluton) to 18.8±1.3 Ma (Yenice Pluton). Age

dates from plutonic rocks collectively suggest a younging age from north to south for plutonism in the

Biga Peninsula, from Late Cretaceous to Early Miocene (Yiğit, 2012).

The structural geology of the Biga Peninsula is intricate. Pre-Cenozoic structures are dominated by

thrust faults associated with ophiolitic rocks. The oldest thrust faults are related to metamorphosed

ophiolites in the Kazdağ Group and melanges in Hodul unit of the Karakaya complex. Cenozoic

structural features are characterized by detachment faulting related to exhumation and core-complex

development of Kazdağ Massif in Oligo-Miocene, and strike-slip faulting started in Early Miocenerelated to development of the North Anatolian Fault Zone (NAFZ).

Neotectonic activity is dominated by dextral-strike slip faulting as well as north-south extension. Based

on interpretation of the geological maps, LANDSAT and ASTER images incorporated with field

observations, NE-, E- and NW trending faults form three major groups in the Biga Peninsula (Yiğit,

2012). The NE- and NW-trending faults are likely conjugate Riedel shears. The most prominent faults

are NE-trending dextral-strike slip systems (~060) (Figure 7.1), related to the western extension of the

NAFZ, which create pull-apart basins that control Oligo-Miocene sedimentation and volcanic activity.

This current tectonic regime forms NE-trending basins and ranges, and forms the northwestern

boundary of volcanic rocks in the Biga Peninsula.

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HALILAĞA

Effective Date

.

Figure 7.1:

Source: Yiği

PRO J ECT

: December 20,

Regional Di

, 2012

PEA TECH

2014

stribution o

NICAL REPO

Metamorp

RT

ic Rocks in the Biga Pe insula

7-3

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HALILAĞA

Effective Date

.

Figure 7.2:

Source: Yiği

PRO J ECT

: December 20,

Regional D

, 2012

PEA TECH

2014

istribution o

NICAL REPO

Volcanic a

RT

d Igneous ocks in the Biga Penin ula

7-4

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7.2 Property Geology

The Halilağa area is mainly underlain by Oligo-Miocene volcanic and sedimentary rocks, overlying a

basement consisting of schists and carbonate rocks that outcrop to the southeast of the Bak ırlik area.

The Kestane porphyry stock was emplaced into the volcano-sedimentary sequence and produced

hornfels halo around its margins. The geology and alteration of the area with units listed from youngest

to oldest are as follows:

Colluvium

The Halilağa property area is extensively covered by colluvium, particularly on the steeper slopes of the

Kunk Tepe, Guvemtasi Tepe, Tasyatak Tepe, and Kumlugedik Tepe. Note that a tepe is defined as a

hill. In road cuts, this colluvial cover can be up to 3 m thick, limiting the total exposure of bedrock

outcrop across the property. The geology and alteration maps do not show this cover and areinterpretations of the underlying bedrock rocks.

Polymictic Conglomerate

Polymictic conglomerates are widely distributed, particularly in the NW of the property. The

conglomerate consists of rounded to sub-rounded clasts of silicified tuff, porphyritic volcanics, and vein

quartz in a chlorite/epidote altered matrix. The timing relationships are not clear. The quartz clasts are

possibly sourced from granites to the SE of Halilağa.

Volcanic Cover Unit

Loose, un-mineralized andesitic/basaltic volcanics ± brownish sediments are distributed to the N and

NE of the Kestane porphyry. The fault contact (110/65N) between the cover unit and quartz feldspar

porphyry (QFP) was observed in some holes and in a trench wall located at HD-38 and HD-47.

Quartz Feldspar Porphyry (QFP Kestane stock)

Quartz-feldspar porphyry intrusive rocks are well-exposed in the Kestane Zone (Central Zone) as

shown in Figure 7.5. This unit consists of a very crowded porphyry with phenocrysts predominantly

made up of white feldspar with lesser amounts of quartz and hornblende, with accessory biotite.

Volcanic/Subvolcanic Rocks

The dominant units in the mapped area are andesitic/dacitic volcanic and sub volcanic rock and a

feldspar-hornblende intermediate felsic subvolcanic porphyry unit with significant quartz. Part of theandesite unit includes tuffs implying, at least in part, an extrusive origin. Where the tuffs can be mapped

over larger areas they have been identified on the geological map as a separate unit.

Inferred discontinuities between the tuff units and the andesite/tuff sequence, and the feldspar-

hornblende-quartz volcanic unit as shown on the maps define a NW-SE-trending structural lineament,

along which the Kestane stock was emplaced.

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Andesitic Tuffs

This unit forms part of the same volcanic pile as the andesite flows but are mapped as a separate unit

where sufficient outcrop exists.

Quartzites and Carbonates

Basement clastic and carbonate rocks in the vicinity of historic mining activity at Kumlugedik Tepe have

been propylitically altered (chlorite, epidote, actinolite, and quartz).

Schistose Basic Volcanics and Sediments

Scattered float on Bakirlik Tepe of un-mineralized schistose basic volcanic and sedimentary rocks

representing the metamorphic basement rocks of the Biga Peninsula, although no outcrops have been

observed.

Figure 7.5 shows the current understanding of the local geology.

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HALILAĞA P RO JE CT – P E A TE CHNICAL RE P O RT


.

Figure 7.3: Halilağa Local Geology


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Figure 7.4: Halilağa Alteration Map


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7-9

.

Alteration

The Halilağa alteration system covers a large area of more than 4 km x 2 km and displays all porphyry

alteration types as well as related epithermal and skarn alteration facies. The Kestane porphyry

outcrops are characterized by potassic overprinted by phyllic alteration whereas Kunk-Kumlugedik Hill-

tops are characterized by silicification surrounded by advanced argillic to argillic and distal propylitic

alteration (see Figure 6.6). Skarn-related alteration is located around the Bak ırlik and Bostanlikbasi

areas.

The details of the alteration styles are described here below.

Propylitic/Sub propylitic

Basement units in the vicinity of historic mining activity at Kumlugedik Tepe have been propyliticallyaltered (chlorite, epidote, actinolite, and quartz) as determined by thin section analysis. Sub-propylitic

alteration of the volcanic and sub-volcanic rocks is typically weak to moderate in intensity. The

occurrence of chlorite, smectite, and varying degrees of carbonate and pyrite define this alteration zone

at Halilağa, which grades out into unaltered rocks to the NE and SW.

Argillic

A broad zone of argillic-altered andesitic volcanic rocks and tuffs are developed around the advanced

argillic alteration zones developed along the Kunk Tepe-Guvemtasi Tepe-Tasyatak Tepe ridge line.

Typically the alteration preserves texture with feldspars altering to kaolin and smectites, with the

groundmass replaced by varying degrees of silica. The best exposures of this style of alteration are

found in excavations for spring water on the NE flanks of Guvemtasi Tepe.

Advanced Argillic (Quartz-Alunite)

Advanced argillic alteration along the NW-trending ridge lines of Guvemtasi Tepe and Tasyatak Tepe

ranges from quartz-alunite altered andesites at Guvemtasi, to near complete silica replacement of tuffs

and volcanics at Tasyatak Tepe. This zone is inferred to represent the distal outflow of acidic fluids from

the porphyry source. Previous rock chip sampling along this zone returned anomalous (< 1 ppm) gold.

Soil sampling returned localized anomalies with respect to gold, bismuth, and mercury.

Commonly, the quartz-alunite alteration is cut by mm-scale goethite-hematite veinlets generally trending

NNE which are well-exposed in saw-cut channels.

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7-10

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Silica-pyrite

Texture preservative and pervasive silica-pyrite alteration is mapped in the SW portion of the property

within andesites and a felsic porphyritic volcanics, probably of dacite composition. A surficial argillic

alteration zone has developed on weathered outcrops due to the oxidation of pyrite. The silica-pyrite

alteration is inferred to be an early stage of alteration resulting from moderate temperature sulphur-rich

and metal-poor fluids derived from a porphyry intrusion. Rock sampling has been limited in this area;

however, soil sampling shows that this area is not anomalous with respect to gold. Later higher-

temperatures and more saline fluids resulted in gold deposition associated with the advanced argillic

alteration developed along NW-trending zones as described above. In this model, the advanced argillic

alteration post-dates and cuts the earlier silica-pyrite alteration.

Silicic

Texturally-destructive silica replacement of tuffs and volcanics occurs on Tasyatak Tepe. The zone here

dips steeply to the NE and is approximately 25 m thick in outcrop.

Phyllic

The quartz porphyry located at Kestane displays variable phyllic alteration (quartz, sericite and pyrite);

Phyllic alteration is both pervasive (in areas near the top and along the margins of the deposit) and

fracture and vein-controlled. It overprints potassic alteration. In near-surface areas it is commonly

overprinted by supergene argillic alteration. Wall rocks at depth, particularly south of the stock, also

display weak pervasive sericite alteration.

PotassicThe Kestane porphyry is dominated by potassic alteration (biotite, K feldspar, and magnetite). Biotite

commonly replaces mafic phenocrysts. K-feldspar is present as a weak pervasive overprint. Magnetite

is found in quartz veins and as fine disseminations in the groundmass. The highest gold and copper

grades in the core are associated with early biotite + abundant magnetite (5-8%) + chalcopyrite

associated with A- and B-type quartz veins. Chalcopyrite, in some cases accompanied by trace

amounts of pyrite, is the dominant sulf mineralide in the potassic alteration. Bornite has not been

observed.

Structural Geology

The Halilağa area is affected by post mineral faults of the North Anatolian Fault System characterized

by major ENE-WSW transtensional structures.

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7-11

.

Ongoing transtensive kinematics has reactivated pre-existing extensional structures within zones

oriented ~N60E to EW.

A first approach on a regional interpretation suggests that the north-tilting of the pre-quaternary geology

could be due to Late Tertiary to movement along kinematics ENE-WSW structures.

These structures could also have constitutive boundaries for Tertiary volcano sedimentary depocentres

by accommodating several asymmetrical and sub-parallel grabens. All post basement volcano-

sedimentary sequences show a clear tilting to the NW, N, and NE by 20º-35º which increases at the

vicinity of a major structure.

Significant faults act as bounding structures for the Kestane Stock (and hence most of the economic

mineralization) including the E-W-striking North Fault and South Fault. These faults show normal

displacement and together form a horst with the porphyry mineralized Kestane stock in the centre.

The South Fault separates propyllitized to phyllic altered volcanic cover to the south, from the

mineralized Kestane stock to the north. It is E-W trending and dips -60º to the south. This structural

contact was intersected in holes HD-41 (between 123.70 m and 173 m), HD-97 (between 124.5 m and

130.9 m), and HD-94A (between 126.7 m and 132.7 m).

The North Fault separates propyllitized volcanic cover and hornfels to the north, from the mineralized

Kestane stock to the south. It is WNW trending and dips approximately 65º to the north. This structural

contact was intersected in holes HD-49 (at 14.6 m), HD-67A (at 229 m), HD-42 (between 238 m and

246 m), HD-82 (at 313.6 m), and HD-87 (at 378.6 m). In addition, there is a low angle structure

between mineralized material stock and the volcanic cover unit that starts at 483250 E line and dips

at -20º to the east.

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8 DEPOSIT TYPES

Halilağa is classified as a copper-gold porphyry system. Advanced argillic alteration and gold

mineralization at Kunk Hill and Pirentepe are classified as high sulphidation epithermal mineralization.

Copper mineralization in the Bakırlik Tepe area is classified as proximal copper skarn. All three types

are related to magmatic-hydrothermal activity associated with intrusion of the Kestane porphyry stock

and other intrusions in the area.

8.1 General Background

Porphyry copper deposits are widely distributed at convergent plate margins, in association with arc-

related volcanic and intrusive rocks of intermediate composition. They typically occur in association with

skarn and intermediate- to high-sulphidation epithermal base and precious metal deposits. They range

in size from tens of millions of tonnes to billions of tonnes of mineralized material. Copper and copper-gold porphyry systems supply a large proportion of the world’s copper, molybdenum, and gold. Some of

the largest Cu-Au porphyry systems include Grasberg (Indonesia), Oyu Tolgoi (Mongolia) and Bajo de

Alumbrera (Argentina). Associated high sulphidation epithermal deposits can also be quite large, with

the best example being Yanacocha in Peru. Typical hypogene grades of porphyry deposits range from

0.2% to 0.8% copper. Copper to gold ratios vary widely in Cu-Au porphyry deposits, but a Cu %:Au ppm

ratio of 1:1 is not uncommon.

A very large volume of literature exists on porphyry deposits because of their large size and economic

importance. The following description of a porphyry deposit comes from a summary by Sillitoe (2010).

Porphyry deposits are typically centred on polyphase stocks and porphyry dyke swarms, with skarn

deposits formed adjacent to and epithermal deposits above the porphyry mineralization (Figure 8.1).

The metal endowment of a porphyry system is related to the geochemistry of the oxidized magmas that

contribute to the formation of the stocks and dykes, with gold and/or molybdenum commonly found in

association with copper. Porphyry deposits typically occur in association with Mesozoic and Tertiary

intrusions, probably as a result of poor preservation of older rocks.

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Figure 8.1: Anatomy of a Telescoped Porphyry System

Source: Sillitoe 2010

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Porphyry systems are typically zoned from a potassic altered (biotite-potassium feldspar) core overlying

barren, calcic-sodic altered rock, upward through phyllic altered (sericite or chlorite-sericite) margins topropylitic altered (chlorite-epidote) rocks (Figure 8.2). Porphyry systems also grade upward into

advanced argillic, argillic and silicic alteration related to epithermal mineralization. Alteration zoning may

be complex and overlapping due to successive injections of magma into country rocks. The vertical

distance between porphyry mineralization and overlying epithermal mineralization may range from one

telescoped kilometre to several un-telescoped kilometres.

Figure 8.2: Generalized Alteration-Mineralization Zoning Pattern for Telescoped Porphyry Copper

Deposits

Source: Sillitoe 2010

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Hypogene copper mineralization is disseminated and veinlet-hosted, and zoned from bornite-rich in the

core through chalcopyrite to pyrite in distal areas. Magnetite (in Cu-Au porphyries) and molybdenite (inCu-Mo porphyries) are common accessory minerals.

Quartz veins and veinlets as stockworks and sheeted arrays are ubiquitous in these systems, and

typically occur in a sequence from early quartz-feldspar A-veins, through quartz-sufide (mainly

chalcopyrite-molybdenite) B-veins with potassic-altered margins to late, sulphide-dominant (primarily

pyrite) D-veins with phyllic altered margins (Gustafson and Hunt 1975). Veining in Cu-Au deposits may

differ slightly, with quartz-magnetite-chalcopyrite and magnetite-dominant M-veins present or dominant

(Arancibia and Clark 1996).

Due to the large amount of disseminated pyrite in most porphyry systems, these systems are

susceptible to supergene weathering and leaching. Copper is oxidized and leached from areas above

the water table and deposited as chalcocite and other supergene copper minerals at or near the watertable, leading to enriched copper grades. Supergene chalcocite enrichment can increase grades locally

by 200% to 300% or more, with a significant impact on the overall economics of these deposits.

Alteration and mineralization associated with high sulphidation epithermal deposits in the upper portions

of porphyry systems consist of pyrite, enargite, and covellite hosted in silicified and often brecciated

silicified volcanic rocks, accompanied by advanced argillic alteration minerals, including pyrophyllite,

alunite, dickite, and kaolinite (Hedenquist et al 2000). Alteration and mineralization at this elevation in

the system comprise a lithocap and may be far more laterally extensive than the porphyry deposit itself.

Proximal skarn deposits are located laterally from porphyry deposits (Meinert, 2000). They consist of

replacement bodies within the causative intrusion (endoskarn) or marginal to it (exoskarn). Skarn may

be particularly well-developed in limestones and other calcium or carbonate-rich rocks. Skarn alteration

assemblages include garnet, pyroxene, wollastonite, magnetite, actinolite, pyrite, magnetite, and

chalcopyrite.

8.2 Halilağa Porphyry Mineralization

Copper-Gold porphyry, skarn, and high-sulphidation epithermal gold alteration and mineralization are all

found in close proximity in the Halilağa area. Recognizing that the high-sulphidation deposits underlying

many of the hills in the area could be overlying or concealing porphyry deposits at depth led to the

discovery of Halilağa in the valley bottom adjacent to the Kunk high-sulphidation epithermal system.

The Kestane Cu-Au porphyry system exhibits alteration and mineralization zoning typically seen in

deposits of this type. This includes a low-grade, potassic-altered core and relatively high copper andgold grades, often associated with a high density of quartz-magnetite-sulphide veins in areas flanking

the core. Mineralization is also associated with an overlap of phyllic and potassic alteration, a small

supergene chalcocite blanket, and adjacent areas of hornfelsing and skarn alteration.

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One departure from the standard porphyry model is a lack of bornite mineralization in the core of the

deposit. Either this represents a departure from the model, or possibly the core of the deposit has notyet been located.

Magmatic-hydrothermal processes related to hypabyssal quartz diorite to quartz monzonite intrusions

have resulted in porphyry copper-gold (proximal, Kestane); epithermal gold (distal, Kunk Hill); and skarn

gold-silver +/-base metal mineralization (Bakırlik + Kestane W) in and near Halilağa.

8.3 Kestane Porphyry Mineralization

At the Kestane porphyry, most quartz veins are ‘B-type’, averaging 5% of the rock by volume, but locally

up to 20% (shown in Figure 8.3), and ‘A-type’ veinlets are rare or difficult to recognize on outcrops. The

fact that B-veins, shreddy biotite, and D-veins can be recognized in an outcrop is significant because

these indicate the presence of moderately intense potassic alteration with a moderate sericitic overprint(Einaudi 2007). Given the tendency for the best grades in porphyry Cu-Au deposits to be associated

with potassic alteration associated with abundant quartz veins, the possibility of high primary grades in

chalcopyrite or chalcopyrite (± magnetite) assemblages can be inferred from these outcrops.

Additionally, the moderate degree of sericitic alteration suggests that chalcocite enrichment below the

leached cap might be present because acidic conditions at the water table favour the formation of

chalcocite rather than copper-oxides, silicates, and carbonates (Einaudi 2007).

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Figure 8.3: B-Veining in Kestane Porphyry


The geometry of gold mineralization at the Central Zone at Kestane shows good continuity and is

approximately 800 m x 1,600 m in the EW direction and porphyry-style copper-gold mineralization has

been intersected in drilling to a depth of approximately 500 m.

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Epithermal – High-Sulphidation

At Kunk Hill, ENE- and ESE- trending ridges are capped by extensive areas of silicified volcanic rocks.

These lithocaps are formed by massive to vuggy silica (quartz alunite), extensive areas of strongly

limonitized breccias, and argillic to advanced argillic alteration which are the host for high sulphidation

gold mineralization. Limited drilling in this area has returned anomalous gold.

Skarn Mineralization

At Bakırlik, (Four km ESE of Kestane) copper-bearing garnet skarn is present. It occurs in

carbonaceous limestone near the contact with a quartz monzonite intrusion. Whether the Bakırlik

occurrence is the same age as Kestane remains to be determined. West of the Central Zone, another

copper-bearing garnet skarn is present which requires further investigation.

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9 EXPLORATION

2002

The property was acquired by Teck Madencilik via auction.

2002-2004

No activity on the property.

2004

Teck Madencilik and Fronteer entered into an option agreement that covered several properties in the

Biga area, including the Halilağa Project that enabled Fronteer to acquire 100% interest in the

properties subject to certain earn-back rights by Teck Madencilik; Fronteer was required to spend$2.0 million on exploration over four years, with a first-year firm commitment of $200,000.

2005-2006

Fronteer conducted an exploration program consisting of geological mapping, surface geochemical

sampling, a pole-dipole IP survey, and a ground magnetics survey. A drill program was initiated on

November 15, 2006.

The soil and rock chip sample results highlighted the porphyry-related mineralization of the Central

Zone at Kestane. Rock chip sampling of oxidized and leached outcrops returned 19 samples (out of 40

collected) with gold values greater than 1.0 g/t.

Forty-three line km of IP Chargeability/Resistivity and 44 line km of ground magnetic surveying were

completed. The most significant feature generated by the surveys was a coincident high chargeability

and high magnetic anomaly associated with the Central Zone.

The Kestane porphyry was discovered with the first drill hole HD-01, and Teck Madencilik decided to

exercise their earn-back, and assumed operations on November 30, 2006.

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2007

Teck Madencilik earned a 60% interest in the property by investing $2.5 million during 2007.

Geological mapping of the Central Zone and surrounding area, was completed by Teck Madencilik at a

scale of 1/10,000. Extensive soil, rock chip, and silt orientation sampling programs were also carried

out. A total of 3,650 soil, 172 rock, and 58 silt orientation samples were collected from the Halilağa

area. The soil results showed that strong surface geochemical anomalies were not only restricted to the

Kestane area, but also occurred to the southeast at Bakırlik and in the central and southern parts of the

property.

Sixty-four line km of IP chargeability/resistivity and 263 line km of ground magnetics were completed.

Subsequent 3-D magnetic modelling of the data by Teck Madencilik (see Figure 9.1) showed a large

magnetic body at depth that comes to surface.

A total of 1,164 soil and 409 rock samples were also collected from the Pirentepe area. A total of 16.3

line km of IP chargeability/resistivity and 58.8 line km of ground magnetics were also completed. The

most significant anomaly generated was a coincident high chargeability and low magnetic anomaly

associated with the NE-trending structures and epithermal gold mineralization.

High magnetic anomalies map the basaltic volcanic rocks SE of Pirentepe, as well as the Kocak ısla

area, suggesting potential for porphyry mineralization.

2008

Matias Sanchez (Sanchez 2008) regionally mapped the Halilağa Project as part of his Ph.D. thesis (seeFigure 9.2). The map includes rock types, alteration, and detailed structural information.

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Figure 9.1: 3-D Magnetic Inversion of Halilağa Data


This anomaly is the coincident high chargeability and high magnetic anomalies associated with the

Central Zone porphyry copper-gold mineralization. High resistivity anomalies map the silica alteration

along the top of ridge line.

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Figure 9.2: Geological Map of Halilağa Property (Sanchez, 2008)

Source: Sanchez 2008

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.

A total of 566 rock samples were collected from selected areas (see Figure 9.3). The 2008 rock

geochemical programs highlighted three new targets:

Kunk North (> 0.5 g/t Au) is part of the Kunk Hill lithocap with strongly silicified, locally vuggy

and brecciated volcanics.

Kumlugedik Hill (> 1.0 g/t Au) has widespread float of silicified and quartz-veined (more

epithermal-low temperature) meta-sediments, and volcanic sediments.

Madendere (> 0.2 g/t Au) has more porphyry style alteration characterized by strong quartz

sericitic volcanic-sub volcanics with a few millimetre-thick quartz veins (possibly B-type). The

size of the selectively sampled area extends more than 2 km long by 200 m wide.

In addition to these areas, additional copper-gold rock geochemistry anomalies at Kizilciktasi,

Osmaniye, and Yaniklar were identified (Ceyhan et al 2009).

Figure 9.3: Rock Geochemistry after 2008 Sampling

Source: Ceyhan et al., 2009

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.

A total of 163 rock-saw samples were collected and returned several gold anomalies up to 7.5 g/t. The

2008 rock sampling program confirmed a greater than 3 km long NE-trending structural corridor. Theprogram also highlighted the Kocakisla area with > 0.1 g/t Au assay results. The rock geochemical

results were also coincident with soil geochemistry results anomalies of > 10 ppb Au along a NE-

trending structure as three separate zones, as well as in the Kocakısla area.

2009 and 2010

A total of 36 rock chip samples were collected during the 2009 field season. Significant highlights of this

program include:

Rock chip sampling results from Kizilciktasi (0.1–0.5 g/t Au); and

Confirmation of anomalous gold in rock-saw samples from north of Kunk Hill (> 0.5 g/t Au).

IP chargeability/resistivity surveys were carried out in 2009 and 2010. The 2009 pole-dipole IP survey

focused on deep targets near Kestane. This data was acquired on six lines (19300N (EW), 82000E,

83300E, 83800E, 84300E, and 84800E (all N-S)). In May 2010, a large IP program was carried out in

the southern portion of the Halilağa licenses in order to extend the IP coverage from previous years.

The data was acquired using a conventional “n” of 6, “a”-spacing of 100 m, pole-dipole configuration

array. These data was subsequently merged with older data sets to create complete chargeability and

resistivity plan maps for the Halilağa area (Figure 9.4 and Figure 9.5). The entire program totaled 54

line-km of data.

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Figure 9.4: Compiled IP Chargeability Plan for 100 m Depth

Source: Teck Madencilik Annual Report 2009 supplied by Pilot Gold

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Figure 9.5: Compiled IP Resistivity Plan for 100 m Depth

Source: Teck Madencilik Annual Report 2009 supplied by Pilot Gold Inc. 2014

On-going metallurgical sampling tested the hornfels units of holes HD-40 and HD-49. Acid Rock

Drainage (“ARD”) tests were performed on the pulps from holes HD-05, HD-09, and HD-48A. Further

petrographic samples were taken from holes HD-40, HD-48A, HD-49, and HD-54. An environmental

base-line study was also initiated over the Halilağa Project.

On December 31, 2009, Teck Madencilik declined to earn an additional 10% interest in the Halilağa

Project.

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.

2011

A total of 162 rock, 143 soil and seven silt samples were collected, primarily in the Halilağa North area

near the village of Halilağa. This was followed up by a 17.4 line-km pole-dipole IP survey and a 16.4

line-km ground magnetic survey. The combination of anomalous soil and rock chip samples and a zone

of moderate chargeability, resistivity and magnetic response suggested the presence of a viable

porphyry target in this area.

2012-2014

No surface work was carried out during this period by Teck Madencilik.

Table 9.1 shows the program summarized, by number and line-km, for geochemical sampling and

geophysics by year.

Table 9.1: Details of the Halilağa Project by Year

Year 2005 2006 2007 2008 2009 2010 2011

Geochemical Sampling ( #) 383 1197 3880 729 36 0 312

IP/Resistivity (line-km) 8 35 64 0 21 33 17.4

Ground Magnetics (line-km) 0 44 263 0 0 0 16.4

Source: Avsan, M., Kilic, D., Keles, S., Kurcan, K., 2011. Halilağa Property Exploration 2011 Year End Report

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10 DRILLINGThe drilling between 2007 and 2012 was performed by Spectra Jeotek of Ankara, Turkey, and was

conducted using two to five, contractor-manufactured drill rigs. The model numbers are D150 and D220

with depth capacities of 1,000 m and 1,500 m of HQ, respectively. In 2011, the drilling was done with

tri-cone bit (Q=120 mm) until the contact with the QFP/hornfels was made and then continued with

HQ core after casing the hole. Between 2007 and 2010, the drilling was mostly HQ drilling which was

then reduced to NQ when ground conditions became difficult. The HRC series drill holes are reverse

circulation (“RC”) type holes.

The drill hole collars for holes HD-01 through HD-35 were surveyed using Total Station methods. The

subsequent holes HD-36 and above were surveyed using a differential global positioning system

(“DGPS”) with a horizontal and vertical accuracy of generally ± 20 cm. Drill hole deviation wasmeasured using Reflex Survey tests taken between 50 m and 100 m intervals down hole to provide

control.

10.1 Halilağa Drilling

Subsequent to property acquisition, all drilling was performed by Fronteer or Teck Madencilik or Truva

Bakır and is listed in Table 10.1.

Table 10.1: Halilağa Drill Summary

Target Year# of Drilled

Holes

Total

(m)

# ofAbandoned

Holes

TotalAbandoned

(m)

Gross #

of Holes

GrossTotal

(m)

Kestane-Kunk 2006-2007 18 5,827 5 519 23 6,346

Kestane+Kunk 2009 14 5,243 4 427 18 5,670

Kestane 2010 11 6,098 14 2,978 25 9,076

Kestane 2011 39 17,973 4 1,159 43 19,132

Kestane 2012 22 8,592 5 293 27 8,885

Kestane Total 2006-2012 104 43,733 32 5,376 136 49,109

Pirentepe 2006-2007 28 4,970 1 37 29 5,007

Bakırlik 2008 18 4,019 2 32 20 4,051

Madendere 2011 1 467 1 467

Madendere 2012 2 207 2 207

Grand Total 2006-2012 153 53,396 35 5,445 188 58,841

Source: JDS 2014

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.

In the 2006-2007 period, a total of 20 diamond holes totaling 5,950 m (including five abandoned holes)

and three reverse circulation holes totaling 396 m were completed. The drilling focused in two mainzones: the Kestane porphyry and the Kunk-Kumlugedik lithocap areas. 17 holes (15 diamond and two

RC) totaling 4,756.1 m were drilled at Kestane. Most of the 2007 drill holes at the Kestane Zone

intersected porphyry-style Cu-Au mineralization with economic grades (for example, HD-01 returned

1.03 g/t Au and 1.03% Cu over 105.4 m). A 25 metre-thick chalcocite blanket (approximately 2% Cu)

was intersected close to the surface in holes HD-01, HD-02, HD-04, and HD-14. A total of six holes (five

diamond and one RC) totaling 1,589.9 m were drilled at the Kunk-Kumlugedik Zone. Holes HD-09 and

HD-10 holes intersected narrow low-grade gold mineralization.

In 2008, a total of 20 diamond holes (including one abandoned hole) totaling 4,051 m were completed.

The main objective of the program was to test the coincidence of magnetic high anomalies and

anomalous Cu and gold values in the soil at the Bakırlik area which is located 4 km SE of Kestane

porphyry Cu-Au mineralization.

A total of nine samples from the holes drilled at Kestane and Bak ırlik zones were petrographically and

mineralogically analyzed (Kuşcu 2008).

In 2009, a total of 18 diamond holes (including four abandoned holes) totaling 5,670 m were completed.

All holes were located at Kestane (central) Zone and tested porphyry Cu-Au mineralization. Drilling

focused on further defining the limits of the Kestane target which steps out to the north, south, and east.

The 2009 drilling footprint increased the extent of the known Kestane mineralization to 1,200 x 450 x

195 m. Significant intersections were encountered in HD-41: 159.5 m @ 0.59 g/t Au and 0.50% Cu; HD-

45: 308.1 m @ 0.34 g/t Au and 0.30% Cu; and HD-38: 103.0 m @ 0.40 g/t Au and 0.70% Cu; all other

significant drill intervals are shown in Table 10.2. A series of consistently eastward, down dropping fault

bound blocks are shown in Figure 10.1 and Figure 10.2.

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.

In 2010, 25 holes (20 diamond and five RC) totaling 9,076.6 m (including 14 abandoned holes) werecompleted. With the exception of holes HD-64 and HD-69, which tested the Kestane-West target, all

holes were located at Kestane (central) Zone. Drill hole HD-54 returned an intercept of 646.5 m @

0.26 g/t Au and 0.33% Cu, representing the longest interval of copper-gold mineralization to date at

Halilağa. The hole collared in a high grade, near surface blanket of chalcocite, then intersected

moderate to high-grade copper zones and then passed into low to moderate-grade copper and gold

mineralization for the remainder of its length (see Table 10.2). The drill hole tested an area where the

QFP intrusive was 20.9 m below surface and was interpreted to represent an up-thrust block related to

the general regional, ENE-WSW transtensional structures of the North Anatolian Fault System. HD-54

successfully tested the northern margin of the intrusive and passed into a down-dropped block of QFP

intrusive at depth.

The southern margin of this same up-thrusted block was tested with HD-57 which returned 348.80 m @0.28 g/t Au and 0.31% Cu (see Figure 10.3). Holes HD-56, HD-60, and HD-61 drilled into the hornfels

aureole on the north and west side of the Kestane Zone and also intersected anomalous molybdenum

over long widths.

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.

The 2011 program included 44 drill holes (including four abandoned holes) for a total 19,599 m ofdrilling. One of those drill holes which extended 467 m was directed toward testing the Madendere area.

The focus of the 2011 drilling was to extend and in-fill the Kestane Zone to demonstrate continuity and

increase confidence in order to calculate an initial resource estimate for the deposit. Significant

intersections included HD-76 with 448.7 m @ 0.35 g/t Au and 0.22% Cu, and HD-80 with 323.3 m @

0.27 g/t Au and 0.32% Cu. HD-86 returned an intercept of 185.6 m @ 0.46 g/t Au and 0.61% Cu.

However, the hole collared in the high grade, near surface blanket of chalcocite, as previously

identified, included 71.3 m @ 0.71 g/t Au and 1.1% Cu. The 2011 drilling has now shown the Kestane

porphyry to be mineralized over a strike length of 1,200 m, a width of 850 m, and a thickness of up to

600 m.

Table 10.2 lists all of the Halilağa drill holes and Table 10.3 summarizes the representative assay

results.

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Table 10.2: Halilağa Drill Hole Collars

Hole IDEast

(UTM m)North

(UTM m)Elevation

(masl)Depth

(m)Azimuth(degrees)

Dip(degrees)

HD-02 483192.54 4419151 349 292.8 0 -60

HD-03 482843.91 4419446 308 242.9 0 -90

HD-04 483120.28 4419342 337.9 320.4 0 -55

HD-05 482568.48 4419104 313.5 285.2 0 -90

HD-06 482852.82 4419252 318 279.1 0 -90

HD-07 483539.88 4419296 360.1 23.3 0 -90

HD-07A 483539.92 4419299 360 330.7 0 -90

HD-08 483886.85 4419361 376.3 434.5 270 -60

HD-09 483817.64 4418745 530.1 355.4 0 -60

HD-10 484022.06 4418850 497.9 297.5 0 -60

HD-11 484201.34 4418248 556.1 296 0 -60

HD-12 483405.77 4418310 531.9 138 0 -60

HD-12A 483405.83 4418312 531.9 353 0 -60

HD-13 483431.97 4419112 376.3 97.6 315 -60

HD-13A 483431.3 4419113 376.3 505.2 315 -60

HD-14 483173.69 4419207 348.8 539.7 315 -60

HD-15 483094.81 4419544 314.1 126.9 0 -60

HD-15A 483094.87 4419544 314.2 133.3 0 -60

HD-16 483015.93 4419154 343.5 602 0 -60

HD-17 485636.11 4418257 440.6 285.5 330 -60

HD-18 485672.97 4418687 394.7 212 0 -50

HD-19 487400.53 4416747 492.4 189.6 325 -50

HD-20 486608.09 4417847 444.1 26 135 -50

HD-20A 486608.89 4417846 444.1 311.4 135 -50

HD-21 486429.68 4417817 486.4 418.4 100 -50

HD-22 485486.2 4417834 536.2 230.6 300 -60

HD-23 485470.78 4417745 527.1 247.2 115 -50

HD-24 486603.5 4417709 428.9 201.4 90 -50

HD-25 486793.99 4417441 361.4 265.1 120 -50

HD-26 484938.87 4416621 561.1 291.5 245 -50

HD-27 487208.92 4416746 546.5 6 335 -50

HD-28 486788.57 4417441 361.4 157.7 300 -50

HD-29 486781.04 4417735 401.9 438.3 180 -60

HD-30 486604.19 4417599 392.2 56 170 -60

HD-31 486134.85 4417905 476.1 196.6 360 -60

HD-32 486805.39 4417270 384.3 256 135 -60

HD-33 486581.39 4417909 442.7 71.5 0 -60

HD-34 486739.11 4416913 400.1 83.4 270 -70

HD-35 486846 4417609 343 106.8 225 -70

HD-36 482951.8 4419558 306.2 149 0 -60

HD-37 483139.2 4419154 336.5 268.5 0 -60

HD-38 483244.55 4419456 326.9 586.1 180 -60

HD-39 483348.49 4419354 341.4 168.5 180 -60

HD-40 483347.28 4419354 341.4 520.7 0 -60

HD-41 483138.7 4419156 336.4 180.6 180 -60

HD-42 483646.47 4419555 337.6 393.5 175 -60

HD-42D 483646.47 4419555 337.6 614 175 -60

HD-43 483845.67 4419575 345 71.9 180 -60

HD-43A 483845.62 4419574 345.1 95.8 180 -60

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HD-49 482950.78 4419551 306.5 541.4 180 -50

HD-50 483031.15 4419254 329 308 270 -45

HD-51 483841.28 4419577 345 76 180 -60

HD-52 483148.93 4419638 307 287 0 -60

HD-53 483742.69 4419554 347 108 160 -60

HD-54 483060.6 4419454 325.6 791 0 -60

HD-55 483249.69 4419044 352 98.2 200 -60

HD-56 482560.78 4419153 316.4 611.1 315 -75

HD-57 483036.78 4419252 329.1 727.3 180 -50

HD-58 482947.9 4419748 296.9 314 0 -90

HD-58A 482944.5 4419746 297.1 431.8 0 -90

HD-60 483250.9 4419349 344.5 798.3 0 -60

HD-61 482854.9 4419353 310.8 502.2 0 -60

HD-62 482854.1 4419255 317.7 625.1 180 -50

HD-63B 483149.9 4419635 306.9 350 0 -90

HD-64 482098 4419434 340.8 683 225 -70HD-65 483548.5 4419554 325 265 0 -90

HD-66 483248.6 4419047 356.7 161.4 0 -90

HD-66A 483249.5 4419048 356.6 183 0 -90

HD-67 483548.7 4419552 324.8 194.5 180 -60

HD-67A 483548.7 4419549 325.2 667.2 180 -60

HD-69 482206.3 4418781 283.3 550 270 -50

HD-70 481944.7 4418506 285.2 595.8 170 -70

HD-71 483246.6 4419043 356.7 71 0 -60

HD-72 484251.4 4419347 379 710.6 0 -90

HD-73 484360.6 4417957 748 632.1 0 -60

HD-74 482742.52 4419539 299.3 515.6 180 -60

HD-75 482848.13 4419047 331.8 46.1 180 -60

HD-75A 482848.13 4419046 331.8 151.4 180 -60

HD-75B 482849.72 4419046 332 300.7 180 -60

HD-75C 482849.91 4419040 331.9 153.6 180 -60

HD-76 483647.91 4419445 353.1 869 135 -60

HD-77 484040.69 4419166 415.8 853.8 0 -65

HD-78 482750.25 4419348 310.8 510 180 -60

HD-79 482744.84 4419162 320.8 500 180 -60

HD-80 483140.42 4419453 331.6 857 0 -65

HD-81 482944.05 4418969 344.8 512.5 180 -60

HD-82 483841.71 4419571 344.5 437 180 -60

HD-82D 483841.71 4419571 344.5 635.8 180 -60

HD-83 484837.09 4419155 317.7 247 0 -90

HD-83A 484837.59 4419157 317.2 167.8 0 -90

HD-84 482447.02 4419457 333.5 709.9 90 -60

HD-85 484540.08 4419262 358.4 473.7 0 -90

HD-86 483047.76 4419358 317.2 504.9 0 -90

HD-87 484049.38 4419561 353.1 884.5 180 -60

HD-88 484737.73 4419057 337.9 427 0 -90

HD-89 482444.97 4419555 328.8 326.6 90 -50

HD-90 482950.15 4419360 314.7 57.1 0 -90

HD-90A 482950.31 4419363 314.6 462.6 0 -90

HD-91 483345.29 4419252 353 449.4 0 -90

HD-92 483346.61 4419555 318.1 60 180 -60

HD-92A 483346.51 4419552 318.1 592.4 180 -60

HD-93 484751.09 4419254 340.9 436 0 -90

HD 94 483246 58 4419052 355 5 453 0 70

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HD-98 483541.56 4419297 359.8 254 90 -60

HD-98A 483540.29 4419296 359.5 503 90 -60

HD-99 484036.09 4419353 383 368 270 -60

HD-99A 484032.95 4419353 383 503.6 270 -60

HD-100 483440.87 4419450 326.8 564.6 90 -60

HD-101 482804.15 4416723 383.5 467.3 0 -90

HD-102 482947.49 4419151 337.8 325 0 -75

HD-103 483741.34 4419553 347.2 505.5 180 -75

HD-104 484043.37 4419167 415.1 275 315 -60

HD-105 482873.74 4416036 403.6 43.6 90 -60

HD-105A 482872.83 4416036 403.8 163.5 90 -60

HD-106 483350.5 4419048 370.5 205 0 -90

HD-107 483650.11 4419446 353.3 435.9 180 -55

HD-108 482853 4419253 318.1 200 0 -60

HD-108A 482852 4419251 318.2 45.9 0 -60

HD-109 483046 4419363 318.2 401.2 345 -65HD-109A 483045 4419364 318.2 30.8 345 -65

HD-110 483349 4419051 370.6 300 0 -65

HD-111 483537 4419295 359.6 378.5 180 -50

HD-112 483740 4419552 347.3 497.4 180 -63

HD-113 483544 4419302 359.7 383.2 140 -50

HD-113A 483547 4419305 359.6 491.1 140 -50

HD-114 484042 4419177 421.2 401.5 0 -90

HD-115 483444 4419251 356.7 413.4 180 -50

HD-116 484042 4419166 413.4 407 250 -60

HD-117 483925 4419359 377.5 497 180 -50

HD-118 483437 4419355 342.3 410.5 0 -50

HD-119 483541 4419457 335.4 20 180 -50

HD-119A 483541 4419456 335.4 472.6 0 -50

HD-120 483925 4419360 377.4 500 180 -75

HD-121 483438 4419449 327 238 180 -75

HD-121A 483438 4419449 327 125 180 -90

HD-121B 483438 4419449 327 50 180 -90

HD-121C 483438 4419449 327 67 180 -90

HD-122 483250 4419050 360 450 180 -80

HD-123 482953 4419561 306 450 0 -90

HD-124 483245 4419254 355 436 0 -75

HRC-01 485580.97 4417853 549.8 150 265 -60

HRC-02 482993.39 4419263 328.1 145.5 0 -60

HRC-03 482914.05 4419177 329.7 100.5 0 -90

HRC-59 483345.3 4419665 304 192 0 -90

HRC-63 483153.6 4419640 307.2 54 0 -90

HRC-63A 483146.7 4419625 307.6 192 0 -90

HRC-68 483137.8 4419153 336.1 143.5 190 -60

Source: JDS 2014

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Table 10.3: Assay Summary

Hole No. Zone From(m)

To(m)

Interval(m)

Au(g/t)

Cu%

Ag(g/t)

Mo%

HD-01

Kestane

0 298.2 298.2 0.5 0.53

Incl 0 105.4 105.4 1.03 1.03 Incl 23.85 49.6 25.75 0.93 2.14

Incl 49.6 287 237.4 0.44 0.43 HRC-02

Kestane

0 145.5 145.5 0.36 0.34 Incl 0 78 78 0.48 0.47 Incl 4.5 25.5 21 0.69 0.92 HRC-3

Kestane

0 100.5 100.5 0.2 0.2 Incl 0 16.5 16.5 0.38 0.36

Incl 9 16.5 7.5 0.51 0.75 HD-02

Kestane

0 292.8 292.8 0.4 0.35 Incl 4.7 233 228.3 0.47 0.41

Incl 15 91.4 76.4 0.59 0.63 Incl 15 32 17 0.73 1.04 Incl 81.3 91.4 10.1 1.35 1.18 HD-03

Kestane0 242.9 242.9 0.14 0.16

Incl 101.8 109.7 7.9 0.46 0.24 HD-04

Kestane

0 320.4 320.4 0.4 0.48 Incl 0 110.8 110.8 0.74 0.89

Incl 22.3 84.8 62.5 0.91 1.34 HD-06

Kestane

0 279.1 279.1 0.13 0.13

Incl 0.6 13.8 13.2 0.09 0.24 Incl 87.2 153.5 66.3 0.16 0.16

Incl 126.5 149.1 22.6 0.21 0.2 Incl 176.3 207.4 31.1 0.15 0.16 HD-7A

Kestane

0 330.7 330.7 0.23 0.17

Incl 61.3 302.5 241.2 0.3 0.22 Incl 78.5 167.5 89 0.33 0.23 Incl 237.5 302.5 65 0.34 0.26

Incl 263.5 273.5 10 0.52 0.35 HD-08

Kestane

167.2 434.5 267.3 0.5 0.26

Incl 208.7 248 39.3 0.36 0.19 Incl 248 381.05 133.05 0.69 0.34 Incl 248 328.2 80.2 0.81 0.38 Incl 258.3 302.2 43.9 0.91 0.41 HD-09

Kunk Hill110 125.8 15.8 0.23 0

And 144 147 3 0.35 0 HD-10

Kunk Hill

70.1 123.7 53.6 0.29 0

Incl 94 113 19 0.48 0.01 Incl 245.1 278.1 33 0.2 0 HD-13

Kestane80.6 97.6 17 0.86 0.6

Incl 83.4 97.6 14.2 1 0.68 HD-13A

Kestane

97 479.95 382.95 0.41 0.34 Incl 97 240 143 0.72 0.5

Incl 106.6 140.5 33.9 1.26 0.76 Incl 180.5 207.8 27.3 0.98 0.69

Incl 305.85 399.25 93.4 0.26 0.32 Incl 407.1 432.45 25.35 0.27 0.28 Incl 447.6 477.4 29.8 0.3 0.28 HD-14

Kestane

0 278.2 278.2 0.42 0.47 Incl 0 211.25 211.25 0.51 0.58 Incl 0 100.2 100.2 0.66 0.8

Incl 0 14.3 14.3 1.49 0.06 Incl 14.3 39.1 24.8 0.96 2.42

Incl 124.9 162.2 37.3 0.65 0.64 Incl 139.6 158.3 18.7 0.83 0.84

HD-16

Kestane

0 314.2 314.2 0.29 0.24 Incl 0 190.2 190.2 0.38 0.29

Incl 54.9 171.8 116.9 0.48 0.35 Incl 54.9 133.6 78.7 0.55 0.4

Incl 54.9 85.7 30.8 0.72 0.54 HD-19

Bostanlikbasi13.3 96 82.7 0.02 0.07

Incl 29.6 50 20.4 0.01 0.11 HD-21

Bakırlik274.9 275.4 0.5 2.8 16.85

And 296.5 298.5 2 0.48 2.01 HD-25

Bakırlik

50.9 53.6 2.7 0.25 1.23 8.29

And 205.6 217.5 11.9 1.29 2.34 34.87 Incl 209.6 215 5.4 2.38 4.51 69.76 HD-37

Kestane

3 183 180 0.72 0.67 Incl 3 162 159 0.77 0.72 Incl 3 124 121 0.83 0.8 Incl 4.2 47.6 43.4 1.11 1.25 HD-38

Kestane

0 111 111 0.39 0.65 Incl 0 85.75 85.75 0.44 0.77

Incl 0 62 620.5 0.93

Incl 14 54.8 40.8 0.55 1.3 Incl 14 34.6 20.6 0.63 2.01

And 327.4 482.1 154.7 0.17 0.2 HD 39 41 4 137 95 6 0 31 0 25

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Incl grade as "0" 7 40.2 33.2 0.87 0.84 HD-41 Kestane-Excluding

"No sampleIntervals"

7 166.5 159.5 0.72 0.62

Incl 7 70 63 0.91 0.84 Incl 7 40.2 33.2 1.1 1.06 HD-42

Kestane

256 393.5 137.5 0.42 0.27

Incl 256 289.4 33.4 0.55 0.32 Incl 311.5 344 32.5 0.56 0.37 HD-42D

Kestane260 401 141 0.4 0.25

Incl 312.6 350 37.4 0.58 0.37 HD-44

Kestane48.8 176 127.2 0.36 0.28

Incl 60 105.8 45.8 0.44 0.32 HD-45

Kestane

103.2 411.3 308.1 0.34 0.3 Incl 204.6 334.5 129.9 0.45 0.35

Incl 323.5 334.5 11 0.67 0.51 HD-46

Kestane

13 154 141 0.23 0.27 Incl 18.1 33.4 15.3 0.19 0.71

And 278.3 292.5 14.2 0.22 0.26 HD-48A

Kestane

293.4 571 277.6 0.15 0.12 87 Incl 293.4 498.5 205.1 0.17 0.12 96

Incl 416.5 472.5 56 0.23 0.19 171 Incl 446.5 474.5 28 0.22 0.17 269 HD-49

Kestane

14.6 300.1 285.5 0.32 0.31 Incl 75.9 182.8 106.9 0.46 0.41 Incl 110 123.5 13.5 0.86 0.56

HD-50

Kestane

0 112.6 112.6 0.4 0.49 Incl 0 56.6 56.6 0.48 0.7 Incl 9.2 33 23.8 0.44 1.17

And 163.2 253.3 90.1 0.22 0.19 Incl 207 231.5 24.5 0.32 0.26 HD-54

Kestane

0 646.5 646.5 0.26 0.33 _ _

Incl 0 392.8 392.8 0.32 0.41 _ _Incl 0 113 113 0.65 0.84 _ _Incl 17 50 33 0.63 1.5 _ _Incl 502.8 646.5 143.7 0.21 0.23 _ 0.011HD-57

Kestane

0 348.8 348.8 0.28 0.31 _ _Incl 0 132.2 132.2 0.42 0.51 _ _Incl 7.1 34.5 27.4 0.47 1.22 _ _And 312.6 438.8 126.2 0.11 0.14 _ 0.013HD-60

Kestane

0 251.2 251.2 0.26 0.33 _ _

Incl 0 61.2 61.2 0.39 0.56 _ _Incl 27.9 54.8 26.9 0.33 1.16 _ _Incl 129.2 251.2 122 0.27 0.32 _ 0.012

Incl 167.2 219.2 52 0.31 0.35 _ 0.022And 167.2 377.2 210 0.18 0.21 _ 0.017HD-61

Kestane

0 200.2 200.2 0.17 0.2 _ 0.022

Incl 90 200.2 110.2 0.21 0.25 _ 0.033

Incl 145.1 194.5 49.4 0.24 0.25 _ 0.046

And 36 329.7 293.7 0.15 0.18 _ 0.024

HD-62

Kestane

2 25.8 23.8 0.09 0.35 _ _

And 133.9 201.2 67.3 0.16 0.15 _ 0.011

And 237.2 348.6 111.4 0.19 0.16 _ _

Incl 274.6 292.6 18 0.34 0.3 _ _

HD-67AKestane

229.4 425.8 196.4 0.33 0.26 _ _

Incl 229.4 292.3 62.9 0.41 0.33 _ _HRC-68 Kestane 1 75 74 0.74 0.63 HD-74 Kestane-W 128.9 244.5 155.6 0.053 0.062 _ 0.011

HD-76

Kestane

180 628.7 448.7 0.35 0.22 _ _Incl 180 383.9 203.9 0.48 0.28 _ _

Incl 207.4 297.8 90.4 0.68 0.37 _ _Incl 572.7 297.8 56 0.36 0.25 _ _HD- 77 Kestane-E 189.3 853.8 664.5 0.14 0.12 _ 0.008Incl Kestane-E 219.3 378.7 159.4 0.17 0.07 _ 0.01

HD-80

Kestane

0 323.3 323.3 0.27 0.32 _ 0.009Incl 0 198.5 198.5 0.35 0.39 _ 0.01

Incl 19.8 39.6 19.8 0.5 0.92 _ -

Incl 99.2 200.5 101.3 0.26 0.31 _ 0.013

And 301.5 358.9 57.4 0.12 0.18 _ 0.01

And 800.7 834 33.3 0.2 0.26 _ 0.008HD-82

Kestane

314.4 435.2 120.8 0.36 0.19 _ _Incl 318 402.3 84.3 0.44 0.22 _ _

Incl 336 363.1 27.1 0.75 0.27 HD-82D

Kestane

314 403.9 89.9 0.49 0.23 _ _

Incl 334.8 363.1 28.3 0.83 0.29 _ _And 513.5 591 77.5 0.25 0.23 _ _And 612.2 635.8 23.6 0.3 0.23 _ _HD-84

Kestane-W

16 136.7 120.7 0.15 _ _

Incl 16 52.7 36.7 0.17 _ _And 119.5 136.7 17.2 0.37 _ _And 197.3 201.9 4.6 1.09 0.15 _ _

HD-86

Kestane

0 185.8 185.8 0.46 0.61Incl 0 137.5 137.5 0.57 0.77

Incl 3 74.3 71.3 0.71 1.1

Incl 5 34 29 0.75 1.67

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HD-94

Kestane

182 453 271 0.27 0.29 0.012Incl. 182 228.5 46.5 0.91 0.77

Incl. 182 398.9 216.9 0.31 0.33Incl. 228.5 398.9 170.4 0.19 0.24 0.013HD-94A

Kestane132.7 250 117.3 0.67 0.54

Incl. 157.2 220.4 63.2 1.07 0.84HD-97

Kestane

120 218 98 0.21 0.17

Incl. 152 167.4 15.4 0.47 0.36Incl. 156.1 156.7 0.6 4.91 4And 218 271.8 53.8 0.28 0.24And 304.2 312.2 8 0.34 0.39HD-98 Kestane 100 254 154 0.33 0.18HD-98A

Kestane

254 425 171 0.23 0.15

Incl. 254 285.1 31.1 0.37 0.25Incl. 315.7 343 27.3 0.3 0.18HD-99

Kestane293.5 368 74.5 0.36 0.2

Incl. 340 368 28 0.53 0.29HD-100 Kestane 229 437.8 208.8 0.28 0.29HD-102 Kestane 0 192.4 192.4 0.22 0.22HD-103 Kestane 360 424.5 64.5 0.32 0.16HD-107 Kestane 151 434 283 0.35 0.19HD-108

Kestane

0 27.6 27.6 0.57 0.62 Incl. 16 25.6 9.6 1.55 1.32 And 35.5 39.1 3.6 2.16 1.24

And 128.7 151.7 23 0.35 0.26 0.01Incl. 138.2 140.5 2.3 0.81 0.53 Incl. 148.7 151.7 3 0.84 0.49

And 174.4 184 9.6 0.23 0.23 0.013

And 190 196 6 0.6 0.59 HD-108A

Kestane

0 27.8 27.8 0.57 0.57

Incl. 17.1 25.8 8.7 1.64 1.2 And 33.8 39.2 5.4 1.96 0.99 HD-109

Kestane

0 119.4 119.4 0.41 0.77 Incl. 5.8 43.5 37.7 0.63 1.77

Incl. 78.4 119.4 41 0.36 0.38 And 192.5 266.7 74.2 0.22 0.24 HD-109A Kestane 0 30.8 30.8 0.73 1.53 HD-110

Kestane122.7 232 109.3 0.44 0.31

Incl. 171.1 211.3 40.2 0.6 0.42 HD-111

Kestane

83.35 252.8 169.45 0.23 0.2 Incl. 83.35 93.1 9.75 0.27 0.42 Incl. 174.2 215.6 41.4 0.27 0.25

And 375.9 378.5 2.6 0.45 0.44 HD-112

Kestane266.4 497.4 231 0.33 0.2

Incl. 285.7 322.8 37.1 0.64 0.29 HD-113

Kestane103 268 165 0.28 0.17

Incl. 104.5 151.3 46.8 0.35 0.22 HD-113A Kestane 69.3 106 36.7 0.26 0.17 HD-114

Kestane

16.5 51.6 35.1 0.32

Incl. 42.2 51.6 9.4 0.76

And 339.2 359.2 20 0.12 0.14 0.009

HD-115 Kestane 93 229.8 136.8 0.64 0.52 HD-116

Kestane SE20.7 65.6 44.9 0.39

Incl. 34.5 56.2 21.7 0.52

HD-117

Kestane

109.2 306.5 197.3 0.28 0.19 Incl. 109.2 171.7 62.5 0.18 0.07

Incl. 171.7 306.5 134.8 0.33 0.25 Incl. 171.7 212 40.3 0.49 0.32 HD-118

Kestane

89.3 407.5 318.2 0.28 0.23 0.01

Incl. 181.4 226.1 44.7 0.38 0.3 Incl. 399.5 405.5 6 0.72 0.46 HD-119A

Kestane119.1 472.6 353.5 0.3 0.24 0.009

Incl. 159.1 252.1 93 0.45 0.34 0.01

HD-120

Kestane

150 468.2 318.2 0.3 0.17

Incl. 238.6 442.4 203.8 0.36 0.23 Incl. 238.6 319 80.4 0.44 0.25 HD-121 Kestane 161.85 238.2 76.35 0.25 0.26 0.017

HD-122 Kestane 259 263.5 4.5 0.19 0.19 HD-123

Kestane

62.5 91 28.5 0.17 0.18 Incl. 62.5 74.45 11.95 0.24 0.25 0.008

And 119.5 379.9 260.4 0.19 0.22 Incl. 119.5 195.5 76 0.24 0.23

Incl. 318.3 329.8 11.5 0.52 0.66 HD-124 Kestane 0 21.8 21.8 1.09

And 21.8 180.5 158.7 0.23 0.26Incl. 21.8 37 15.2 0.43 1.04

Incl. 52.5 68 15.5 0.43 0.28And 305.4 436 130.6 0.2 0.23 0.005

Source: JDS 2014

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10.2 Pirentepe DrillingIn 2006, a total of 890.1 m of drilling was completed at the Pirentepe area. Ten holes were drilled with

the first two holes in the program intersecting gold mineralization (PD-1: 1.83 g/t Au over 46.90 m,

starting at 17 m depth, and PD-2: 1.82 g/t Au over 38.0 m starting at 24 m). In 2007, thirteen (13)

diamond holes totaling 3,031.9 m and six reverse circulation holes totaling 1,084.7 m were completed.

The objectives of the drilling program included:

Test the main NE-trending structural corridor where the highest gold in soil coincides with IP

chargeability; and

Test the Kocakısla area for any hidden porphyry Cu-Au and/or epithermal gold potential.

10.3 West Kestane Drilling

HD-56, HD-64, and HD-69 were drilled to the west of Central Zone at Kestane to test bulls-eye high

magnetics anomalies and deep high chargeability targets.

HD-56 intersected disseminated pyrite in hornfels. HD-64 was located 1 km west of Kestane and

intersected narrow high grade Zn (1.90 m @ 16.76% Zn) mineralization with anomalous gold and Cu

(see Table 10.4). The skarn mineralization occurs at the contact between hornfels and recrystallized

limestone. Another garnet skarn zone was intersected by historic hole HD-03 at the NW edge of

Kestane. It was recommended to drill test the area between HD-03 and HD-64 for possible economic

skarn mineralization. HD-69 intersected chlorite-altered porphyritic volcanics and returned no significantmineralization.

Table 10.4: HD -64 Skarn Zone Assays

Hole IDSampleFrom(m)

SampleTo(m)

SampleInterval

(m)

Au(ppm)

Cu(%)

Zn(%)

HD-64 501.7 502.6 0.9 1.002 1.464 1.01

HD-64 502.6 504.5 1.9 0.099 0.192 16.76

HD-64 504.5 506 1.5 0.013 0.045 2.48


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11 SAMPLE PREPARATION, ANALYSES, AND SECURITY

11.1 Sample Preparation and Analyses

Collars were set up under the direct supervision of Truva Bakır staff and were drilled with HQ and PQ

diameter core. The holes were reduced to NQ when problems were encountered due to difficult ground

conditions and/or thick fault zones. Core was placed in plastic boxes with depth markers for every drill

run of up to 3 m.

At the Halilağa Project, core recoveries are considered by the authors to be good and within tolerance

to include in a resource estimate.

Figure 11.1 and Figure 11.2 show the scatter plots for Recovery% versus Cu and gold grade,

respectively. There does not appear to be any bias toward high grade versus low grade recoveries. Drillholes HD-41, HD-44, HD-46, and HD-94A returned relatively low recoveries in relation to other drill

holes. These holes are within the central volume used for resource estimation. The grades were visually

checked and compared against adjacent data and no issues were identified. Other holes with low

recoveries were identified but were outside the resource area. Recoveries in drill holes HD-101, HD-105

and HD-105A were relatively low in comparison to other holes. For holes HD-108 and newer, 4% or 157

samples returned recoveries of <50% and 10% (or 514 samples) had recoveries between 50% and

80%. It is the author’s opinion that the recoveries are very good overall and do not pose an issue.

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Figure 11.1: Core Recovery vs. Copper Grade (%)

Source: Gray and Kirkham, 2012

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Figure 11.2: Core Recovery vs. Gold Grade (g/t)

Source: Gray and Kirkham, 2012

Boxes were securely sealed and brought by truck to the core facility at the exploration camp in Etili

once a day by the drilling company or Truva Bak ır staff. Surveys were taken with a Reflex tool at 50 m

intervals down-hole to provide control on drill hole orientation.

At the core handling facility, drill holes were logged by Truva Bak ır geologists recording observations

using the Anaconda method (Einaudi 1997) and then entered into the database using Acquire®

software. Prior to logging, the geologist and the field technicians performed the following tasks:

Inspected core boxes;

Recorded missing boxes and footage errors;

Replaced footage markers with clean, clear markers;

Digitally photographed all boxes;

Recorded rock quality designation (RQD) and core loss; and

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.

(dry weight)

(dry weight) – (weight in water)

Logged core; information included engineering comments regarding the competency of core

and a fracture analyses that included quantitative measurements of primary fractures, gougematerial, veins, and dominant fracture patterns.

Specific gravity (SG) measurements from drill cores were routinely carried out for both oxide and

sulphide mineralization. Pieces (10 cm-20 cm) of solid core were used for SG measurements. Both

mineralized and un-mineralized zones were measured. The solid core was cleaned and washed and

dried in an oven. Samples were dried at 105°C for 8-12 hours and then weighed in air (dry weight).

They were then coated with paraffin, which was allowed to dry, and then the samples were reweighed

in water.

The SG values were identified by the following formula:

SG=

11.2 Sample Preparation

Sample preparation and analysis of core from the first 35 drill holes at Halilağa was conducted by

independent, ISO certified ALS Minerals (ALS). In addition, samples from three early reverse circulation

holes were similarly prepared and analyzed by ALS. Since October 18, 2009, or from drill hole HD-36

onward, the drill core samples were prepared and analyzed by independent, ISO certified AcmeAnalytical Laboratories (Acme).

Preparation at ALS Minerals

Before October 2009, all core samples were prepared at the on-site preparation lab at Truva Bak ır’s

Etili camp which was set up, managed, and run by personnel from the ALS Minerals sample preparation

facility in Izmir, Turkey (Cook, 2007). Samples were weighed on arrival, dried, crushed in their entirety

to at least 70% passing 2 mm using a TM Terminator crusher, and riffle splitting to 1,000 g. This amount

was pulverized to 85% passing 200 mesh in a TM Max2 pulverizer and disk mill. The master pulp

remained in Turkey while an approximately 100g pulp packet was forwarded to the ALS laboratory in

North Vancouver, Canada for analysis and assay.

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Preparation at Acme Labs

Core samples were prepared at Acme Lab’s sample preparation facility in Ankara, Turkey, from October

2009 onward. The R200-1000 sample preparation method were used, which includes barcoding and

weighing of the sample on arrival, drying, coarse crushing of the entire core sample to at least 80%

passing 10 mesh (2 mm) in a TM Terminator, riffle splitting to approximately 1,000 grams, and

pulverizing this material in a LM-2 carbon steel disk mill to 85% passing 200 mesh. An approximately

100 g pulp packet was forwarded by independent transport to Acme’s main laboratory in Vancouver,

Canada for analysis and assay, with the remaining master pulp material for each sample staying in

Ankara, Turkey. Rejects and pulps were stored either on-site at the Etili core shack or in Ostim, Ankara,

Turkey. Documentation of preparation, analysis, and QA/QC protocols for the Halilağa Project was

provided by Cook and Houle (2009).

In mid-2011, Acme Labs opened a new laboratory in Ankara, and since April 18, 2011 all samplepreparation and subsequently gold fire assay work, has been carried out at this location.

11.3 Assay and Analysis

Assay and Analysis at ALS Minerals

Core samples from the first 35 drill holes at Halilağa (up to HD-35) were analyzed at the ALS laboratory

in North Vancouver, Canada. Core samples were analyzed by a general aqua regia-digestion

geochemical suite, a gold fire assay technique, a copper assay technique and, where warranted, a

sequential copper procedure to determine supergene copper.

The following is a synopsis of the specific methods used:

Geochemical Analysis – ME ICP41 suite (aqua regia digestion/ICP-ES). This method was

originally employed to screen samples for copper assay, prior to assaying all samples within the

deposit.

Gold fire assay – Au-AA24 method. This lead-collection fire assay technique uses a 50 g

sample split size with AAS finish, and has a stated detection limit of 0.005 ppm.

Copper Assay – Cu-AA62 method. This four-acid digestion/AAS method is a total copper assay

method. The initial base metal assays of the first 35 drill holes by ALS were for copper assay

only. No molybdenum, iron, or silver assays have been conducted on these samples.

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Sequential Copper Assay - These were conducted, where warranted, on all core samples with

visible supergene copper mineralization. The method includes three sequential leaches,including the total copper assay of the residual component, and a total copper assay on a

separate sample split. Sequential copper assays were also applied to the full length of the first

eight Halilağa drill holes. The process includes the following procedures, the first three of which

are conducted on a single sample split:

o AA06 sulphuric acid leach;

o AA16s cyanide leach;

o AA62s total copper on the residual (4-acid digestion/AAS); and

o AA62 total copper on a separate sample split (4-acid digestion/AAS).

Assay and Analysis at Acme Labs

The Halilağa core samples were analyzed at Acme Analytical laboratories in Vancouver, Canada,

beginning with HD-36. Core samples were analyzed for a general aqua regia-digestion geochemical

suite to provide information on deleterious and pathfinder elements, a gold fire assay technique, a base

metal assay technique and, where warranted, a sequential copper procedure for to determine

supergene copper. The following is a synopsis of the specific analytical methods used:

Geochemical Analysis – Group 1DX 01 multi-element suite (aqua regia digestion/ICP-MS). This

method provides low-level concentrations of aqua regia-digestible commodity, deleterious and

pathfinder elements such as As, Sb, Hg, Bi, and S on a 0.5 g sample weight;

Gold Fire Assay – Group 3B01 method. This lead-collection fire assay technique uses a 30 g

sample split with ICP-ES instrumental finish. The method, which has a stated detection limit of

2 ppb, is suitable for gold concentrations < 10 g/t. As noted, since mid-2011, all gold fire assay

work has been carried out at Acme’s new laboratory in Ankara;

Base Metal Assay – Group 7TD02 method. This base metal total assay method uses a strong

four-acid digestion followed by an ICP-ES instrumental finish to provide data for 23 elements

including Cu, Mo, Ag, and Fe; and

Sequential Copper Assay – Group 9 04 method. These were conducted, where warranted, on

all core samples with visible supergene copper mineralization. The method includes three

sequential leaches, including the total copper assay of the residual component, and a totalcopper assay on a separate sample split. All sequential copper methods were conducted at the

Acme Laboratory in Santiago, Chile.

In Acme-Santiago, the Group 8TD method (4-acid digestion/AAS) was used for total copper assay in

place of the usual Group 7TD method. The Acme Labs sequential leach process includes the following

procedures, the first three of which are done on a single sample split:

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Group 9 04 sulphuric acid leach;

Group 9 04 cyanide leach;

Group 9 04 total copper on the residual (Group 8TD 4-acid digestion/AAS); and

Group 8TD total copper on a separate sample split (4-acid digestion/AAS).

Acme laboratories operate according to the guidelines set out in ISO/IEC Guide 25 – “General

requirements for the competence of calibration and testing laboratories”.

When sample shipments are received by the laboratory, an emailed notification is sent. Throughout the

program, no problems were encountered during transport.

Note that Total Sulphur by Leco analyses had not been routinely carried out on Halilağa core at either

ALS or Acme Labs. Leco Sulphur has been determined on only a single drill hole namely, HD-02. Alsoa small number of low-concentration core samples from earlier drill holes (i.e. 19 samples from ALS,

and 1 from Acme Labs), originally analyzed for Cu and other elements, were done by aqua regia/ICP-

ES or ICP-MS analysis only and have since been re-analysed for Cu and Mo. Following

recommendations, Group 2A15 – Sulphide, Total sulphur minus sulphate, were performed on

subsequent samples.

Reverse Circulation Drilling Sampling

All reverse circulation drilling samples were subjected to quality control procedures that ensured best

practice in the handling, sampling, analysis, and storage of the drill samples

Reverse circulation (“RC”) samples were collected and split using a 24-slot rotary splitter at the drill siteand then sealed in plastic bags. Samples were collected continuously at 1.0 m-1.5 m intervals. The

splitter was cleaned between each sample with a compressed air hose. The RC drill samples were

taken and kept under constant supervision by Truva Bakır personnel.

There are no known factors related to drilling and sampling that would materially impact the accuracy

and reliability of the results in the opinion of the author.

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11.4 Quality Assurance and Quality Control Programs

Core boxes were securely sealed and brought, by truck, to the core facility at the Etili camps once a day

by the drilling company or Truva Bakır. Here they were logged, cut, bagged, tagged and stored prior to

being shipped to the analytical laboratory. All core samples were prepared at the ALS sample

preparation laboratory located at the Etili camp and processed within the secure confines of the camp

prior to the pulp packets being transported by commercial air carrier to the ALS laboratory in North

Vancouver, Canada.

Following the switch to Acme Labs in late 2009, all Halilağa core samples were first trucked from Etili to

the Acme preparation facility in Ankara by independent transport. Pulp packets were subsequently

transported by commercial air carrier to Acme Labs in Vancouver, Canada, for assay and analysis.

Quality assurance and quality control measures used on the Halilağa Projects were employed at allstages of work in the core shed, the sample preparation facility, and in the analytical laboratory.

Evaluation of QA/QC results was done systematically and promptly to ensure that only the best quality

data was entered into the project database. Umpire, or external check, assays have been carried out as

a further means of data verification. At all times this work, whether in the field, the lab, or the exploration

office, was consistent with best practices currently in use in the mineral exploration industry.

Core Shed Procedures

Quality control measures and data verification procedures were applied to the acquisition of drill data

and to the sampling of core. The following procedures were employed to ensure quality assurance and

quality control:

Standardized sample booklets were used at all times. All booklets were marked up, prior to use,

with the standards, field blanks, and core duplicates locations clearly defined;

Standards and field blanks were entered every 20th sample. Core duplicate samples (1/4 core)

are entered into the sample flow, at the discretion of the geologist, approximately every 20

samples;

All holes were sampled from top to bottom, with most samples being taken every 2 m;

For each sample interval, all required entries (from-to) of the standard sample card were filled

in and half of the sample number tag was placed at the starting point of the sample interval in

the core box; The second half of the tag was put into the sample bag (labelled on both sides with the sample

number) by the splitters when they were taking the sample;

The beginning of a sample is clearly marked with a black marker, and a line perpendicular to the

core; and

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The geologist’s double-checked the samples once they were cut, and to verify that all samples

were properly labelled, with the sample tags inside of the sample bags.

At the Halilağa Project, inserting quality control samples took place in the core shack before samples

were shipped to the lab. These samples were routinely inserted and used to check for accuracy,

precision and cleanliness in the analytical laboratory. At the beginning of sampling, sample tags were

pre-marked before logging with locations for standards, core duplicates, and field blanks.

The process was as follows:

Core duplicate samples are taken every 20 samples within the sample series (5%). Core

duplicate samples are used to evaluate combined field, preparation and analytical precision.

The core duplicate samples are quarter-spilt cores sampled on-site before the samples leave

camp;

Field blanks comprised non-mineralized limestone material collected from a local source,

broken with a hammer, and inserted into the sample series every 20 samples (5%). Field blanks

are inserted to test for any potential carry-over contamination which might occur in the crushing

phase of sample preparation, as a result of poor cleaning practices;

Standards are used to test the accuracy of the assays and to monitor the consistency of the

laboratory over time. Commercially-available Cu-Au assay standards were purchased from

CDN Resource Laboratories Ltd., Vancouver, Canada. These standards are inserted into the

sample sequences approximately once every 20 samples (5%). The identities of the standardsare blind to the laboratory; and

Pulp blanks, purchased commercially from CDN Resource Laboratories Ltd., are also inserted

in the analytical suites in order to assess any potential carry-over contamination in the analytical

lab. The pulp blanks are not involved in sample preparation, and do not test the preparation

stage of the sample stream as the field blanks do.

The author has relied on data and information related to quality assurance and quality control that has

been prepared by a QP as defined by the requirements of NI 43-101. The QA/QC data was prepared

and evaluated by Truva Bakır personnel.

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Commercial standards sourced from CDN Resource Laboratories Ltd., Vancouver, Canada, were used

to test the accuracy of the assays and to monitor the consistency of the laboratory over time. Allstandards listed here are Cu-Au standards with recommended values (between-lab mean ± 2 standard

deviations) for copper, gold, and in some cases, molybdenum. These standards were randomly inserted

into the sample sequences approximately once every 20 samples. Table 11.1and Table 11.2 show the

standards used for the Halilağa Project in 2009 and 2010, along with their recommended mean gold

and copper concentrations.

Table 11.1: Recommended Gold Concentrations of Standards Used at Halilağa

Standard ID Au (ppb)

CDN-CGS-7 950

CDN-CGS-16 140

CDN-CGS-15 570

CDN-CGS-12 290

CDN-CGS-11 730

CDN-CGS-10 1730

Source: Teck Madencilik 2010

Table 11.2: Recommended Copper Concentrations of Standards Used at Halilağa

Standard ID Cu (%)

CDN-CGS-10 1.55

CDN-CGS-11 0.683

CDN-CGS-12 0.265CDN-CGS-15 0.451

CDN-CGS-16 0.112

CDN-CGS-7 1.01

Source: Teck Madencilik 2010

Sample Preparation Lab Procedures

Crusher duplicate samples were prepared in the sample preparation labs by laboratory personnel, at

Truva Bakır’s request, and inserted into the laboratory sample batches for subsequent assay and

analysis. Crusher duplicates were prepared after the crushing stage, but prior to the pulverization stageof sample preparation. Their purpose was to test the combined preparation and analytical precision in

the laboratory, and was designed to measure the quality of sample preparation.

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Analytical Laboratory Procedures

In addition to standards, core duplicates, field blanks, and pulp blanks which were inserted in the core

shed, and the crusher duplicate samples which was prepared in the prep lab, laboratory pulp duplicate

samples were also prepared during final sample analysis. In this case the pulp duplicates were part of

the laboratory’s (ALS and Acme Labs) own internal QA/QC procedures; these results were reported by

the labs and subsequently captured as a measure of analytical precision at the final digestion and/or

fusion and instrumental stages of analysis.

Umpire (External Check) Assays

A protocol has been initiated to send 5% of all assayed sample pulps to a second laboratory for

analysis. Prior to 2009, ALS was the main lab at the project with Acme Labs providing external check

assays. In 2009, Acme became the main lab with ALS stated to provide external check assays. Umpire,or external check, assays are an additional means of data verification and are useful in identifying any

substantial biases between laboratories, which may have been introduced during the course of the

project. Although 5% external checks are part of the QA/QC protocol, in reality much less than that has

been completed.

Correlation plots are shown for Hole HD-02 between the two laboratories for copper (Cu), gold (Au) and

molybdenum). Figure 11.3 shows the umpire, external checks, for Cu geochemistry (ALS;ME-ICP41

and Acme;Group 1DX) and Cu assay (ALS;Cu-OG-62 and Acme;Group 7AR) which illustrates an

excellent correlation between the two laboratories.

Figure 11.4 shows the umpire, external checks for gold assay (ALS;Au-AA24 and Acme;Group 3B).

With the exception of a small number of outliers, the data between the two laboratories shows goodagreement.

Figure 11.5 shows the umpire, external checks, for Mo geochemistry (ALS;ME-ICP41 and Acme;Group

1DX) which illustrates a good correlation between the two laboratories.

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Figure 11.3: Umpire Check Assays for Drill Hole HD-02; Cu (%)


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Figure 11.4: Umpire Check Assays for Drill Hole ID HD-02; Au (ppm)


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Figure 11.5: Umpire Check Assays for Drill Hole HD-02; Mo (%)


Heberlein Geoconsulting Ltd. was contracted to audit the quality control procedures and data for the

Halilağa Project which resulted in the report “Halilağa Project; An Audit of Quality Control Procedures

and Data for Teck Resources Inc.” dated July 20, 2012 (Heberlein 2012). This audit has shown that the

quality control procedures employed at Halilağa are consistent with ”Industry Best Practices” as

required under NI 43-101.

Evaluation of QA/QC Results

Results for standards, field blanks, pulp blanks, and various types of duplicate samples were reviewed

and evaluated quickly after the batch results were received. Several types of control charts are used to

plot and monitor the data. Decisions were made immediately, based on simple criteria, as to whether or

not the quality control results were acceptable.

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Standards

Failure of a standard implied that all routine samples within its sphere of influence were also considered

to have failed, and must be re-analyzed at the same primary laboratory. Two criteria for the passing and

failing of standards results have been used on the Halilağa Project:

Standards were considered to have failed if the reported copper or gold assay concentration

was greater or less than three standard deviations from the recommended mean value for that

standard; and

Standards were also considered to have failed if the reported copper or gold assay

concentration was greater or less than two standard deviations from the recommended mean

value for that standard, in two or more successive insertions.

In the case of failure of any standard, the procedure was to re-assay the block of samples within itssphere of influence. In practice, this meant that all consecutively listed samples, down list from the

failing standard to the next passing standard, and up list from the failing standard to the next prior

passing standard, were considered to have failed, and were to be re-assayed.

Upon receipt of re-assay data, the standards were re-evaluated, and if they passed the data is similarly

deemed to have passed. At this point, the failed data was ejected from the project database and

replaced with the new, passing data.

During the significant drilling campaigns of 2009-2011, the copper standards had a failure rate of

between 1% - 3%, and molybdenum standards had a failure rate of 1%. However, the gold standards

exhibited a failure rate of between 10% –11% for 2009, 2010, and 2011 which is very high; however,these are a result of two consistently failing single gold standards: CDN-CGS-15 and CDN-CGS-16 and

due to the adoption of a low failure threshold. It is important to note that the majority of failures were

attributed to these failing standards. In addition, note that when a standard fails, the complete assay

run, before and after the fail, was re-run until the standard passes.

The overall failure rate at ALS was 19 in 733 analyses, or 2.6% as shown in Table 11.3. This compares

to the Acme rate of 126 in 1,864 analyses, or 6.8% as shown Table 11.4. The gold by fire assay failure

rate was 3.2% and 6.3% for ALS and Acme, respectively.

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Table 11.3: Standards Performance for the Halilağa Project (2004 - 2009)

STD-ID NNo of Cu%by ME41

(ICP)

No of Cu%By AA62b

(AA)

No. of Auppm byAA24(FA)

No. of Mo %by ME41

(ICP)

Cu Limits(% Cu)

Mo Limi(% Mo

CDN-CGS-10 80 30 (0) 19 (0) 31 (1) 30 1.505-1.955 NA

CDN-CGS-11 95 37 (2) 19 (0) 39 (0) 37 0.644-0.722 NA

CDN-CGS-12 104 42 (1) 16 (0) 46 (6) 42 0.243-0.288 NA

CDN-CGS-7 87 34 (0) 18 (0) 35 (2) 34 0.905-1.115 NA

CDN-CGS-8 82 30 (0) 19 (0) 33 (3) 30 0.093-0.117 NA

CDN-CM-3 32 1(0) 0(0) 30 (0) 1(1) 0.517-0.580 0.025-0.0

CDN-GS-1P5 25 24 14 25 (0) 24 NA NA CDN-GS-2A 30 29 28 30 (0) 29 NA NA

CDN-GS-2B 6 6 4 6 (0) 7 NA NA

CDN-GS-2C 23 23 0 23 (0) 0 NA NA

CDN-GS-5B 57 54 48 57 (0) 54 NA NA

CDN-GS-P3 58 57 51 58 (1) 57 NA NA

CDN-GS-P5 50 50 41 50(0) 50 NA NA

CDN-GS-5C 4 3 0 4(2) 3 NA NA

733174 (3) 91 (0) 467(15) 1 (1)

1.80% 0% 3.20% 100%


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Table 11.4: Standards Performance for the Halilağa Project (2009 – 2012)

STD-ID NNo of Cu by

1DXNo of Au by

3BNo of Cu by

7TDNo of Moby 7TD

No of MoBy 1DX

Cu% Limits Mo%

CDN-CGS-10 113 3 (3) 56 (2) 54(1) 53 3 1.505-1.955 N

CDN-CGS-11 120 11 (7) 56 (3) 53 (1) 3 11 0.644-0.722 N

CDN-CGS-12 133 48 (5) 69 (6) 16 (0) 16 48 0.243-0.288 N

CDN-CGS-15 221 47 (25) 96 (8) 78 (1) 76 47 0.421-0.481 N

CDN-CGS-16 465 42 (18) 233 (20) 190 (3) 184 42 0.104-0.112 N

CDN-CGS-7 12 4 (0) 4 (0) 4 (0) 4 4 0.905-1.115 N

CDN-CGS-21 166 0 (0) 84 (2) 82 (2) 80 0 1.174-1.426 N

CDN-CM-2 209 0 (0) 73 (6) 70 (4) 66(1) 0 0.949-1.078 0.026CDN-CM-3 335 0(0) 117 (5) 110 (3) 108(0) 0 0.517-0.580 0.025

CDN-CM-5 90 0(0) 30(0) 30(0) 30(0) 0 0.289-0.349 0.043

Total 1,864155(58) 818(52) 687(15) 204(1)

37.40% 6.30% 2.20% 0.50%


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There were a relatively high failure rates for the gold standards which appeared to be attributed to two

particular standards. However, any failures were recorded and the lab requested the complete job bere-run. Additional check assays were also sent to the ALS Chemex Lab in North Vancouver as a further

means of data verification. An internal review of the QA/QC procedures of the lab and Truva Bak ır's

protocols was performed in 2012 and is discussed below. Subsequently, alterative gold standards were

sourced and used for the 2012 program.

In 2012, a total of 161 standards for gold (lead fire assay) and copper (4-Acid Digestion), 89 for

molybdenum along with 33 standards for molybdenum and sulphur were analyzed during the 2012 drill

programs as shown in Tables 11.5, 11.6 and 11.7, respectively.

Table 11.5: Standards Performance for the Halilağa Project (2012) - Au (Pb Fire Assay)

Standards Au Std value ± 2 std dev Sent Failure

CDN-CGS-16 0.14 ± 0.046 g/t 20 1

CDN-CGS-27 3 0

CDN-CM-2 1.42 ± 0.13 g/t 3 0

CDN-CM-3 0.46 ± 0.06 g/t 9 1

CDN-CM-5 0.294 ± 0.046 g/t 18 0

CDN-CM-13 0.740 ± 0.094 g/t 11 0

CDN-CM-16 0.294 ± 0.046 g/t 31 0

CDN-CM-20 0.278 ± 0.044 g/t 17 0

CDN-GS-P2A 1 0

CDN-GS-P3B 1 0OREAS-501 0.204 ± 0.022 g/t 47 1

TOTAL 161 3

Source: JDS 2014

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Table 11.6: Standards Performance for the Halilağa Project (2012) - Cu (4-Acid Digestion)

Standards Cu Std value ± 2 std dev Sent Failure

CDN-CGS-16 0.112 ± 0.005 % 20 6

CDN-CGS-27 3 1

CDN-CM-2 1.013 ± 0.043 % 3 0

CDN-CM-3 0.548 ± 0.021 % 9 1

CDN-CM-5 0.319 ± 0.020 % 18 0

CDN-CM-13 0.786 ± 0.036 % 11 1

CDN-CM-16 0.184 ± 0.014 % 31 0

CDN-CM-20 0.316 ± 0.016 % 17 0

OREAS-501 2708 ± 0.0164 ppm 47 0

TOTAL 159 9

Source: JDS 2014

Table 11.7: Standards Performance for the Halilağa Project (2012) - Mo (4-Acid Digestion)

StandardsMo Std value ± 2 std

devSent Failure

CDN-CM-2 0.029 ± 0.002 % 3 0

CDN-CM-3 0.029 ± 0.003 % 9 0

CDN-CM-5 0.050 ± 0.005 % 18 0

CDN-CM-13 0.044 ± 0.004 % 11 0

CDN-CM-16 0.016 ± 0.002 % 31 0

CDN-CM-20 0.030 ± 0.002 % 17 0

OREAS-501 59 ± 4.2 ppm 33 8

TOTAL 122 8

Source: JDS 2014

The OREAS molybdenum - sulphur standards were chosen to use in data verification however it is clear

that these standards cannot be relied upon and should not be used in the future due to the extremely

high failure rate.

The author considers that the adequacy of sampling, security, and analytical procedures is satisfactory.

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Field Blanks

Field blanks are used to check the level of cleanliness at a laboratory, and to more specifically check for

the presence of any carry-over contamination during the crushing phase of sample preparation. Proper

cleaning of the coarse crushers between samples, and between sample batches, should ensure that

there is no carry-over of material between samples, producing negligible gold and Cu results on a

consistent basis. Field blanks are typically created from barren rock material, preferably of hardness

similar to the target lithologies. At Halilağa, non-mineralized limestone material collected from a local

source, broken with a hammer, and inserted into the sample series every 20 samples (5%). In addition

to the field blanks, commercial CDN Resource Laboratories Ltd. pulp blanks were randomly inserted

into the sample series every 20 samples as a check on any possible carry-over contamination in the

analytical laboratory.

In general, field blanks exhibited a failure rate of 1% for each of the years, which indicated that carry-over contamination at the crusher stage has not been a problem on this project.

Duplicate Samples

Both core-duplicate and crushed-duplicate samples are created for insertion in Halilağa core assay

batches. The core duplicate samples are 1/2 split of core which are cut on-site before the sample

leaves camp. Acme Labs has also prepared coarse-crusher duplicates and inserted these at a rate of

one every 20 samples. Core-duplicate samples are taken every 20 samples within the sample series.

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12 DATA VERIFICATION

12.1 Geological Site Visit

Several of the other named QP’s have visited the site as outlined in Section 2.5, however with respect

to the geology, Garth Kirkham, P. Geo., visited the property between August 13 and August 16, 2011.

The site visit entailed inspection of the camp, accommodations, core logging facilities, offices, active

drill sites, outcrops, historic drill collars, core storage facilities, core receiving area, core sawing stations,

and tours of major centres and surrounding villages most likely to be affected by any potential mining

operation.

The tour of the offices, core logging and storage facilities showed a clean, well-organized, professional

environment. On-site staff led Kirkham through the chain of custody and methods used at each stage of

the logging and sampling process. All methods and processes are to North American, industrystandards and no issues were identified

Kirkham also visited two active drill sites. One of the drills was down due to mechanical issues.

However, the second drill was observed and the drill site processes and procedures were

demonstrated. The drill sites are clean and well-maintained. The process of extracting the core and

core boxing is standard practice. The boxes are sealed in plastic core boxes, labelled with permanent

marker, and sealed securely with bailing wire. The boxes are loaded into the drillers’ truck or into the

geologist’s vehicle and taken to the camp.

A visit to collar locations showed the collars were well marked and labelled, therefore easily identified.

The earlier drill holes were cased and permanently cemented.

Four complete drill holes: HD-01, HD-08, HD-92A, and HD-94A were selected at random from the

database by Kirkham and laid out at the core storage area. Site staff supplied the logs and assay

sheets for verification against the core and the logged intervals. The data correlated with the physical

core and no issues were identified. In addition, Kirkham toured the complete core storage facilities

pulling and reviewing core throughout. No issues were identified and recoveries appeared to be very

good to excellent.

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Kirkham is confident that the data and results are valid based on the site visit and inspection of all

aspects of the project, including methods and procedures used. It is the opinion of the independentKirkham that all work, procedures, and results have adhered to best practices and industry standards

required by NI 43-101. At the time of the 2011 site visit, no duplicate samples were taken to verify assay

results, but Kirkham is of the opinion that the work is being performed at that time was by a well-

respected, large, multi-national company that employed competent professionals that adhered to

industry best practices and standards.

Kirkham has not completed a current site visit subsequent to the 2012 work program. The author

believes that the above remains valid, as no exploration work has occurred on site during the period of

2012-2014, however; personal validation has not been completed as at the effective date of this report.

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13 MINERAL PROCESSING AND METALLURGICAL TESTING

13.1 Metallurgical Testing

The existing metallurgical test work is considered suitable for this level of study and the comminution

data is considered adequate for a conceptual milling circuit design. The design of the processing

circuits is based on test work data in conjunction with reasonable assumptions based on typical industry

values when insufficient data were available. The Halilağa mineralized material is of moderate

competency and hardness, and amenable to grinding in a conventional SAG-ball milling circuit (SABC).

The copper mineralogy is fine grained and test work indicates a requirement to re-grind to a fine particle

size to achieve adequate liberation for cleaner flotation.

Test work Programs

The following list of metallurgical and mineralogy test reports were reviewed for this study:

G&T Metallurgical Services Ltd., Preliminary Metallurgical Assessment of Samples from the

Halilağa Project KM2020, 11th July 2007;

G&T Metallurgical Services Ltd., Mineralogy Testing on Samples from the Halilağa Project

KM2157, 1st April 2008;

G&T Metallurgical Services Ltd., Metallurgical Testing KM2908, 15th April 2011; and

ALS Metallurgy Kamloops. Metallurgical Testing on Samples from the Halilağa Deposit

KM3897, 26th September 2013.

The first test program (KM2020 - G&T 2007) tested composites made from two drill holes HD-01 and

HD-04 for preliminary flotation to produce copper concentrates and 25 variability rougher flotation tests.

The second report (KM2157 – G&T 2007) covers mineralogy speciation on 25 samples that were tested

in the first test program. In the majority of samples tested copper occurs mainly as chalcopyrite. Five

of the samples, which contain higher than average copper content (supergene zone), have copper

occurring mainly as chalcocite and covellite.

For the third test program (KM2908 - G&T 2011), 15 samples were composited from drill holes HD-40

and HD-49. A master composite called “Typical Ore” was then formed from the 15 composites. (The

use of the term “Ore” represents the common name for plant feed material used to identify a sample but

in the case of this project, does not refer to mineral reserve material). Bond Ball Mill Work Index

comminution testing was performed on six of the composite samples and all 15 composite sampleswere tested for preliminary flotation to produce copper cleaner concentrates. Two locked cycle tests

were performed on the “Typical Ore” composite.

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For the fourth test program (KM3897 - ALS 2013), three master composite samples were compiled,

representing material types to be mined in Year 1 (supergene/hypogene – covellite, chalcocite andchalcopyrite mineralization), Year 2-5 and Year 6-14 (hypogene - chalcopyrite mineralization). These

composite samples were specifically prepared for locked cycle flotation testing to produce

representative cleaner tails material, for subsequent cyanide leach testing and analysis.

The location of the metallurgical samples and drill holes are considered to be reasonably representative

of the deposit for a PEA study stage given the nature and continuity of rock types and mineralization of

the deposit.

Where no test work data are available, reasonable assumptions, based on operating data or test work,

from other projects, has been used to develop the process design criteria.

G&T July 2007 Test Program (KM-2020)

The first test program conducted in 2007 was based on sample intervals from two drill holes. Bench

scale flotation testing was conducted on two master composites as well as 25 discrete variability

composites.

The composites were made up of the reject samples from two drill holes designated HD-01 and HD-04.

The location of these two drill holes in the economic mineralized zone, relative to the proposed pit

outline, is shown in Figure 13.1. HD-01 comprised intervals from 23.85 m in depth through 200.6 m and

HD-04 comprised intervals from 22.3 m in depth through 200.8 m.

The 25 discrete composite samples were prepared by combining individual intercepts of the two drill

holes with a target total composite weight of 20 kg, 13 composites from drill hole HD-01and 12composites from drill hole HD-04.

The two master composites were made by taking 2 kg from each of the 25 variability composites and

are designated Master HD-01 and Master HD-04. The variability composite sample numbers are listed

in Tables 13.3 and 13.4.

The two main phases of G&T 2007 metallurgical testing undertaken on the Halilağa samples were:

Bench scale flotation testing, rougher and cleaner testing on master composites; and

Bench scale flotation testing consisting of rougher flotation to map the variability of metallurgical

performance, supplemented with mineralogical examination.

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Figure 13.1: Location of Drill Hole

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Flotation Test Work Results

Flotation test work commenced with rougher and cleaner open circuit tests on the two major composites

and rougher flotation on the 25 variability composites. Copper sulphide flotation was undertaken to

produce a copper concentrate. Analysis of the test work constitutes a part of the information used in the

plant process design criteria and estimates of concentrate copper grade, and copper and gold recovery.

The 2007 flotation test work program was divided into three phases:

Rougher flotation on the two master composites, HD-01 and HD-04.

Open circuit cleaner flotation on the two master composites; and

Rougher flotation on the 25 variability composites.

Rougher Flotation

Initially, a series of rougher kinetic flotation tests were conducted on the master composites to

determine the sensitivity of the mineralized materials to grind size and reagent scheme. These tests

were supplemented with mineralogical examination (Second G&T Metallurgical Report KM-2157) on the

two master composites by modal analysis to determine fundamental mineral liberation and mineral

speciation.

These tests indicated the following:

The sulphide minerals occur in order of abundance as pyrite, chalcopyrite and

chalcocite/covellite;

Rougher flotation chalcopyrite liberation appears to be relatively insensitive to primary grind size

in the range of 100 µm and 200 µm;

Overall copper sulphide liberation should be adequate to achieve good metallurgical

performance in the rougher flotation circuit. A target primary grind size of 150 µm was

established through rougher kinetic testing at a primary grind size of 100 µm, 150 µm, and 200

µm; and

Mineralized material hardness increases with depth in the drill hole.

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The results from the rougher tests are summarized in Table 13.1.

These tests indicate that the mineralized material is amenable to flotation, resulting in good recovery of

target mineral species into a low mass concentrate stream using potassium amyl xanthate (PAX), lime,

and methyl isobutyl carbinol (MIBC).

Table 13.1: Rougher Flotation Results for Halilağa Master Composite Samples

Hole IDGrindSizeµm

MassRecovery

%

Feed Concentrate Grade Concentrate Recovery

% Cu g/t Au % Cu g/t Au % S % Cu % Au % S

HD01 152 14.2 0.75 0.62 4.91 3.26 26.3 93.8 75.2 97.3

HD01 152 18.5 0.73 0.66 3.83 2.91 19.9 96.6 81.6 98.5

HD01 152 15.8 0.76 0.59 4.67 3.18 24.4 96.6 84.4 98.7

HD01 94 14.2 0.76 0.64 5.18 3.82 28.5 96.5 84 98.1

HD01 192 17.2 0.74 0.68 10.3 6.83 24.8 95.9 89.1 99.4

HD04 151 11.2 0.65 0.46 4.19 2.65 15.2 71.5 64.4 54.3

HD04 151 22.3 0.65 0.51 2.79 1.98 16.5 95.7 86.7 98.7

HD04 151 15.5 0.64 0.53 4.01 2.87 22.2 96.2 84 97.5

HD04 93 15.4 0.63 0.5 3.95 2.62 22.1 96.1 79.9 97.8

HD04 186 14.9 0.63 0.46 3.98 2.58 22.7 94.3 83.4 96.4

Source: Simmons 2014

Cleaner Flotation

Cleaner flotation was conducted on the master composite samples at a target primary grind size of 80%

passing 150 μm and incorporated a rougher re-grind stage to increase mineral liberation. Varying

re-grind times, reagent dosages, and pH in cleaners were tested to determine optimum flotation

conditions. The cleaner flotation reagent scheme was not changed from that tested in the rougher tests.

Re-grinding of rougher concentrates would be required to achieve adequate cleaner concentrate

grades, as indicated by: 1) the fine-grained structure of the mineralized material identified by the Modal

Analysis, and 2) the increase in rougher grade and recovery with finer re-grind. Concentrate re-grinding

was therefore incorporated in all subsequent cleaner testing. Multiple stages of cleaning were

incorporated to target higher concentrate grades, typically with an elevated pH level in the cleaning.

The results from optimization tests are summarized in Table 13.2 and show the relationship of copper

recovery versus grade and copper recovery versus gold recovery for both master composites.

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The tests where rougher concentrate was re-ground to approximately 20 µm achieved the best copper

grades at 31.2% Cu to 39.2% Cu and the best recoveries of 89.7% and 91.1% from master compositesamples from HD-04 and HD-01 respectively. This is clearly illustrated in Figure 13.2 and Figure 13.3.

Gold recoveries ranged from 53.5% to 71.3%.

Table 13.2: Cleaner Flotation results for Halilağa Master Composite Samples

Hole IDGrindSizeµm

MassRecovery

%



HD01 55 3 0.73 0.6 17.5 11.3 39.9 72.5 57.2 27

HD01 47 2.4 0.74 0.52 26.5 12.9 39.3 87 59.7 24.9

HD01 22 1.8 0.77 0.52 39.2 18.84 32.2 91.1 65 14.8

HD04 46 2.7 0.63 0.46 17.5 10.2 42.4 74.4 58.8 33

HD04 40 1.7 0.63 0.46 26.4 13.54 41.2 69.3 53.5 19.3

HD04 23 1.8 0.63 0.45 31.2 17.53 28.4 89.7 71.3 14.4


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HALILAĞA

Effective Date

.

Figure 13.

Source: Sim

PRO J ECT

: December 20,

: Master Co

mons 2014

PEA TECH

2014

mposite HD

NICAL REPO

1 Copper R

RT

ougher Rec very vs. Copper Grade

13-7

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HALILAĞA

Effective Date

.

Figure 13.

Source: Sim

Variability

Twelve of

between 9

78.0% to 9

Copper-go

near the s

Figure 13.

drill hole H

The 12 sa97.1%. GoTable 13.4

PRO J ECT

: December 20,

: Master Co

mons 2014

Rougher F

the 13 sa

3.8% and 9

1.8%.

ld recovery

rface which

is a plot of

D-01.

mples fromld recovery t.

PEA TECH

2014

mposite HD

lotation Tes

ples from

.6%. Gold

nd grade fo

may be oxid

copper and

drill hole Ho rougher c

NICAL REPO

4 Cleaner

ting

rill hole H

ecovery to r

r sample C1

ized.

gold recove

-04 producncentrate f

RT

opper Reco

-01 produc

ougher conc

were low; t

ry for each

d rougherr samples C

very vs. Cle

d rougher

entrate for s

is is likely d

f the 13 var

oncentrate1 to C12 ra

ner Gold R

concentrate

amples C2 t

ue to the loc

iability comp

recoveries bged from 7

covery

copper rec

o C13 rang

ation of the

osite sampl

etween 92..5% to 92.6

13-8

overies

d from

sample

s from

% and. See

7/24/2019 2015 Halilaga




.

Twenty five samples were subject to rougher flotation variability testing. The results are summarized in

Table 13.3 and Table 13.4.

Table 13.3: Variability Rougher Flotation Testing Discrete Composites Drill Hole HD-01

Comp

GrindSize

MassRec


µm % % Cu g/t Au % Cu g/t Au % S % Cu % Au % S

HD01C1 124 10.7 1.86 0.76 11.1 3.92 15 64.1 55.6 40.6

HD01C2 134 7.6 2.4 0.9 9.06 2.9 24.8 97.4 83.5 99.4

HD01C3 121 9.3 0.68 0.56 2.68 1.98 25.3 96.1 85.7 99.5

HD01C4 132 6.8 1.03 1.08 5.34 4.92 22.7 97.6 85.7 99.3

HD01C5 140 5.2 1 1.14 6.02 6.23 24.9 97.4 88.2 99.4

HD01C6 152 4.1 0.97 1.01 4.65 4.35 28.5 94.6 85.3 98.6

HD01C7 143 12.9 0.4 0.39 1.94 1.53 27.3 95.4 78 98.7

HD01C8 136 4.1 0.25 0.24 20.9 1.36 24.8 93.8 78.9 97.6

HD01C9 155 9.7 0.36 0.45 1.93 2.05 18.7 96.4 83.5 99

HD01C10 131 7.8 0.38 0.38 2.09 1.89 17.8 96.7 86.9 100

HD01C11 136 7.6 0.21 0.19 0.97 0.86 18.9 95.1 91.8 99.3

HD01C12 136 6.12 0.2 0.21 1.2 1.03 14.6 95.8 79.5 99.1

HD01C13 116 7.3 0.16 0.13 0.84 0.63 16.3 94.9 87.4 99.2


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HALILAĞA

Effective Date

.

Table 13.4

Comp

HD04C1

HD04C2

HD04C3

HD04C4

HD04C5

HD04C6

HD04C7

HD04C8HD04C9

HD04C10

HD04C11

HD04C12

Source: Sim

Figure 13.

Source: Sim

PRO J ECT

: December 20,

: Variability

GrindSizeµm

M

146

149

136

133 1

146

150 1

144 1

139 1145

146 1

110 1

111 1

mons 2014

: Copper a

mons 2014

PEA TECH

2014

ougher Flo

assec% % C

0.1 1.15

3.9 1.85

9.7 1.15

9.7 0.72

0.5 0.51

9.6 0.51

6.6 0.37

3.4 0.1914 0.34

6.5 0.35

8.2 0.21

5.6 0.25

d Gold Rou

NICAL REPO

tation Testi

Feed

g/t Au

0.63

1.44

0.86

0.81

0.49

0.45

0.37

0.220.27

0.21

0.13

0.16

her Recove

RT

g Discrete

Conce

% Cu

3.73

7.5

3.73

3.49

2.42

2.52

2.16

1.32.35

2

1.1

1.46

ry Variabilit

omposites

trate Grade

/t Au %

1.87 15

5.59 22

2.47 19

3.5 24

2.06 23

1.91 1

1.76 16

1.26 171.59 23

1.1 14

0.56 10

0.82 11

HD-01

Drill Hole H

Con

S % Cu

.6 97.6

.3 97.2

.2 96.4

.1 95.9

.4 96.9

97.1

.1 96.1

.2 94.4

.9 96

.5 94.5

.4 94.2

.3 92.8

-04

entrate Rec

% Au

88.9

92.6

85.3

85.1

85.5

83.8

79.6

76.581.2

87.9

80.6

79.1

13-10

very

% S

99.9

99.5

99.1

98.7

99.3

99.2

97.4

96.997.8

98.4

98.5

97.1

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HALILAĞA

Effective Date

.

Figure 13.hole HD-0

Figure 13.

Source: Sim

Mineralo

G&T perfo

testing in

Digital Ima

PRO J ECT

: December 20,

shows cop.

: Copper a

mons 2014

y Examina

rmed miner

M2020. A

ging System

PEA TECH

2014

per and gol

d Gold Rou

tion (G&T 2

logical exa

summary of

) analysis is

NICAL REPO

recovery fo

her Recove

011, KM-21

ination on

the metal a

summarized

RT

r each of th

ry Variabilit

57)

25 samples

say and mi

in Table 13.

12 variabili

HD-04

that were

neralogical

.5.

ity composit

reviously su

ata from th

samples fr

bjected to fl

ADIS (Aut

13-11

m drill

lotation

mated

7/24/2019 2015 Halilaga




.

Table 13.5: Mineralogy Examination Results

SampleMetal Assay - percent Mineral Assay – Weight Percent

Cu Fe S Cp Bn Ch/Cv Py Ma He Gn

HD01-C1 1.7 3 3.7 0.2 <0.1 2.1 6 0.1 0.1 91.6

HD01-C2 2.2 4.7 5.9 0.9 <0.1 2.1 9.6 <0.1 <0.1 87.4

HD01-C3 0.7 5.9 6.2 1 <0.1 0.2 10.8 1.2 <0.1 86.7

HD01-C4 0.9 6.5 3.6 2.5 <0.1 <0.1 5.5 4 0.1 87.9

HD01-C5 0.9 7.9 3.5 1.5 <0.1 <0.1 5.7 4.2 0.2 88.5

HD01-C6 0.9 7 5.2 2.6 <0.1 <0.1 8.6 4.1 <0.1 84.8

HD01-C7 0.4 5.4 4.7 0.9 <0.1 <0.1 8.2 0.6 <0.1 90.2

HD01-C8 0.3 5.2 2.9 0.7 <0.1 <0.1 8.4 1.8 0.1 88.9

HD01-C9 0.3 5.4 2.7 1.1 <0.1 <0.1 4.3 1.8 <0.1 92.8

HD01-C10

0.3 5.4 2.2 1 <0.1 <0.1 4.9 1.9 <0.1 92.2

HD01-C11

0.2 5.1 2.8 0.6 <0.1 <0.1 7.6 1.4 <0.1 90.5

HD01-C12

0.2 4.5 1.6 0.6 <0.1 <0.1 2.6 0.7 <0.1 96

HD01-C13

0.2 4.2 2.2 0.3 <0.1 <0.1 3.8 0.6 <0.1 95.1

HD04-C1 1.1 4.6 4.6 0.3 <0.1 1.6 7.8 0.1 <0.1 90.2

HD04-C2 1.9 5.5 5.1 0.3 <0.1 2.1 8.5 <0.1 0.1 89

HD04-C3 1.1 4.8 4.9 0.7 <0.1 1.1 10.3 <0.1 0.1 87.7

HD04-C4 0.7 6 4.4 1.2 <0.1 0.3 12.7 0.3 1.3 84.2

HD04-C5 0.5 6.3 4.2 0.9 <0.1 0.1 7.3 0.5 <0.1 91.2

HD04-C6 0.5 5.8 2.9 0.9 <0.1 0.1 5.5 1.3 <0.1 92.2

HD04-C7 0.4 4.4 2.5 1 <0.1 <0.1 5.2 1.3 <0.1 92.5

HD04-C8 0.2 4 1.9 0.6 <0.1 <0.1 3.3 0.9 <0.1 95.2

HD04-C9 0.3 4.7 2.9 0.3 0.31 <0.1 8.6 1.3 <0.1 89.5

HD04-C10

0.4 3.1 2 0.8 <0.1 0.1 3.1 0.1 <0.1 95.9

HD04-C11

0.2 2.6 1.6 0.5 <0.1 <0.1 2.6 0.2 0.1 96.6

HD04-C12

0.2 2.4 1.7 0.7 <0.1 <0.1 7.3 1.3 <0.1 90.7


Cp-Chalcopyrite, Bn-Bornite, Ch-Chalcocite, Cv-Covellite, Py-Pyrite, Ma-Magnetite, He-Hematite, Gn-Non-sulfideMinerals

A series of photomicrographs, depicting the samples that contain secondary copper sulphides and

some without, are appended to the report.

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.

Flotation Test Work Results (G&T 2011, KM-2908)

The third test program conducted in 2011 was based on two shipments of samples. The first shipment

contained 203 samples weighing a total of 1.2 t. The second shipment contained 115 samples weighing

a total of 672 kg. A total of 15 composites were made using the received samples. The samples were

collected from drill holes HD-40 and HD-49.

Figure 13.6 shows the location of these drill holes in the pit outline.

A global composite designated “Typical Ore” was made by taking 4 kg from each Composite 1, 3, 8,

and 9 for a total 16 kg sample.

Two main phases of metallurgical testing were undertaken on the Halilağa mineralized samples:

Determine the mineralized material hardness of 6 of the 15 variability composites; and

Open circuit cleaner flotation testing.

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.

Figure 13.6: Location of Drill Holes HD-40 and HD-49

Comminution Test Work

The current comminution dataset consists of seven Bond ball mill work index tests, one test each for

Composites 1-7. Composites 8-15 were not tested. Table 13.6 summarizes the outcome of the

comminution laboratory test work undertaken in this test program.

7/24/2019 2015 Halilaga


7/24/2019 2015 Halilaga


HALILAĞA

Effective Date

.

Modeling

1-7 whereequation i

required t

composite

For the tw

8.8 to a hi

the real an

two drill ho

Figure 13.

Source: Sim

PRO J ECT

: December 20,

as conduct

the actualFigure 13.

produce t

8-15.

drill holes

h of 12.7 k

d modeled

les. A high

: Summary

mons 2014

PEA TECH

2014

d by plottin

(Real) BM-) from this

he same fl

sed in the

h/t. Figure

ata. The ba

r degree of

Real and M

NICAL REPO

laboratory

i informatiata and ap

tation feed

&T 2011 ca

13.7 is a g

ll mill work i

ariability mi

deled BM-

RT

grind time v

n was genlied to com

P80 of 150

paign, HD-

aphical repr

dex numbe

ht be expec

i Data

s. P80 particl

rated. Anosites 8-15,

µm, to esti

40 and HD-

esentation o

rs are not si

ted on a res

e size (150

equation w using the la

imate the B

9, BW-Wi r

f the BM-Wi

nificantly di

ource wide

µm) for com

s generateboratory gri

M-Wi numb

nges from

variability b

ferent betw

asis.

13-16

posites

(Seed time

ers for

low of

etween

en the

7/24/2019 2015 Halilaga


HALILAĞA

Effective Date

.

In Figure 1

surface.

Figure 13.

Source: Sim

To date n

Test), Bo

undertake

metallurgic

PRO J ECT

: December 20,

3.8 the dat

: Summary

mons 2014

Drop Wei

d Crushing

. Addition

al developm

PEA TECH

2014

shows a

Depth Down

ht Index Te

Index, Bo

l deposit w

ent studies.

NICAL REPO

odest decre

hole (meter

sting (either

d Rod Mil

ide commin

RT

ase in BM-

s) vs. BM-W

the JK Dro

l Index or

ution sampl

i for the co

i (kWh/t)

p Weight T

Bond Abra

es need to

mposites th

st or SAG

ion Index

be acquire

t are close

edia Com

esting hav

to support

13-17

to the

etency

been

future

7/24/2019 2015 Halilaga




.

Flotation Test Work Results

Flotation test work included open circuit cleaner flotation testing and culminated in two locked cycle

tests of the “Typical Ore” composite to produce a copper concentrate. The test work was used to assist

in developing the plant process design data and provide preliminary estimates of concentrate grade,

copper and gold recovery.

The flotation test work program was divided into two phases:

Open circuit cleaner flotation; and

Locked cycle cleaner flotation.

Open Circuit Cleaner Flotation

Cleaner flotation was conducted at a target primary grind size of 80% passing 150 μm and incorporated

a rougher re-grind stage to a target grind size of 80% passing 20 μm to increase mineral liberation.

Multiple stages of cleaning were incorporated to target high concentrate grades, typically with an

elevated pH level in the cleaning. The results from selected optimization tests are summarized in Figure

13.8.

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.

Table 13.8: Cleaner Flotation results for Halilağa Composite Samples

Comp ID

Re-GrindSizeµm

MassRec%



Comp 1 21 0.7 0.3 0.18 33.2 11.4 33.8 80.4 44.3 13.5

Comp 2 21 0.6 0.21 0.2 33.4 15.8 32.2 76.6 38.5 13.8

Comp 3 28 0.8 0.27 0.13 27.7 10.7 32 79.5 62.5 15.5

Comp 3 17 0.8 0.29 0.15 32 13.4 30.9 83.5 68 14.3

Comp 4 23 0.7 0.27 0.24 33.4 17.6 33.4 80.8 47.8 13.9

Comp 5 24 0.8 0.26 0.13 32 11.3 32.2 80.6 58 15.1

Comp 6 40 2.2 0.16 0.06 9.9 2.94 12 69.2 52.5 4.8

Comp 7 29 1 0.18 0.06 21.9 5.75 22.3 69.3 50.6 3.5

Comp 8 17 1.6 0.27 0.19 32.9 18.9 33.2 75.5 64.2 14.9

Comp 9 24 0.9 0.36 0.33 30 19.6 36.7 77.2 55.1 21.9

Comp 10 19 1.4 0.47 0.39 30.9 19.9 34 78 60.8 25.9

Comp 11 18 1.2 0.4 0.3 32.5 22.6 32.7 80.2 72.9 31.5

Comp 12 18 1.1 0.44 0.47 33.4 22.2 33.6 79.9 50.2 28.4

Comp 13 20 0.9 0.35 0.2 32.8 16.2 34.3 74.6 65.8 19.3

Comp 14 20 0.7 0.21 0.12 27.7 13.5 36.6 76.6 64.5 14.8

Comp 15 24 0.8 0.2 0.13 26.3 13.2 35.9 75.1 59.4 13


Copper recovery ranged from 74.6% to 83.5% (excluding composites 6 and 7) with copper concentrate

grade ranging from 26.3% Cu to 33.4% Cu. The low values for composites 6 and 7 are mainly

attributed to the coarser re-grind of 29 µm and 40 µm. Figure 13.9 shows a plot of copper and gold

recovery for the composite samples.

7/24/2019 2015 Halilaga


HALILAĞA

Effective Date

.

Figure 13.

Source: Sim

Figure 13.

rougher co

PRO J ECT

: December 20,

: Cleaner C

mons 2014

10 shows r

ncentrate to

PEA TECH

2014

opper and G

-grind P80

a P80 = 20µ

NICAL REPO

old Recover

s. Cu conc

before cle

RT

y for Comp

entrate gra

ner flotatio

site Sampl

e. There i

.

s

is a clear n

eed for re-g

13-20

rinding

7/24/2019 2015 Halilaga


HALILAĞA

Effective Date

.

Figure 13.

Source: Sim

Locked C

Following

conducted

locked cyc

types.

Flotation r

previously.

A flowshe

open circu

tests were

Locked cy

cleaner tai

these stre

can be ma

PRO J ECT

: December 20,

0: Re-grind

mons 2014

cle Test W

the comple

on the “Ty

le test to d

sults, using

t incorporati

it cleaner te

performed o

le tests ar

lings, on th

ms and co

e of the ov

PEA TECH

2014

P80 vs. Cu C

rk

ion of ope

ical Ore” co

termine clo

the optimiz

ng rougher c

ting was sel

n the “Typic

used to de

e overall gr

bining the

rall perform

NICAL REPO

oncentrate

circuit cle

mposite. Thi

ed circuit g

d test cond

oncentrate r

ected. To te

l Ore” com

termine the

de-recover

with conce

nce of a full

RT

rade

ner flotatio

s phase tes

rade-recove

itions, highli

e-grinding a

t the propo

osite sampl

effects of r

performan

trates from

scale plant

test work,

ted the opti

ry performa

ht the ben

nd multiple

ed flowshe

.

cycling inte

e of the m

a subseque

operation.

two locke

ized flows

ce of the

fit of fine re

tages of cle

t performan

rmediate str

ineralization

nt flotation t

cycle test

eet conditio

ineralized

grinding, as

ning, simila

e two locke

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type. By r

st, an asse

13-21

were

ns in a

aterial

stated

r to the

d cycle

as the

taining

ssment

7/24/2019 2015 Halilaga




.

Two locked cycle tests were undertaken for the “Typical Ore” composite, the results are summarized in

Table 13.9.

Table 13.9: Locked Cycle Test Results on “Typical Ore” Composite

Test No. Product

Re-GrindSizeµm

MassRecovery

Concentrate Grade Recovery

% % Cu g/t Au % S % Cu % Au % S

KM2908-17

FlotationFeed

20

100 0.3 0.3 1.64 100 100 100

Concentrate 0.9 25.4 15.5 40.1 80.7 49.2 23.2

1st CleanTails

8.8 0.41 0.84 13.3 12 25 72

RougherTails

90.2 0.02 0.08 0.09 7.3 25.8 4.9

KM2908-18

FlotationFeed

16

100 0.31 0.3 1.61 100 100 100

Concentrate 0.9 29.6 20.6 31.7 85.4 60.81 17.6

1st CleanTails

9.3 0.27 0.46 13.4 8.1 14 77.6

RougherTails

89.8 0.02 0.09 0.09 6.5 25.2 4.8


Analysis of these results indicates that recoveries of target minerals are acceptable and generally in line

with those achieved in the open circuit cleaner testing.

Flotation Test Work Results (ALS 2013)

The fourth test program, conducted in 2013, was based on a single shipment of samples taken from

three drill holes, HD-109, HD-115 and HD-124. The shipment consisted of 523 kg of quarter cut drill

core. The drill core was prepped and composited into three metallurgical samples for testing. Figure

13.11 shows the location of these drill holes.

The composites were taken from drill intervals, at various elevations, to represent mining progression

through the projected life of the resource. Composites were designated as:

Year 1 Composite;

Year 2-5 Composite; and

Year 6-14 Composite

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.

Head analysis of the three composites is presented in Table 13.10.

The Year 1 Composite targeted the supergene copper and gold zone at Halilağa which was known to

contain higher grade copper and gold values. Significant CuOX and CuCN copper values were

assayed in the Year 1 Composite head sample. Since this core was stored outside for approximately

two years (under tarps) before sampling – how much of the copper mineral oxidation is due to natural

(in-ground) oxidation vs. oxidation in the core boxes is unknown.

The head assays for the Year 2-5 and Year 6-4 Composites, taken from the hypogene zone of

mineralization, indicate minimal oxidation in the core boxes and subsequent flotation testing confirm this

assessment, as the test results do not indicate any sign of detrimental metallurgical performance.

Table 13.10: Head Analysis

CompositeAssay – percent or g/tonne

Cu CuOX CuCN Au Ag S Fe

Year 1 0.85 0.19 0.47 0.45 1.1 3.39 5

Year 2-5 0.34 0.007 0.016 0.37 1 2.92 3.29

Year 6-14 0.24 0.006 0.014 0.26 0.9 2.79 6.05


The principle objectives of this laboratory test program were to:

Perform flotation testing culminating in locked cycle flotation tests generating a cleaner tailings productfor subsequent cyanide leach testing; and

Conduct cyanidation bottle roll tests on the cleaner tailings products from the locked cycle tests andanalyze the cyanide liquors for dissolved metal content and cyanide speciation.

This program utilized a flowsheet developed in past metallurgical programs at ALS Metallurgy,

previously known as G&T Metallurgy who conducted the previous metallurgical test campaigns and was

utilized for this program. Three composites were tested from areas designated for mining representing

three periods of concentrator feed; the Year 1, the Years 2-5, and the Years 6-14.

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.

Figure 13.11: Location of Drill Holes HD-109, HD-115 and HD-124

Locked Cycle Flotation Testing

Batch rougher and cleaner tests were conducted using the same flowsheet developed in pastmetallurgical programs at G&T Metallurgy for the locked cycle tests. A small amount of Aero Promoter404 was used in the Year 1 composite flotation testing to improve flotation of tarnished sulphides in thissample. A summary of locked cycle test results are shown in Table 13.11.

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.

Table 13.11: Summary Locked Cycle Flotation Testing

ProductWeight

%

Assay – percent or g/tonne Distribution - percent

Cu Fe Ag Au Cu Fe Ag Au

Year 1 Comp – 10 Cycles

Flotation Feed 100 0.85 4.54 1 0.41 100 100 100 100

Cu Concentrate 1.9 32.3 20.7 31 11 71.2 8.5 61.5 49.8

Cu 1st Cleaner Tail 8.6 0.7 8.8 1.1 0.59 7.1 16.6 10.3 12.3

Cu Rougher Tail 89.6 0.2 3.8 0.3 0.17 21.7 74.9 28.1 37.9

Year 2-5 Comp – 8 Cycles

Flotation Feed 100 0.34 3.7 0.9 0.32 100 100 100 100

Cu Concentrate 1 29.1 27.8 50 16.4 89 7.9 59.2 52.9Cu 1

st Cleaner Tail 11.3 0.08 18.5 1.2 0.63 2.8 57.1 15.6 22.4

Cu Rougher Tail 87.7 0.03 1.5 0.3 0.09 8.2 35.1 25.2 24.7

Year 6-14 Comp – 9 Cycles

Flotation Feed 100 0.22 5.77 0.8 0.22 100 100 100 100

Cu Concentrate 0.7 29.7 28.2 75 19 90.6 3.3 63.3 58.7

Cu 1st Cleaner Tail 9 0.1 23.2 1.7 0.36 4.2 36.1 19.7 14.7

Cu Rougher Tail 90.4 0.01 3.88 0.1 0.06 5.1 60.6 17 26.6


Concentrate grade produced from these tests were high, averaging about 31% Cu and about 11 to

19 g/t Au. However, recoveries varied, ranging between 71% and 91% for Cu and 50% to 59% for Au. A

substantial portion of the gold reported to the rougher tail, about 25% to 38%. Approximately 12% to

22% of the gold reported to the first cleaner tailing.

The Year 1 Composite performance was below expectation, based upon previous locked cycle test

results. The Year 1 composite is a mix of supergene/hypogene material containing chalcocite, covellite

and chalcopyrite mineralization. The poor flotation performance is linked to the considerably high acid

and cyanide soluble Cu mineralization in the composite (See Table 13.10). Since this sample was

stored outside (under tarps) for two years, prior to sampling for metallurgical testing, it is unknown how

much of the Cu mineral oxidation is due to natural (in-ground) oxidation or oxidation in the core boxes.

It was recommended that fresh samples be drilled for additional testing of the mixedsupergene/hypogene zone as the project progresses forward.

Year 2-5 and Year 6-14 (hypogene chalcopyrite mineralization) composite samples performed as well

or better than expected. Low mass pull and high grade Cu concentrate was the norm for these two

composites. The locked cycle test results for these two composite samples do not appear to be

affected by the extended outside storage conditions.

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.

Cleaner Tails Cyanide Leaching

Cyanide bottle roll tests were conducted on the first cleaner tailings products generated in the locked

cycle tests. The tests were conducted at two sodium cyanide (NaCN) concentrations, 500 and 1,000 g/t.

The sodium cyanide concentration was maintained from 0 to 12 hours during the leach and allowed to

degrade following this duration. The total leach time was 24 hours. The results of the leach tests are

summarized in Table 13.12.

Table 13.12: Summary Cleaner Tails Cyanide Leach (24 Hrs)

Comp IDNaCNg/tonne

Gold Silver OverallRecovery**(%)Au (g/tonne) Recovery (%) Au (g/tonne) Recovery (%)

LeachFd

LeachTl

FromCT*

FromFd

LeachFd

LeachTl

FromCT*

FromFd

Au Ag

Year 1Year 1

5001000

0.550.48

0.320.28

41.843.1

5.15.3

1.21.2

1.21.2

2.12.0

0.20.2

54.955.1

61.761.7

Year 2-5Year 2-5

5001000

0.600.49

0.140.13

76.573.4

17.116.4

1.61.6

0.60.5

62.368.0

9.710.6

70.069.3

68.969.8

Year 6-14Year 6-14

5001000

0.290.30

0.060.05

79.184.8

11.612.5

1.81.8

0.50.5

71.775.4

14.114.9

70.371.2

77.478.2

Source: Simmons 2014Fd – Feed, Tl – Tail, CT – Cleaner tails

The cyanide leaching on the Year 1 Composite, show lower extraction and higher cyanide

consumptions than the other composites. This was expected as the Year 1 Composite head contained

higher cyanide soluble copper, which consumes cyanide in a leach scenario. Gold extraction from the

cleaner tails was about 43% representing approximately 5% of the gold in the flotation feed. This

brought overall gold extraction from the flotation feed to about 55%.

Cyanide leach tests were more successful on the other composites. Gold extracted from the Year 2-5

and Year 6-14 Composites, first cleaner tailing, ranged between 73% and 85%. On a plant feed basis,

the Year 2-5 Composite gold recovery represented approximately 17% of the flotation feed gold and

about 12% of the flotation feed gold for the Year 6-14 Composite.

Overall gold recovery (gold recovered in the flotation 3 rd cleaner concentrate plus the cleaner tails

cyanide leach) for the Year 2-6 Composite averaged about 70% and for the Year 6-14 Composite

averaged about 71%.

The first cleaner tailings silver grade was very low averaging about 1.5 g/t for all three composites.

Although silver extraction was relatively high, ranging from 62-78%, the resource grade is too low to

represent any significant value to the project.

Cyanide and lime consumptions were measured for all tests and the results are shown in Table 13.13.

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Table 13.13: Cyanide and Lime Consumption

Comp ID

Cleaner Tail – kg/tonne Plant Feed – kg/tonne

Cyanide Lime Cyanide Lime

500 ppm 1000 ppm 500 ppm 1000 ppm 500 ppm 1000 ppm 500 ppm 1000 ppm

Year 1 2.8 5.7 4.5 4.8 0.2 0.5 0.4 0.4

Year 2-5 1.6 1.9 5.6 4.5 0.2 0.2 0.6 0.5

Year 6-14 1.7 3.3 11.5 9.3 0.2 0.3 1 0.8


Similar gold extraction results were achieved at both concentrations of sodium cyanide, in most cases

slightly higher metal extractions were recorded at the higher concentration. Higher cyanideconsumptions were also recorded at the higher concentration. Given the low leach feed gold grade, the

differences in tail grades were relatively low.

Cleaner Tails Cyanide Leach Kinetics

Leach kinetic samples were taken at 4, 8, 12 and 24-hours to measure incremental gold extraction and

cyanide consumption. Total gold leach extraction and kinetic gold leach extraction were only mildly

impacted by increasing the NaCN concentration. The tests were conducted at two sodium cyanide

(NaCN) concentrations, 500 and 1,000 g/t. The sodium cyanide concentration was maintained from 0

to 12 hours during the leach and allowed to degrade following this duration. The total leach time was

24-hours. The results of the leach tests are summarized in Table 13.12

Kinetic leach data analysis indicates very little benefit to leaching beyond 12-hours. See Figure 13.12.

In fact, gold extraction of composite Year 2-5 and Year 6-4 marginally declined after 12- hours of

leaching where gold extraction from the Year 1 composite marginally increased.

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HALILAĞA

Effective Date

.

Figure 13.

Source: Sim

Cyanide c

6-14 com

13.13.

PRO J ECT

: December 20,

2: Au Extra

mons 2014

nsumption i

osites. Th

PEA TECH

2014

ction vs. Le

ncreased by

Year 1 co

NICAL REPO

ch Time

over 50% i

posite cya

RT

the last 12-

ide consu

hours of lea

ption was fl

hing for the

at after 12

Year 2-5 a

ours. See

13-28

d Year

Figure

7/24/2019 2015 Halilaga


HALILAĞA

Effective Date

.

Figure 13.

Source: Sim

Due to the

cleaner tail

PRO J ECT

: December 20,

3: Cyanide

mons 2014

levelling off

ls leach circ

PEA TECH

2014

Consumptio

of gold extra

it was analy

NICAL REPO

n vs. Leach

ction results

zed using th

RT

Time

and increas

e 12 hour ki

ing cyanide

etic leach d

onsumptio

ata. See Ta

after 12 ho

ble 13.14.

13-29

rs, the

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.

Table 13.14: Twelve Hour Cleaner Tails Leach Analysis

ALS Lab Results (12 Hrs Cleaner Tails Leach)

Item500 ppm NaCN

Yr 1 Yr 2-5 Yr 6-14 Yr2-14

Composite Feed Grade

Cu (%) 0.85 0.34 0.22 0.28

Au (g/t) 0.41 0.32 0.22 0.27

Ag (g/t) 1.0 0.90 0.80 0.85

Flotation Concentrate Grade

Cu (%) 32.3 29.1 29.7 29.4

Au (g/t) 11.0 16.4 19.0 17.7

Ag (g/t) 31.0 50.0 75.0 62.5

Flotation Concentrate Recovery (%)

Cu Recovery 71.2 89.0 90.6 89.8

Au Recovery 49.8 52.9 58.7 55.8

Ag Recovery 61.5 59.2 63.3 61.3

Cleaner Tails Leach (CTL)

CT Mass - % of Flot Feed 8.6 11.3 9.0 10.2

CTL Au Grade (g/t) 0.17 0.63 0.36 0.50

Total Au Reporting to CTL Feed 12.3 22.4 14.7 18.6

Au Rec (% of CTL Feed) 41.5 80.9 81.3 81.1

Au Rec (% of Total) 5.1 18.1 12.0 15.0

Overall Flotation + CTL Recovery (%)

Au Rec (% of Total) 54.9 71.0 70.7 70.8

Reagent Consumption

NaCN (k/t of Plant Feed) 0.240 0.220 0.167 0.194

Lime (kg/t of Plant Feed) 0.387 0.633 1.035 0.834


For the hypogene portion of the resource, combined Years 2-14, the cleaner tails leach achieved anaverage gold extraction of 81.1% which is equal to 15.0% of total gold contained in the flotation feed

samples.

Corresponding cyanide consumption was 1.94 kg/t of cleaner tails leach feed and 0.194 kg/t of flotation

plant feed.

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13.2 Grade and Recovery Predictions

Analysis of the open circuit cleaner and locked cycle tests have been undertaken to estimate flotation

performance predictions. Three locked cycle tests were selected as the basis for the design recoveries

of the target metals copper and gold, the results are summarized in Table 13.15.

Table 13.15: Summary Locked Cycle Flotation Data

Comp ID ZoneHead Grade Re-grind Weight Clr Conc Assay Recovery (%)

Cu (%) Au (g/t) P80 µm % Cu (%) Au (g/t) Cu Au

Yr 1 Secondary/Transition 0.85 0.45 24 1.9 32.3 11.0 71.2 49.8

Yr 2-5 Primary Sulphide 0.34 0.37 26 1.0 29.1 16.4 89.0 52.9

Yr 6-14 Primary Sulphide 0.24 0.36 26 0.64 29.7 19.0 90.6 58.7

Typ Comp Secondary/Primary 0.31 0.30 16 0.9 29.6 20.6 85.4 60.8

Source:

Copper Recovery and Grade Models

The Year 1 composite was left out of the modeling due to its poor performance and knowing that this

secondary/transition composite sample was stored outside for two years before being sampled for

testing, making the results suspect.

Copper recovery versus copper feed grade was plotted for the selected cleaner tests in Figure 13.14,

along with the three locked cycle test results and the predicted copper recovery curve. An equation for

copper recovery was developed vs. copper feed grade. The equation established for the open circuitcleaner tests was adjusted to account for the improvement seen with recycling streams during locked

cycle testing. As a result of recycling streams copper recovery increased by approximately 7%

compared to the equivalent batch cleaner test. The predicted copper recovery vs. copper feed grade is

shown in the following equation:

Copper Recovery (%) = 96.452(Copper Feed Grade)0.0889

Similar methodology was use for correcting Cu concentrate grade. As expected the locked cycle

flotation testing improved Cu recovery at a slightly reduced Cu grade. See Figure 13.15. The predicted

concentrate copper grade vs. copper head grade is shown in the following equation:

Concentrate Copper Grade = 34.902(Copper Feed Grade)0.1272

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HALILAĞA

Effective Date

.

Figure 13.

Source: Sim

PRO J ECT

: December 20,

4: Copper

mons 2014

PEA TECH

2014

eed Grade

NICAL REPO

s. Copper R

RT

ecovery

13-32

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HALILAĞA

Effective Date

.

Figure 13.

Source: Sim

Gold Rec

Gold reco

the locked

between t

presented

PRO J ECT

: December 20,

5: Copper

mons 2014

very and G

ery was mo

cycle flota

e open circ

in Figure 13.

PEA TECH

2014

oncentrate

rade Model

deled in a s

ion test res

uit cleaner

16.

NICAL REPO

Cu Grade M

imilar fashio

ults. Unlik

nd locked

RT

del

n to copper

copper th

ycle flotatio

using the o

gold reco

n tests. Th

en circuit cl

ery was es

e model res

eaner flotati

sentially th

ults are gra

13-33

on and

same

hically

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HALILAĞA

Effective Date

.

Figure 13.

Source: Sim

Since theflotation t

recovery is

Au Recov

In the next

of the min

Copper co

Gold Grad

methods p

Cu/Au Ratgrade, via

vs. Cu/Au

Concentr

PRO J ECT

: December 20,

6: Gold Fee

mons 2014

locked cyclst results, t

modeled us

ry (%) = 1.

phase of th

ralized mat

ncentrate g

e in Conce

roduced acc

io predictivlocked cycl

Ratio is sho

te Au Grad

PEA TECH

2014

d Grade vs.

e tests arehe open cir

ing the follo

702 (Gold

project, m

rial to confir

ld grade wa

trate and 2

eptable gol

model. S testing, fro

n in the foll

= -6.0715 (

NICAL REPO

Gold Recov

almost a pcuit and m

ing equatio

eed Grade)

re locked c

m the desig

s graphicall

.) Feed Gra

grade pred

e Figure 1 what was

wing equati

Copper/Gol

RT

ry

rfect symmdeled gold

n:

+ 57.782

cled tests a

gold recov

modeled u

e Cu/Au R

ictive result

.17. The rshown in th

on:

d Ratio) + 2

etrical overlrecovery e

re required

ries.

ing two me

atio vs. Gol

. For this s

esults indice previous s

4.355

ay of the ouations ar

n samples f

hods: 1.) G

Grade in

tudy, it was

te a modestudy. The p

pen circuitthe same.

rom differen

ld Feed Gr

oncentrate

decided to

t reductionedicted gol

13-34

leanerGold

zones

de vs.

. Both

se the

in goldgrade

7/24/2019 2015 Halilaga


HALILAĞA

Effective Date

.

Figure 13.

Source: Sim

13.3 C

The coppresults are

The coppefrom lockecycle testi

PRO J ECT

: December 20,

7: Copper

mons 2014

oncentrate

r concentratpresented i

r concentratd cycle testig should in

PEA TECH

2014

oncentrate

Quality

e from lock Table 13.1

was free ong of the Ylude additio

NICAL REPO

Au Grade M

d cycle test.

elements tar 1, Yearal copper c

RT

del

18 was su

at might inc2-5 and Yencentrate i

jected to a

ur smelter pr 6-14 compurity analy

detailed ch

nalties. Noposite sampsis.

mical analy

data was ales. Future

13-35

is; the

ailablelocked

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.

Table 13.16: Quality of Concentrate

Element Units Locked Cycle Test 18 Concentrate

Aluminum % 0.27

Antimony % 0.04

Arsenic g/t 163

Bismuth g/t 26

Cadmium g/t 20

Calcium % 0.21

Carbon % 0.13

Cobalt g/t 72

Copper % 30.6

Fluorine g/t 51

Gold g/t 19.7

Iron % 29.2

Lead % 0.06

Magnesium % 0.08

Manganese % <0.001

Mercury g/t 3

Molybdenum % 0.17

Nickel g/t 92

Phosphorus g/t 50

Selenium g/t 80

Silicon % 1.13

Suphur % 32.1

Silver g/t 50.6

Zinc % 0.1


Any deleterious elements identified at this time exist at concentrations below threshold smelter penalty

levels.

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13.4 Further Metallurgical Test Work Required

Further work will be required for a PFS/FS in order to provide more confidence in the process design

criteria. It is recommended to conduct additional comminution and flotation variability test work across

the mineralized zones and use this additional data to develop grade/recovery models that take into

account variations in head grade and material type.

Recommended additional metallurgical testing includes the following:

Test additional variability samples that provide greater spatial representation of the mineable

resource and which represent the complete volume of material to be mined. In particular, the

secondary enrichment zone needs better characterization and variability composite testing.

Additional comminution test work, including JK drop weight, SMC, Bond Crusher work index,

Bond Ball Mill work index, Bond abrasion index testing;

Review of comminution circuit selection and design by an external consultant;

Conduct additional reagent and flotation optimization flow sheet testing to confirm process

design criteria and potentially improve metallurgical performance;

Test the potential for increasing gold recovery into the rougher concentrate. This will result in

improved overall combined gold recovery, from the combination of gold in the final copper

concentrate, plus additional gold recovered in the first cleaner tails leach. As significant gold

reports directly to the rougher tails, another recommended approach is to evaluate pyrite/gold

flotation, post copper rougher flotation, to determine if additional cost effective Au recovery can

be achieved from this product;

Additional flotation variability and lock cycle testing to improve metallurgical models and confirmregrind size;

Optimization of cyanide leach conditions on first cleaner tails products produced from locked

cycle or pilot plant tests;

Test work to confirm concentrate thickening rates and concentrate thickener sizing;

Test work to confirm tailings thickening rates for tailings thickener selection and sizing;

Test work to confirm concentrate filtration rates and filter selection;

Rheology test work to confirm tailings pumping, pipeline and distribution design;

Bulk materials handling test work to optimize design of the ROM bin, conveyors, chutes,

feeders, crushed plant feed material stockpile and reclaim facility;

Conduct a concentrate marketing study to identify potential Cu Smelter destinations andassociated treatment costs, refining charges and penalties;

Additional mineralogy on metallurgical variability composites, final concentrates and tails; and

Work towards developing a 3D Geometallurgy model for the resource.

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14 MINERAL RESOURCE ESTIMATES

14.1 Introduction

This report documents the update of the initial Halilağa mineral resource documented in an NI 43-101

Technical Report in March of 2012; this updated resource includes all drill results available since that

time. This resource was estimated by ordinary kriging, using Gemcom® software as opposed to the

geometric method of inverse distance weighting used for the initial resource. The geologic model used

for this resource was again prepared by Teck staff and is conceptually the same as that used for the

initial resource with the addition of an altered porphyry unit recognized in 2012 fieldwork. Geologic

control for estimation was based on rock type as well as structural zonation on the flanks of the

porphyry unit as it was for the 2012 resource. Copper, gold and molybdenum grades were estimated

using 2.0 m composited drill data.

The resource is tabulated within the same optimized pit shell as was generated and used for the 2012

reported resource. The impact of drilling since the initial resource has been to increase confidence as

reflected by the increase in Indicated Mineral Resource as a portion of the total resource. Table 14.1

compares the 2014 updated sulphide resource with the initially reported numbers; the 0.43 g/t AuEq

cut-off approximately corresponds to the 0.2% CuEq cut-off used in the 2012 disclosure.

Table 14.1: Comparison to Initial Estimate at 0.43 g/t AuEq Cut-off

ResourceModel

Indicated Inferred

Tonnes(Millions)

Cu(%)

Au(g/t)

Mo(%)

AuEq(g/t)

Tonnes(Millions)

Cu(%)

Au(g/t)

Mo(%)

AuEq(g/t)

Update 182.7 0.27 0.30 0.0057 0.90 178.7 0.23 0.24 0.0087 0.77

Initial 168.8 0.30 0.31 0.0054 0.97 199.6 0.23 0.26 0.0067 0.78

Difference +8% -10% -4% +6% -8% -10% -1% -7% +30% -2%

Source: Gray 2014

14.2 Available Data

This Halilağa resource update is based on assay data available as of July 4, 2013. Results from 115

holes have been used for this estimate, of these 112 were core holes and three were RC. Figure 14.1

shows drill hole locations as well as the limits of the resource model. The block model geometry is listed

in Table 14.2 and is unchanged from the initial model.

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Figure 14.1: Halilağa Drilling and Resource Model Outline (New Holes in Red)

Source: Gray 2014

N

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Table 14.2: Resource Block Model Setup

Block: X (Easting) Y (Northing) Z (Elevation)

Origin(1)

in UTM Coordinates 481,800 4,418,560 600

Size (m) 20 20 10

n blk (Number of Blocks) 135 72 105

Non-rotated: 1,020,600 blocks(1)

SW model top, block edge

Source: Gray 2014

14.3 Geologic Model

The geologic model was again prepared by site geology and Teck exploration staff. Conceptually the

model was essentially the same as the 2012 version; one additional unit was recognized for this

estimate. Five lithologic units were interpreted and constrained within three-dimensional solids honoring

drill hole contacts. These reflect the porphyry intrusion into felsic volcanic rock to the south, a barren

cover unit, which consisted of a mix of loose unmineralized andesitic/basaltic volcanics and sediments to

the north, and a hornfels unit that variably surrounds the porphyry. The northern side of the porphyry

body is interpreted to be in faulted contact with overlying cover material. For this estimate, the porphyry

was subdivided into an altered and an unaltered variety.

Contacts of the oxide and supergene volumes were based primarily on soluble copper assays and were

updated for this estimate. The top of supergene zone (bottom of oxide zone) was marked down-hole by

an abrupt increase in total and cyanide soluble copper grade. A pronounced reduction in cyanide

soluble to total copper assay was used to delineate the bottom of the supergene enriched zone.Twenty-two holes were used to estimate grade in the oxide and supergene zones.

Overburden was not modeled separately. Its thickness is generally low (< 5 m) and overburden is often

not differentiated from the cover unit in logs. Rock type assignment to individual blocks was on a

majority rules basis as this was determined to adequately reproduce solid volumes. An example section

is shown in Figure 14.2.

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Figure 14.2: Sectional Interpretation 483,550E (view to west)

Source: Gray 2014

14.4 Assay Compositing

Sample data was composited to a down-hole length of 2.0 m. The choice of composite length was

based primarily on its relation to the length of samples assayed. Sixty-eight percent of samples were

either 1 or 2 m in length (54% - 2 m, 14% - 1 m).

A total of 140 composites of less than half length (1.0 m) were removed from the dataset used for grade

estimation, after it was determined that this did not fundamentally affect the grade statistics by rock

type. A total of 21,502 composites were used for grade estimation.

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14.5 Grade Capping

Grade capping is used to control the impact of extreme, outlier high-grade samples on the overall

resource estimate. The levels selected as being outliers to the general population were determined by

examining histograms and probability plots by rock type and by metal. Copper, gold, and molybdenum

grades in the mineralized units generally display a log-normal distribution with very little scatter at the

highest portion of the distribution. For copper and gold, the majority of composites that required grade

capping were outside the main mineralized porphyry unit.

Capping levels applied to composite grades are listed in Table 14.3. Capped versus uncapped

composite statistics are presented in Table 14.4.

Table 14.3: Grade Capping Levels

Cap Levels

Rock TypeCu(%)

Au(g/t)

Mo(%)

8 Oxide 0.21 -- 0.016

9 Supergene 3.40 1.40 0.020

30 Hornfels 0.60 2.00 0.100

50 Felsic Volcanics 0.13 0.30 0.020

60 Porphyry 1.30 1.80 0.100

70 Altered Porphyry -- 2.00 0.040

80 Cover 0.01 0.08 0.002

Source: Gray 2014

The amount of metal removed through the top-cutting process was calculated by comparing capped

and uncapped nearest neighbor models. In total, metal removed by capping composite grades

amounted to 0.4% copper, 1.0% gold, and 1.3% molybdenum.

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Table 14.4: Composite Statistics

Cap Impact

Rock Type

Cu(%)

CuCap(%)

count mean max CV n cap'd mean max CV

8 Oxide 140 0.06 0.27 0.9 2 0.06 0.21 0.8

9 Supergene 259 1.28 5.97 0.6 1 1.27 3.40 0.5

30 Hornfels 6,739 0.07 1.01 1.5 11 0.07 0.60 1.5

50 Felsic Volcanics 1,835 0.01 0.39 2.6 9 0.01 0.13 2.4

60 Porphyry 7,037 0.15 3.04 0.8 1 0.15 1.30 0.8

70 Altered Porphyry 2,028 0.33 1.66 0.7 0 0.33 1.66 0.7

80 Cover 3,464 0.00 0.26 5.1 13 0.00 0.01 1.1Total 21,502 37

Rock TypeAu (g/t) AuCap (g/t)


8 Oxide 140 0.61 1.96 0.7 0 0.61 1.96 0.7

9 Supergene 259 0.63 4.06 0.6 6 0.61 1.40 0.5

30 Hornfels 6,739 0.07 3.67 1.8 3 0.07 2.00 1.8


60 Porphyry 7,037 0.17 2.79 0.9 3 0.17 1.80 0.9


80 Cover 3,464 0.00 0.35 4.2 13 0.00 0.08 3.0Total 21,502 40

Rock TypeMo (%) MoCap (%)


8 Oxide 140 0.004 0.030 1.1 4 0.004 0.016 0.9

9 Supergene 259 0.003 0.027 1.2 3 0.003 0.020 1.1

30 Hornfels 6,739 0.004 0.168 1.9 4 0.004 0.100 1.8


60 Porphyry 7,037 0.005 0.177 1.5 5 0.005 0.100 1.5


80 Cover 3,464 0.000 0.007 1.6 2 0.000 0.002 1.4

Total 21,502 27

Source: Gray 2014

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14.6 Grade Interpolation

Grades for this updated resource were estimated by ordinary kriging (OK) using Gemcom® software.

Grades were interpolated by rock type and within dipping zones flanking the main porphyry body.

The estimation search strategy was established based mainly on directions of grade continuity noted in

plan and section. Most notably copper and gold grades trend roughly parallel to the margins of the

porphyry unit. Plots of grades versus distance from the porphyry contact led to the creation of north-

dipping and south-dipping domains that extended 20 m into the low grade units to the north and south

(cover and felsic volcanics), and up to 100 m into the hornfels or into the core of the porphyry. This

process adequately constrained most of the anomalously high grades in the cover and felsic volcanic

units.

Spatial continuity of capped composite data was analyzed using Supervisor® software. Copper, goldand molybdenum data were subdivided by rock type and within north and south-dipping domains to

establish suitable variogram model parameters for use in estimation. Variogram models used are listed

in Tables 14.5 to Table 14.7 for copper, gold, and molybdenum respectively.

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Table 14.5: Variogram Models - Copper

DomainRotation Direction

(dip/azimuth)NuggetEffect

Spherical Component 1 Spherical Component 2

(axis) (RHR) Sill Range (m) Sill Range (m)

8Oxide

Z 0 00/000

0.07 0.20

12

0.73

25

X 0 00/270 12 25

Z 90 90/000 12 25

9Supergene

Z 145 00/305

0.17 0.59

25

0.24

50

X -85 -85/215 35 100

Z 0 05/215 55 110

31HornfelsN.Dipping

Z 0 00/090

0.07 0.23

15

0.70

145

X -50 -50/000 15 300

Z 0 40/000 20 100

32

HornfelsS.Dipping

Z -10 00/100

0.06 0.22

15

0.72

200

X 60 -60/190 20 350Z 0 -30/010 10 100

33Hornfels

Z 5 00/085

0.06 0.22

35

0.72

50

X 20 -20/175 20 65

Z 0 -70/355 10 70

50FelsicVolcanics

Z 0 00/000

0.07 0.08

50

0.85

150

X 0 00/270 50 150

Z 90 90/000 50 150

61PorphyryN.Dipping

Z 20 -65/340

0.15 0.50

25

0.35

440

X -65 00/250 115 325

Z 90 25/340 20 30

62Porphyry

S.Dipping

Z 0 00/090

0.13 0.27

35

0.60

95

X 60 -60/180 30 100

Z 0 -30/000 30 90

63Porphyry

Z 35 -70/145

0.10 0.25

35

0.65

175

X 70 00/235 15 120

Z -90 -20/325 10 135

71Alt'dPorphyryN.Dipping

Z 30 00/060

0.10 0.23

30

0.67

150

X -60 -60/330 15 135

Z 0 30/330 4 25

72Alt'dPorphyryS.Dipping

Z -35 00/125

0.15 0.20

25

0.65

130

X 75 -75/215 30 65

Z 0 -15/035 25 65

73AlteredPorphyry

Z -40 00/130

0.10 0.15

10

0.75

100

X -65 -65/040 20 100

Z 0 25/040 10 55

80BarrenCover

Z 0 00/000

0.03 0.06

15

0.91

100

X 0 00/270 15 100

Z 90 90/000 15 100

Source: Gray 2014

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Table 14.6: Variogram Models - Gold





8 / 9Oxide/Supergene

Z 145 00/305

0.24 0.24

10

0.52

100

X -65 -65/215 5 35

Z 0 25/215 10 60

31HornfelsN.Dipping

Z 0 00/090

0.23 0.15

25

0.62

120

X -50 -50/000 35 195

Z 0 40/000 20 75

32HornfelsS.Dipping

Z -10 00/100

0.27 0.19

25

0.54

120

X 60 -60/190 35 180

Z 0 -30/010 15 60

33Hornfels

Z 5 00/085

0.22 0.38

35

0.40

50

X 20 -20/175 20 65Z 0 -70/355 35 230

50FelsicVolcanics

Z 0 00/000

0.09 0.14

60

0.77

100

X 0 00/270 60 100

Z 90 90/000 60 100

61PorphyryN.Dipping

Z 20 -65/340

0.13 0.28

30

0.59

100

X -65 00/250 15 150

Z 90 25/340 10 80

62PorphyryS.Dipping

Z 0 00/090

0.20 0.18

30

0.62

95

X 60 -60/180 40 70

Z 0 -30/000 15 75

63

Porphyry

Z 35 -70/145

0.15 0.19

25

0.66

210

X 70 00/235 25 210

Z -90 -20/325 10 15071Alt'dPorphyryN.Dipping

Z 30 00/060

0.17 0.36

55

0.47

95

X -60 -60/330 80 110

Z 0 30/330 30 40

72Alt'dPorphyryS.Dipping

Z -35 00/125

0.18 0.34

25

0.48

130

X 75 -75/215 30 65

Z 0 -15/035 25 65

73AlteredPorphyry

Z -40 00/130

0.13 0.15

20

0.72

100

X -65 -65/040 35 100

Z 0 25/040 15 55

80BarrenCover

Z 0 00/000

0.35 0.13

10

0.52

20

X 0 00/270 10 20

Z 90 90/000 10 20

Source: Gray 2014

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.

Table 14.7: Variogram Models - Molybdenum





8 / 9Oxide/Supergene

Z 50 00/040

0.49 0.28

35

0.23

120

X -30 -30/310 20 60

Z 0 60/310 8 33

31HornfelsN.Dipping

Z 0 00/090

0.36 0.08

75

0.56

100

X -50 -50/000 5 50

Z 0 40/000 85 180

32HornfelsS.Dipping

Z 0 00/090

0.22 0.27

75

0.51

170

X 80 -80/180 25 150

Z 0 -10/000 25 65

33Hornfels

Z 5 00/085

0.23 0.22

20

0.55

125

X 15 -15/175 20 85Z 0 -75/355 20 105

50FelsicVolcanics

Z 0 00/000

0.30 0.13

30

0.57

100

X 0 00/270 30 100

Z 90 90/000 30 100

61 / 71PorphyryN.Dipping

Z -140 07/031

0.30 0.36

20

0.34

95

X -20 19/299 8 120

Z -160 70/140 8 25

62 / 72PorphyryS.Dipping

Z 40 00/050

0.23 0.39

10

0.38

120

X 75 -75/140 5 15

Z 0 -15/320 20 30

63 / 73Porphyry &

Alt'd Porph

Z -40 -63/110

0.25 0.48

50

0.27

150

X -80 -25/315 35 80

Z 65 10/040 30 55

80BarrenCover

Z 0 00/000

0.08 0.09

10

0.83

65

X 0 00/270 10 65

Z 90 90/000 10 65

Source: Gray 2014

Search distances in the dipping zones were established based on a visual assessment of sections as

well as through consideration of variogram model ranges.

Grade relationships among rock types were assessed using boundary/grade plots across all contacts

for each metal; grades were generally not interpolated across rock type contacts. The exception was for

molybdenum grades across the altered/unaltered porphyry contact. Plots of grade versus distance to

that contact confirmed the use of a soft contact for Mo grade interpolation.

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.

Blocks outside the dipping corridors were interpolated with a 150 m spherical search except in the core

of the porphyry where an oblate E-W, vertically-oriented search was deemed most appropriate. Blocksin the upper weathered zone were interpolated in a single pass with a 75 m spherical search. Generally

a two pass search approach was used; units for which an anisotropic or 75 m search was used in

Pass 1 had a second 150 m spherical search applied in Pass 2 to adequately fill blocks. Search details

are listed by rock type in Table 14.8.

Table 14.8: Interpolation Parameters

Rock Type orDirectional Domain

Search Pass 1Search (m)

X/Y/Z

Search Pass 2Search (m)

X/Y/Z

8 Oxide 100/100/100

9 Supergene 100/100/100

30 Hornfels 150/150/150

north dipping 75/75/75 150/150/150

south dipping 75/75/75 150/150/150

50 Felsic Volcanics 150/150/150

south dipping 75/75/75 150/150/150

60/70 Porphyry (unalt. & alt.) 150/65/100 150/150/150

north dipping 75/75/75 150/150/150

south dipping 75/75/75 150/150/150

80 Cover 150/150/150

north dipping 75/75/75 150/150/150

Source: Gray 2014

Copper, gold, and molybdenum grades were estimated according to the above search criteria using a

minimum of three samples, a maximum of 16 samples, and a maximum of five samples per hole. This

approach resulted in approximately 80% of classified blocks being estimated by two or more holes.

14.7 Density Assignment

In total, 8,025 dry bulk density measurements were used to calculate an average density per rock type.

These samples were coded for rock type based on drill hole intersections with the geologic solids.

Probability plots of density by rock type were examined for any obviously spurious values; six were

detected and removed from the original dataset.

Table 14.9 lists details of density measurements and the mean values assigned to the seven rock type

domains.

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Table 14.9: Average Rock Type Density

Rock TypeDensity (t/m3)

Count Mean Min Max

8 Oxide 87 2.30 1.93 2.78

9 Supergene 154 2.43 1.80 2.81

30 Hornfels 2,723 2.65 1.83 3.98

50 Felsic Volcanics 504 2.57 1.92 2.97

60 Porphyry 3,056 2.60 1.97 3.16

70 Altered Porphyry 1,134 2.56 1.90 3.01

80 Cover 367 2.41 1.76 2.78

Total 8,025 2.59 1.76 3.98

Source: Gray 2014

14.8 Model Validation

A nearest neighbor (NN) model was estimated using the same rock type matching as in the inverse

distance estimate, but without the influence of the north and south-dipping structural corridors; this

model was estimated using an isotropic 150 m search. The NN model used a 2 m block height

reflecting the composite length. The NN model was re-blocked to the resource model grid (5:1) and

used to check various aspects of the estimation process.

A second check model was estimated by inverse distance squared (ID) weighting. This matches the

approach that was used for the 2012 reported resource.

Visual assessment, comparing sample composite values and block grades on plans and sections,

showed good correlation for copper, gold and moly grades. Copper and gold equivalent grade variables

were calculated based on block estimated (OK) grades using parameters listed in Table 14.10.

Table 14.10: Metal Equivalence Parameters

Metal Metal Price Recovery

Cu US$ 2.90 / lb 90%

Au US$ 1200 / oz 70%

Mo US$ 12.5 / lb 50%

Source: Gray 2014

A more quantitative validation was made by generating swath plots by block model northings, eastings,and elevations to spatially compare the resource model against NN and ID results. Plots weregenerated by resource class, globally and within the 2012 optimized pit shell. Plots of all Indicatedblocks are presented in Figure 14.3 for copper grade and in Figure 14.4 for gold grade. All plots showreasonable spatial correlation between estimated blocks and the underlying composite data.

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Figure 14.3: Copper Grade Swath Plots Comparing OK, ID and NN Estimates

Source: Gray 2014

0

10,000

20,000

30,000

40,000

50,000

60,000

0.00

0.05

0.10

0.15

0.20

0.25

0.30

482,200 482,400 482,600 482,800 483,000 483,200 483,400 483,600 483,800 484,000 484,200 484,400

T o n n e s ( 1 , 0

0 0 s )

C u ( % )

East

Block Model Easting: 43,800 Indicated Blocks

OK; mean: 0.13

ID2; mean: 0.13

NN; mean: 0.12

Tonnes

0

10,000

20,000

30,000

40,000

50,000

0.00

0.05

0.10

0.15

0.20

0.25

4,418,800 4,419,000 4,419,200 4,419,400 4,419,600 4,419,800 4,420,000

T o n n e s ( 1 , 0

0 0 s )

C u ( % )

North

Block Model Northing: 43,800 Indicated Blocks

OK; mean: 0.13

ID2; mean: 0.13

NN; mean: 0.12

Tonnes

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.

Source: Gray 2014

0

10,000

20,000

30,000

40,000

50,000

0.00

0.05

0.10

0.15

0.20

0.25

-300 -200 -100 0 100 200 300 400 500

T o n

n e s ( 1 , 0

0 0 s )

C u ( % )

Elevation

Block Model Benches: 43,800 Indicated Blocks

OK; mean: 0.13

ID2; mean: 0.13

NN; mean: 0.12

Tonnes

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.

Figure 14.4: Gold Grade Swath Plots Comparing OK, ID and NN Estimates

0

10,000

20,000

30,000

40,000

50,000

60,000

0.00

0.05

0.10

0.15

0.20

0.25

0.30

482,200 482,400 482,600 482,800 483,000 483,200 483,400 483,600 483,800 484,000 484,200 484,400

T o n n e s ( 1 , 0

0 0 s )

C u ( % )

East

Block Model Easting: 43,800 Indicated Blocks

OK; mean: 0.13

ID2; mean: 0.13

NN; mean: 0.12

Tonnes

0

10,000

20,000

30,000

40,000

50,000

0.00

0.05

0.10

0.15

0.20

0.25

4,418,800 4,419,000 4,419,200 4,419,400 4,419,600 4,419,800 4,420,000

T o n n e s ( 1 , 0

0 0 s )

C u ( % )

North

Block Model Northing: 43,800 Indicated Blocks

OK; mean: 0.13

ID2; mean: 0.13

NN; mean: 0.12

Tonnes

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.

Source: Gray 2014

14.9 Resource Classification and Tabulation

The resource estimate was classified based on spatial parameters related to drill density and

configuration, and inclusion within the previously optimized pit shell. The classification criteria applied to

the Halilağa resource are listed in Table 14.11. These parameters are: (i) the minimum number of holes

used to estimate grade, (ii) the average distance to samples used to estimate grade, (iii) the distance to

the closest sample used in the estimate, (iv) the distance to the second closest sample, and (v) the

distance of the third closest sample.

Classification criteria were established iteratively by visually assessing the impact of parameter

adjustment on resultant maps of classified blocks. The goal was to have reasonably cohesive volumes

rather than a scattered patchwork of Indicated and Inferred blocks, while assigning the Indicated

category in a justified pattern based on sampled locations.

0

10,000

20,000

30,000

40,000

50,000

0.00

0.05

0.10

0.15

0.20

0.25

-300 -200 -100 0 100 200 300 400 500

T o n n e s ( 1 , 0

0 0 s )

C u ( % )

Elevation

Block Model Benches: 43,800 Indicated Blocks

OK; mean: 0.13

ID2; mean: 0.13

NN; mean: 0.12

Tonnes

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.

Table 14.11: Resource Classification Criteria

CategoryNo. Holes

min.Max. Distance to (m) Avg. Distance

max. (m)closest hole 2nd

closest hole 3rd

closest hole

Indicated 2 25

2 50

3 25

3 80

3 100

4 100

Inferred 1 100 otherwise, remainder estimated and within pit shell

Source: Gray 2014

For the initial Q1 2012 estimate, measures were taken to ensure the resource met the condition of

“reasonable prospects of economic extraction” as required under NI 43-101. At that time a Lerchs-

Grossman pit shell was generated for the purpose of resource tabulation. This pit volume was

generated using MineSight ® software using a copper price of US$4.0/lb applied to copper equivalent

grade and an overall pit slope of 45°. Only blocks within that pit volume are included in the resource

tabulation. New drilling was primarily in-fill and the extent of mineralized blocks beyond the initially

generated pit shell is minimal.

The Halilağa sulphide Mineral Resource is presented in Table 14.12. A gold equivalent cut-off of

0.43 g/t was felt to be reasonable based on a production rate of 50,000 to 70,000 t/d from a pit feeding

a mill and flotation plant where total operating costs would be in the range of $10-12/t. Also, this cut-off

approximately corresponds to the 0.2% copper equivalent cut-off at which the initial resource was

reported. At the 0.43 g/t AuEq cut-off, the strip ratio is 2.4:1 within the 2012 resource shell. The updatedresource is compared to the initial resource in Table 14.14. Metal equivalence parameters are listed in

Table 14.10.

As in the initial resource, oxide blocks above the supergene zone have been tabulated as a separate

Inferred gold resource. This reflects the assumption that base metals within the oxide zone will not be

recoverable. This material is tabled by gold grade cut-off in Table 14.13; a cut-off of 0.2 g/t gold is

judged as reasonable based on other heap leach gold projects including Alamos Gold’s nearby Ağı

Dağı project.

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.

Table 14.12: Halilağa Sulphide Mineral Resource by Gold Equivalent Cut-off*

Cut-OffGrade(g/t AuEq)

Indicated Inferred

Tonnes(Millions)

Cu(%)

Au(g/t)

Mo(%)

AuEq(g/t)

AuEQOunces(1,000s)

Tonnes(Millions)

Cu(%)

Au(g/t)

Mo(%)

AuEq(g/t)

AuEQOunces(1,000s)

0.10 249.6 0.22 0.24 0.005 0.74 5,931 279.8 0.17 0.19 0.007 0.59 5,335

0.25 230.7 0.24 0.26 0.006 0.79 5,822 239.7 0.2 0.21 0.008 0.66 5,117

0.43 182.7 0.27 0.3 0.006 0.9 5,287 178.7 0.23 0.24 0.009 0.77 4,431

0.50 160.1 0.29 0.32 0.006 0.96 4,947 152.5 0.24 0.26 0.009 0.82 4,039

0.75 87.2 0.38 0.41 0.006 1.25 3,503 73.3 0.31 0.34 0.009 1.04 2,460

1.00 44.7 0.51 0.51 0.005 1.62 2,330 30.7 0.38 0.42 0.009 1.29 1,269

1.25 25.4 0.65 0.6 0.003 2.01 1,642 10.1 0.51 0.55 0.007 1.67 539

1.50 16.7 0.79 0.66 0.003 2.35 1,260 4.9 0.62 0.64 0.005 1.99 315

1.75 11.6 0.92 0.69 0.003 2.67 999 2.8 0.73 0.69 0.004 2.27 205

2.00 8.7 1.05 0.7 0.003 2.94 821 1.6 0.86 0.74 0.004 2.59 130

Source: Gray 2014

* The Halilağa sulphide Mineral Resource is presented in Table 14.12. A gold equivalent cut-off of 0.43 g/t was felt to be reasonable

based on a production rate of 50,000 to 70,000 t/d from a pit feeding a mill and flotation plant where total operating costs would be in the

range of $10-12/t. Also, this cut-off approximately corresponds to the 0.2% copper equivalent cut-off at which the initial resource was

reported. At the 0.43 g/t AuEq cut-off, the strip ratio is 2.4:1 within the 2012 resource shell. The updated resource is compared to the

initial resource in Table 14.14. Metal equivalence parameters are listed in Table 14.10.

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.

Table 14.13: Oxide Gold Resource

COG(g/t Au)

Inferred

Tonnes(1,000s)

Au(g/t)

0.1 3,783 0.63

0.2 3,774 0.63

0.3 3,674 0.64

0.4 3,375 0.66

0.5 2,550 0.73

0.6 1,828 0.80

0.7 1,049 0.91

0.8 629 1.02Source: Gray 2014

Table 14.14: Halilağa Sulphide Resource: Update vs. Initial at 0.43 g/t AuEq Cut-off

ResourceModel

Indicated Inferred

Tonnes(Millions)

Cu(%)

Au(g/t)

Mo(%)

AuEq(g/t)

Tonnes(Millions)

Cu (%)

Au(g/t)

Mo(%)

AuEq(g/t)

Update 182.7 0.27 0.30 0.0057 0.90 178.7 0.23 0.24 0.0087 0.77

Initial 168.8 0.30 0.31 0.0054 0.97 199.6 0.23 0.26 0.0067 0.78

Difference +8% -10% -4% +6% -8% -10% -1% -7% +30% -2%

Source: Gray 2014

The mineral resource estimates could be affected by environmental, permitting, legal, title, taxation,

political and other risks; however, at this preliminary level there do not appear to be any material issues.

A list of the major project risks is included in Section 24.

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15-1

.

15 MINING METHODS

15.1 Mining Context

Geotechnical Considerations

This section estimates suitable overall pit slope angles for the 2014 25kt/d pit. Data used for this PEA

level study include:

Geological block model;

Drill hole database containing Total Core Recovery (TCR), Rock Quality Designation (RQD) and

Fault Zones;

Core box photographs; and 2012 Halilağa NI 43-101 Report by Kirkham and Gray.

Lithological units across the pit are divided into Kestane Porphyry (quartz porphyry), Hornfels Halo, with

felsic volcanics either side. Porphyry alteration types are observed across the deposit and are often

quite intense.

Data suggests that RQD is largely independent of lithology, although a slight improvement in the rock

mass is seen in the porphyry relative to the volcanics. Histograms of RQD per lithology highlights the

large proportion of low quality rock mass (RQD<20%) encountered on site. This is likely due to the

intense alteration which reduces rock mass quality. Presence of structures may also reduce the rock

mass quality locally.

Examination of core box photographs and RQD down drill hole in 3D space suggests RQD values in the

upper 100-200 m are generally less than 20% and rock mass quality appears very low. At depths

greater than 100-200 m the rock mass appears to be more moderate in quality with RQD ranges from

20% to 60%. Views of RQD down hole relative to the northern and southern walls are shown in Figure

15.1.

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HALILAĞA

Effective Date

.

Figure 15.

Source: SR

PRO J ECT

: December 20,

: View of R

Pit Show

2014

PEA TECH

2014

QD Data Plo

in Grey)

NICAL REPO

tted Down

RT

rill Hole be ind Northern and South

rn Walls (2

15-2

kt/d

7/24/2019 2015 Halilaga


HALILAĞA

Effective Date

.

Geotechn

Two basic

produced

(Figure 15.

Figure 15.

Source: SR

The Upper

Mass Rati

moderate

domains a

PRO J ECT

: December 20,

cal Domain

geotechnica

on a sectio

2). The maj

: Section V

RQD Val

2014

Domain co

g (RMR89 (

uality rock

re shown in

PEA TECH

2014

s

l domains h

-by section

rity of the 2

iew Showin

es Down Dr

sists of a p

Bieniawski 8

mass with

igure 15.3.

NICAL REPO

ve been ge

basis whic

kt/d pit is

3D Surfac

ill hole

or quality, v

9)) would b

n RMR aro

RT

erated bas

separates

ithin the Up

Defining th

ery broken,

around 20

nd 40-60.

d on the R

the pit into

er Domain.

e Upper and

weak and hi

-30. The Lo

ypical exa

D data. A 3

an Upper

Lower Dom

ghly altered

er Domain

ples of drill

surface ha

nd Lower

ain by Hono

rock mass.

represents

core from t

15-3

s been

omain

uring

Rock

a more

he two

7/24/2019 2015 Halilaga


HALILAĞA

Effective Date

.

Figure 15.

Source: SR

PEA Pit Sl

PEA level

shows the

are given i

Table 15.1

Design Sec

Upper Dom

Lower Dom

Source: SR

PRO J ECT

: December 20,

: Typical E

2014

ope Design

pit slope an

Upper and

n Table 15.1

: SRK Slope

or

in

in

2014

PEA TECH

2014

xample of R

gles are giv

ower Doma

.

Angles per

NICAL REPO

ck Mass Q

n for the t

ins clipped t

omain

RT

ality in Drill

o geotechni

o the 25kt/d

Core from t

ical domain

pit. Recom

Com

he Upper an

discussed

ended com

posite Slope

36

48

d Lower Do

above. Figu

posite slope

Angle (°)

15-4

ains

re 15.4

angles

7/24/2019 2015 Halilaga


HALILAĞA

Effective Date

.

Figure 15.

Source: SR

Seismicit

Turkey is

sense alth

15.2 O

Input Par

The 3D m

deriving th

20 m x 10

Estimatescosts and

on calcula

optimizatio

waste mini

the cost es

PRO J ECT

: December 20,

: Geotechn

2014

ituated in a

ugh much

pen Pit Op

meters

ineral resou

e economic

m.

were maderoyalties. Mi

ed processi

n (see Tabl

ng, where v

timate base

PEA TECH

2014

ical Domain

seismically

ore detail w

imization

rce block m

shell limits

for gold andning, proces

g throughp

15.2). The

ariations in

on in-coun

NICAL REPO

Clipped to t

ctive region.

ill be neede

odel, as de

or the Halil

copper pricsing, and g

t and, alon

OP mining

aulage profi

ry contract

RT

he 25 kt/d pi

This was c

at the next

eloped by

ğa Project.

e, mining dilneral admin

with geotec

costs were

les and equi

ining budg

t

nsidered in

level of stud

ames Gray

The block

lution, heapistration OP

hnical para

estimated f

ipment sele

tary quotes.

the facilities

y.

, A was us

odel dimen

leach proceEX were als

eters, form

r both plan

tion were ta

designs in

d as the b

sions were

ss recovery,o calculated

d the basis

feed mater

ken into acc

15-5

broad

sis for

0 m x

offsitebased

for OP

ial and

ount in

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15-6

.

Table 15.2: Input Parameters Used in the LOM Open Pit Optimization

Item Unit ValuesMetal Prices

Gold US$/oz 1,250

Copper US$/lb 3.00

Recovery to Cu Concentrate

Gold % var.w/ Au grade

Copper % var.w/ Cu grade

Recovery CIL

Gold (cleaner tails CIL) % 15

Cu Concentrate Grade (“conc.”)

Gold g/t var.w/ Au and Cu grade

Copper % 30

Moisture content 8

Smelter PayablesGold in Dore % 99

Gold in Cu conc % 96

Gold deduction in Cu conc. g/t in conc 1

Copper in Cu conc % 96

Treatment and Refining Costs (TC/RC)

Cu conc treatment $/dmt conc 75.00

Cu refining charge $/lb pay Cu 0.075

Au refining charge $/oz pay Au 7.00

Transport, marketing, ins., etc.

Ocean freight to Europe $/wmt conc 30.00

Truck freight to Port $/wmt conc 20.00

Port charges $/wmt conc 6.69

Marine/transportation insurance % 1Transport, marketing, ins, etc. $/wmt conc 57.69

Transport, marketing, ins, etc. $/dmt conc 62.70

Other Parameters

Grade factor (variable) % 95

Royalties % 4

Operating Costs

O/P Waste mining Cost $/waste tonne 2.00

OP Mineralized material Mining Cost $/mill feed tonne 2.00

OP Processing and G&A Cost $/milled tonne 8.54

Pit Slope Angles overall degrees 36 to 48

Dilution % 5

Mining recovery % 100

Strip ratio (est.) t:t 1.3

Processing rate tpd milled 25,000

Processing rate t/yr milled 9,125,000

*The values in this table vary slightly from those used in the economic model as parameters were further refined in theeconomic model. The differences are not considered material to pit shape definitionSource: JDS 2014

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15-7

.

The mineral inventory block model for the Halilağa deposit was then used with CAE Mining NPV

Scheduler (“NPVS”) OP optimization software to determine optimal mining shells. This evaluationincluded the aforementioned parameters. The economic shell limits included indicated and inferred

mineral resources. Inferred mineral resources are considered too speculative geologically to have the

economic considerations applied to them to be categorized as mineral reserves, and there is no

certainty that the inferred resources would be upgraded to a higher resource category.

Cut-off Grade

Table 15.3 summarizes the parameters used, along with incremental (or mill) COG calculations (based

on NSR) and mining dilution. The incremental (or mill) COG incorporates all OPEX except mining. This

incremental cut-off is applied to material contained within an economic pit shell where the decision to

mine a given block was determined by the NPVS optimization. This mill cut-off was applied to all of the

estimates that follow.

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15-8

.

Table 15.3: Cut-off Grade Calculations used in Pit Optimization

Item Unit

Parameters

Ext. COG(includes

mining cost)

Incr. COG(excludes

mining cost)

Revenue, smelting & refining

Copper Price $/lb Cu 3.00 3.00

Payable metal % Cu 96 96

TC/RC/Transport $/lb Cu payable 0.28 0.28

Royalties @4% $/lb 0.1 0.1

Net Cu price $/lb Cu 2.5 2.5

On Site Costs

OP Mining Cost $/t waste mined 2.00 2.00

Strip ratio (estimated) Wt:Ot 1.3 -

OP Mining Cost $/t milled 4.6 -

Milling Cost $/t milled 7.56 7.56

G&A/TSF $/t milled 0.98 0.98

Total Site Costs $/t milled 13.14 8.54

Process and Mining Losses

Process Recovery (average) % Cu 87.20 87.20

Dilution % 5 5

Cut-off grade

Insitu Cut-off Cu grade (Cu only)* % Cu 0.29 0.19

By-product contribution (est) % of Cu value 35 35

Insitu Cut-off Cu grade (equivalent)* % Cu 0.21 0.14

Insitu Cut-off NSR (includes dilution) $/t 13.8 8.97

*estimates only as NSR cut-off used in NPVS

Source: JDS 2014

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15-9

.

Optimization Results

A series of optimized shells were generated for the Halilağa deposit based on varying revenue factors.

The results were analyzed with shells chosen as the basis for ultimate limits and preliminary phase

selection. Refer to Table 15.4 and Figure 15.6 to Figure 15.8 for Inferred and Indicated resources.

NPVS produces both “best case” (i.e. mine out shell 1, the smallest shell, and then mine out each

subsequent shell from the top down, before starting the next shell) and “worst case” (mine each bench

completely to final limits before starting next bench) scenarios. These two scenarios provide a bracket

for the range of possible outcomes. The shells were produced based on varying revenue factors (0.3

through to 1.4 of base case) to produce a series of nested shells with the NPV results shown in Figure

15.5 through Figure 15.8. The NPV values noted here are based on a gold price of $1,250/oz and a

copper price of $3.00/lb as per Table 15.2 and do not include CAPEX and were used only to determine

the mining shapes.

The actual NPV of the project is summarized in the Economics section of this report.

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.

Table 15.4: Overall Optimization Results (excluding capital costs)

Pit#RevFac

Life Total Diluted Mill feed Waste Strip Total

(yr) (Mt) Cu (%) Au (g/t) Cu (Mlb) Au (koz) (Mt) Ratio (Mt)

Pit 1 (5) 0.10 0.0 0.0 1.92 1.09 0 0 0.0 0.60 0.0

Pit 2 (6) 0.12 0.0 0.1 1.64 0.92 3 2 0.0 0.56 0.1

Pit 3 (7) 0.14 0.1 0.7 1.51 0.81 23 18 0.6 0.83 1.2

Pit 4 (8) 0.16 0.3 3.0 1.34 0.69 90 67 1.9 0.63 4.9

Pit 5 (9) 0.18 0.6 5.2 1.24 0.63 142 105 2.9 0.56 8.1

Pit 6 (10) 0.20 0.6 5.9 1.20 0.63 155 118 3.1 0.52 8.9

Pit 7 (11) 0.22 0.7 6.8 1.14 0.62 170 135 3.2 0.47 10.0

Pit 8 (12) 0.24 0.8 7.3 1.10 0.62 178 146 3.3 0.45 10.6

Pit 9 (13) 0.26 1.0 8.7 1.02 0.61 196 171 3.4 0.39 12.1

Pit 10 (14) 0.28 1.0 9.3 0.99 0.61 203 180 3.4 0.37 12.7 Pit 11 (15) 0.30 1.2 10.8 0.93 0.59 221 204 4.1 0.38 14.9

Pit 12 (16) 0.32 1.3 12.3 0.87 0.57 236 227 4.5 0.37 16.8

Pit 13 (17) 0.34 1.6 14.3 0.81 0.56 256 256 4.9 0.35 19.2

Pit 14 (18) 0.36 1.8 16.6 0.75 0.54 277 289 5.4 0.32 22.0

Pit 15 (19) 0.38 2.0 17.9 0.73 0.53 288 306 5.5 0.31 23.5

Pit 16 (20) 0.40 2.3 21.2 0.67 0.51 314 350 6.6 0.31 27.8

Pit 17 (21) 0.42 2.4 22.3 0.66 0.51 323 364 7.2 0.33 29.5

Pit 18 (22) 0.44 2.6 23.8 0.64 0.50 335 382 7.7 0.32 31.5

Pit 19 (23) 0.46 2.9 26.3 0.61 0.48 352 408 7.9 0.30 34.2

Pit 20 (24) 0.48 3.2 28.8 0.58 0.47 369 436 8.7 0.30 37.4

Pit 21 (25) 0.50 3.3 30.5 0.57 0.46 381 454 9.4 0.31 39.9

Pit 22 (26) 0.52 3.8 34.8 0.54 0.45 411 503 12.9 0.37 47.7

Pit 23 (27) 0.54 4.1 37.6 0.52 0.44 429 532 14.3 0.38 51.9 Pit 24 (28) 0.56 5.2 47.7 0.48 0.43 503 665 33.0 0.69 80.7

Pit 25 (29) 0.58 5.4 49.7 0.47 0.43 515 685 34.4 0.69 84.1

Pit 26 (30) 0.60 5.8 52.7 0.46 0.42 536 720 39.9 0.76 92.6

Pit 27 (31) 0.62 6.0 55.0 0.45 0.42 551 745 43.4 0.79 98.4

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.

Pit 28 (32) 0.64 6.3 57.9 0.44 0.41 568 772 46.1 0.80 104.0

Pit 29 (33) 0.66 6.5 59.0 0.44 0.41 575 783 47.6 0.81 106.7 Pit 30 (34) 0.68 6.7 60.9 0.44 0.41 584 797 48.2 0.79 109.1

Pit 31 (35) 0.70 7.1 64.7 0.43 0.40 606 831 53.0 0.82 117.7

Pit 32 (36) 0.72 7.4 67.9 0.42 0.40 624 862 58.2 0.86 126.1

Pit 33 (37) 0.74 7.6 69.6 0.41 0.39 634 877 60.2 0.86 129.9

Pit 34 (38) 0.76 8.1 73.5 0.41 0.39 656 910 66.8 0.91 140.2

Pit 35 (39) 0.78 8.6 78.7 0.39 0.38 684 950 72.4 0.92 151.2

Pit 36 (40) 0.80 8.8 80.0 0.39 0.37 689 959 73.3 0.92 153.3

Pit 37 (41) 0.82 13.6 124.5 0.34 0.34 922 1,360 157.4 1.26 282.0

Pit 38 (42) 0.84 14.0 127.5 0.33 0.34 935 1,379 159.1 1.25 286.6

Pit 39 (43) 0.86 14.8 135.1 0.33 0.33 975 1,441 173.5 1.28 308.6

Pit 40 (44) 0.88 16.7 152.2 0.32 0.32 1,065 1,584 210.1 1.38 362.3

Pit 41 (45) 0.90 19.0 173.2 0.31 0.32 1,168 1,781 259.4 1.50 432.6

Pit 42 (46) 0.92 19.2 175.0 0.30 0.32 1,176 1,793 261.1 1.49 436.1 Pit 43 (47) 0.94 20.7 188.5 0.30 0.31 1,239 1,907 290.0 1.54 478.4

Pit 44 (48) 0.96 21.1 192.3 0.30 0.31 1,255 1,933 294.0 1.53 486.3

Pit 45 (49) 0.98 21.3 194.7 0.29 0.31 1,264 1,946 295.4 1.52 490.0

Pit 46 (50) 1.00 22.2 202.8 0.29 0.31 1,302 2,000 309.7 1.53 512.6

Source: JDS 2014

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HALILAĞA

Effective Date

.

Figure 15.

Source: JDS

PRO J ECT

: December 20,

Pit Optim

2014

PEA TECH

2014

ization Resu

NICAL REPO

lts Graph

RT

15-12

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HALILAĞA

Effective Date

.

Figure 15.

Source: JDS

PRO J ECT

: December 20,

: Open Pit

2014

PEA TECH

2014

ptimization

NICAL REPO

Incremental

RT

Value Results (excludi g capital costs)

15-13

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HALILAĞA

Effective Date

.

Figure 15.

Source: JDS

For the Ha

overall pit,

determine

understan

assumed

CAPEX w

PRO J ECT

: December 20,

: Open Pit

2014

lilağa deposi

but have hi

the optimu

ing of the

maximum

re added.

PEA TECH

2014

ptimization

it, shells bey

her increm

shell on

deposit, the

processing

NICAL REPO

Incremental

ond Pit Shell

ntal strip ra

hich to ba

shells wer

rate of 9.1

RT

Tonnage R

l 37(41) add

tios with mi

se the pha

e analyzed

t/yr. No st

sults

mineralized

imal positiv

ing and sc

in a prelim

ckpiles wer

rock and wa

impact on

heduling an

inary sched

e used in t

ste tonnage

the NPV. T

d to gain a

ule. The sc

e analysis

15-14

to the

better

better

hedule

and no

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.

Based on the analysis of the shells and preliminary schedule, Pit Shell 37(41) was chosen as the base

case shell for further phasing and scheduling for the Halilağa deposit. This shell contains 124 Mt ofmineralized material above cut-off with an average copper grade of 0.34% and 920 Mlbs contained

copper along with a gold grade of 0.34 g/t and 1,357 koz of contained gold. The total waste tonnage in

the shell is 158 Mt for a strip ratio of 1.3:1.

Table 15.5 summarizes the tonnages and grades contained within the shell limits (using the incremental

cut-off value of $8.97/t). There is a total of 3Mt of oxide material that is included in the waste tonnages

above. This material would be segregated in order to provide ability to process in the future.

Table 15.5: LOM Plan Summary

Description Unit Value

Mine Production Life yr 14Process Feed Material Mt 124

Diluted Copper grade % 0.34

Contained Cu Mlbs 920

Diluted Gold grade g/t 0.34

Contained gold koz 1,357

Waste Mt 158

Total material Mt 282

Strip ratio t:t 1.3

Source: JDS 2014

15.3 Open Pit Mine Design

Open pit mine design was conducted using a combination of software packages, including MINTEC Inc.

MineSight™, GEMS™, and NPVS. The base 3D block model was analyzed using GEMS™. The phase

selection and production scheduling was undertaken with the use of MineSight™ and NPVS software.

For the Halilağa deposit, the ultimate shell limits, along with the associated phasing, were based on the

shell analysis described in this report. Preliminary rock storage facilities were then designed to account

for the material produced in each mining phase and shell.

Shell 37 (41) was chosen as the mining shape limit for the Halilağa deposit.

Figure 15.8 and Figure 15.9 represent plan and section views of the ultimate pit shape.

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Figure 15.9: Section View of Halilağa Pit Shell Showing Phase (Stage) Pushbacks

Source: JDS 2014

Note: Only blocks above incremental cut-off shown (i.e. Cu >0.14 %)

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15.4 Mine Sequence/Phasing

The preliminary shells for Halilağa were further analyzed and optimizations were conducted in order to

better define the possible phase designs within the ultimate shell limit. It was decided to divide the pit

sequence into four phases for the mine plan development to maximize the grade in the early years,

reduce the pre-stripping requirements, and to maintain the process facility at full production capacity.

The shell tonnages, grades, and contained metal of the preliminary phases (stages) are summarized in

Table 15.6.

Table 15.6: Halilağa Pit/Phase Tonnages and Grades

StageTotal Diluted Mill Feed Waste Strip Total

(Mt) Au (g/t) Au (koz) Cu (%) Cu (Mlb) (Mt) Ratio (Mt)

Phase 1 30.8 0.44 440 0.55 373 11.4 0.37 42.2Phase 2 27.3 0.35 304 0.30 180 31.2 1.14 58.5

Phase 3 28.0 0.29 261 0.27 164 44.6 1.59 72.6

Phase 4 38.4 0.29 355 0.24 205 70.2 1.83 108.6

Total 124.5 0.34 1,360 0.34 922 157.4 1.3 282.0

Source: JDS 2014

Figure 15.10 further illustrates the phase designs for Halilağa, with tonnes, grades, and strip ratios

shown.

The phases were based on the optimized shells summarized above. Shells selected provide reasonable

pushback widths with mining starting in the higher grade mineralized zone and progressing outwards

from the initial phase (primarily to the east).

During the active mining and processing of the Halilağa deposit, the waste would be placed into the

RSF immediately adjacent (east) to the final shell limits. All mineralized material would be hauled to the

primary crusher immediately to the north east of the deposit.

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HALILAĞA

Effective Date

.

Figure 15.

Source: JDS

15.5 O

The OP m

the basis f

was $2.00

and was

supervisio

and comp

local condi

variations i

PRO J ECT

: December 20,

0: Halilağa

2014

P Mine Op

ining activiti

or this PEA.

t however,

sed in the

. The cost

red to build-

tions. The O

n haulage p

PEA TECH

2014

Project - Ph

ration

s for the Ha

The LOM a

pon further

project econ

stimate was

ups from firs

P mining co

ofiles and e

NICAL REPO

se Summar

lilağa pit we

erage unit

review, $1.8

omics, for

based on di

t principles

sts for both

uipment sel

RT

re assumed

ining cost u

5/t of materi

it and dum

rect quotes

s well as ex

ineralized

ection.

to be under

sed in the d

al mined w

operations

eceived fro

perience of

aterial and

aken by a

evelopment

s determine

, road main

mining co

similar sized

waste minin

ining contra

of the mine

d to be appr

tenance, an

tractors in t

OP operatio

g take into

15-19

ctor as

shapes

opriate

mine

e area

ns and

ccount

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.

Equipment

The major owner-supplied open pit mining equipment requirements are indicated in Table 15.7 and are

based on similar sized OP operations, the proposed processing rate of 9.1 Mt per annum (“Mtpa”),

along with deposit and pit geometry constraints. Note that these estimates are built-up from first

principles based on mining contractor quotes and typical fleet sizes used in the area and are only as a

guide to approximate open pit mining fleet requirements.

The mining contractor will ultimately determine overall mining fleet requirements.

The mining fleet has an estimated maximum capacity of 70,000 t/d total material, which would be

sufficient for the mine production plan.

Table 15.7: Major OP Equipment Assumptions (Contractor Estimates may Vary)

Equipment Type No. of initial units

250 mm dia. Rotary, Crawler Drill (diesel) 3

115 mm dia. Rotary, Crawler Drill (diesel) 1

7 m3 Front Shovel (diesel) 4

7 m3

Wheel Loader (diesel) 2

40 t Haul Truck 35

D9-class Dozer 4

14H-class Grader 3

814F-class Rubber Tire Dozer 240 t Water Truck 1

Source: JDS 2014

Unit Operations

250 mm diameter blast hole drills are planned to perform the bulk of the production drilling in the mine

(both mineralized and waste rock). The hydraulic drill with a 115 mm diameter bit would be used for

secondary blasting requirements and may be used on the tighter spaced patterns required for pit

development blasts. The main loading and haulage fleet is planned to consist of 40 t haul trucks, loaded

primarily with the diesel powered 7 m3 front shovels or the 7 m3 wheel loader, depending on pitconditions.

As pit conditions dictate, the D9-class dozers are planned to rip and push material to the excavators

and maintaining the waste dump.

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.

The additional equipment listed in Table 15.7 is planned to be used to maintain and build access roads

and to meet various site facility requirements, including stockpile maintenance and further explorationdevelopment.

15.6 Mine Schedule

The production schedule for the Halilağa deposit was developed with the aid of NPVS and MineSight™

software, and incorporated the various pits and phases mentioned above.

Indicated and Inferred resources were used in the LOM plan, with Indicated resources making up 69%

of the total LOM tonnage processed. The resources included in the mine plan have an estimated

external dilution factor of 5%.

With the higher grade mineralized material near surface at Halilağa, Year -1 represents thecommencement of pre-stripping as well as the stockpiling of mineralized material. The LOM maximum

planned amount of total material to be moved is approximately 65,000 t/d. The average mining rate is

planned to be 55,000 t/d.

Table 15.8 is a summary of total material movement by year for the LOM open pit production schedule

(both as totals, as well as by each phase) and the proposed processing schedule.

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Table 15.8: Proposed LOM Open Pit Production Schedule

Item Unit TotalYear

-1 1 2 3 4 5 6 7 8 9 10 11 12 13 14

Mining Total

Mineralized Material Mt 124.3 1 8.1 9.1 9.1 9.1 9.1 9.1 9.1 9.1 9.1 9.1 9.1 9.1 9.1 5.6

Gold feed grade g/t 0.34 0.64 0.46 0.46 0.41 0.36 0.33 0.32 0.35 0.27 0.33 0.28 0.31 0.28 0.26 0.31

Contained gold koz 1,357 21 121 134 120 107 96 95 103 80 97 83 91 81 75 55

Copper feed grade % 0.34 0.95 0.73 0.57 0.38 0.31 0.28 0.28 0.31 0.26 0.29 0.23 0.25 0.24 0.24 0.26

Contained copper Mlb 920 21 131 116 76 63 56 57 63 52 59 47 50 49 49 33

Waste Material Mt 157.6 8.7 9.3 13.8 14.3 14.7 14.7 14.7 14.7 14.8 14.7 14.7 7.4 0.9 0.2

Total Material Mt 281.9 9.7 17.4 23 23.5 23.8 23.8 23.9 23.8 23.9 23.8 23.8 16.6 10 9.3 5.6

Strip ratio t:t 1.3 8.7 1.1 1.5 1.6 1.6 1.6 1.6 1.6 1.6 1.6 1.6 0.8 0.1 0

Total Material Mined tpd 26,486 47,712 62,918 64,321 65,317 65,210 65,371 65,273 65,597 65,234 65,231 45,388 27,420 25,433 15,468

Phase 1

Mineralized Material Mt 30.8 1 8.1 9.1 7.8 4.8

Gold feed grade g/t 0.44 0.65 0.46 0.46 0.43 0.37

Contained gold koz 441 21 121 134 108 57

Copper feed grade % 0.55 0.94 0.73 0.57 0.4 0.34

Contained copper Mlb 373 21 131 116 69 36

Waste Material Mt 11.6 4.9 4.8 1.9

Total Material Mt 42.5 5.9 12.9 11 7.8 4.8

Strip ratio t:t 0.4 4.9 0.6 0.2

Phase 2

Mineralized Material Mt 27.3 1.3 4.1 7.4 5.6 8.9

Gold feed grade g/t 0.35 0.3 0.35 0.33 0.36 0.35

Contained gold koz 304 13 47 79 64 101

Copper feed grade % 0.3 0.25 0.29 0.29 0.31 0.31

Contained copper Mlb 180 7 26 47 38 62

Waste Material Mt 33.3 3.7 2.2 5.7 12.1 4.7 2.5 2.3

Total Material Mt 60.6 3.7 2.2 5.7 13.5 8.8 9.9 7.9 8.9

Strip ratio t:t 1.2 9.2 1.1 0.3 0.4

Phase 3

Mineralized Material Mt 28 0.3 1.7 3.6 0.2 9.1 9 4.1 0

Gold feed grade g/t 0.29 0.38 0.3 0.27 0.25 0.27 0.33 0.25 0.17

Contained gold koz 261 3 17 31 2 80 96 34 0

Copper feed grade % 0.27 0.15 0.24 0.24 0.24 0.26 0.29 0.26 0.21

Contained copper Mlbs 164 1 9 19 1 52 58 23 0

Waste Material Mt 42.6 2.3 6.3 2.2 10.2 9.7 5.2 0.1 5.3 0.9 0.4

Total Material Mt 70.6 2.3 6.3 2.2 10.5 11.4 8.7 0.3 14.4 9.9 4.5 0

Strip ratio t:t 1.5 40.4 5.7 1.5 0.5 0.6 0.1 0.1

Phase 4

Mineralized Material Mt 38.1 0.1 5 9.1 9.1 9.1 5.6

Gold feed grade g/t 0.29 0.27 0.3 0.31 0.28 0.26 0.31

Contained gold koz 352 1 49 91 81 75 55

Copper feed grade % 0.24 0.23 0.21 0.25 0.24 0.24 0.26

Contained copper Mlb 204 0 23 50 49 49 33

Waste Material Mt 70.4 2.4 7.2 14.7 9.5 13.8 14.2 7.4 0.9 0.2

Total Material Mt 108.6 2.4 7.2 14.7 9.5 13.9 19.3 16.6 10 9.3 5.6

Strip ratio t:t 1.8 182.5 2.8 0.8 0.1 0

Mill Schedule

Mill Feed Total

Mill Feed Mt 124.3 9.1 9.1 9.1 9.1 9.1 9.1 9.1 9.1 9.1 9.1 9.1 9.1 9.1 5.6

Gold feed grade g/t 0.34 0.48 0.46 0.41 0.36 0.33 0.32 0.35 0.27 0.33 0.28 0.31 0.28 0.26 0.31

Contained gold koz 1,357 141 134 120 107 96 95 103 80 97 83 91 81 75 55

Copper feed grade % 0.34 0.76 0.57 0.38 0.31 0.28 0.28 0.31 0.26 0.29 0.23 0.25 0.24 0.24 0.26

Contained copper Mlb 920 152 116 76 63 56 57 63 52 59 47 50 49 49 33

Source: JDS 2014

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HALILAĞA

Effective Date

.

The Halila

158 Mt offocuses o

in the sc

stockpiled

phases ar

Figure 15.

15.12 illus

15.13 illus

average, i

approxima

Figure 15.

Source: JDS

PRO J ECT

: December 20,

a deposit is

waste (1.3: achieving c

edule, bala

material wa

active in a

11 summari

rates the fe

rates annu

estimated

ely 68 koz

1: Process

2014

PEA TECH

2014

planned to

overall strionsistent pr

cing grade

s included i

y one year.

zes process

ed tonnage

l bench adv

to produce

er year.

Tonnes, Wa

NICAL REPO

roduce a to

p ratio) ovecessing fee

and strip r

this prelimi

plant tonn

by phase a

ance and m

otal recove

te Tonnes

RT

tal of 124 M

a 14 yearproduction

atios, while

nary sched

ge, waste t

d period, a

aterial mov

ed copper

nd Strip Ra

of minerali

mine operat rates, minin

trying to

le. In order

onnages, a

s well as ov

ment. Duri

f 58 Mlbs

io by Perio

ed process

ing life. Theg of higher

aximize NP

to achieve t

d strip rati

erall gold g

g full produ

er year and

plant feed r

current LOrade materi

V. No blen

argets, up t

by period.

ades, while

ction, the m

recovered

15-23

ck and

M planl early

ing of

o three

Figure

Figure

ine, on

gold of

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HALILAĞA

Effective Date

.

Figure 15.

Source: JDS

PRO J ECT

: December 20,

2: Mineraliz

2014

PEA TECH

2014

ed Tonnes

NICAL REPO

nd Grade b

RT

Phase and Period

15-24

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HALILAĞA

Effective Date

.

Figure 15.

Source: JDS

To further

rock stora

The Halila

pit is sche

while tryin

PRO J ECT

: December 20,

3: Annual

2014

illustrate the

e facilities a

a deposit p

uled to be

to maximiz

PEA TECH

2014

enches and

progression

t the end of

rovides max

ined out in

e the NPV of

NICAL REPO

Total Materi

of mining,

arious perio

imum return

series of p

the project.

RT

al Mined by

igures 15.1

ds.

when the

sh-backs in

Phase and

4 to 15.18 p

arious phas

order to ac

eriod

rovide snap

es are mine

ieve the req

hots of the

concurrent

uired proces

15-25

pit and

ly. The

s feed,

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.

Mine Development Schedule

Year -1: Development of the Halilağa Project is planned to commence with pre-stripping and

mine production of the open pit. A total of 1.0 Mt of mineralized plant feed is scheduled to be placed in

stockpile while 8.7 Mt of waste are pre-stripped from Phase 1 and 2. The average copper grade is

estimated to be 0.95% and the gold grade is estimated to be 0.64 g/t.

Year 1: The 9.1 Mtpa target plant feed is envisioned to be attained with mining in Phases 1 through 3

(waste only mining in Phases 2 and 3). The average mine grade in Year 1 is estimated to be 0.73% Cu

and 0.46 g/t Au. A total of 9.3 Mt of waste rock is scheduled at an average strip ratio of 1.1:1.

Year 2: Process plant feed production is scheduled to be maintained at the target of 9.1 Mtpa (or

25kt/d). Mining continues in Phases 1 through 3 with mill feed being produced from Phase 1. Total

waste planned to be mined from the active phases is 13.8 Mt. The average gold grade is estimated to

be 0.46 g/t Au, with a copper feed grade of 0.57 %Cu. Stripping of push backs is planned to increaseoverall strip ratio to 1.5:1. Production rates are envisioned to increase to an average of 63,000 t/d total

material for the year.

Year 3: Mining of Phase 1 is reduced to strictly mill feed tonnes while mining continues in Phase 2 and

3. Process plant feed production is planned to be maintained at steady state of 9.1 Mt total delivered.

14.3 Mt of waste is planned to be mined at an overall strip ratio of 1.6:1. Production rates for the year

are scheduled to peak at 65,000 t/d total material.

Years 4 to 9: Mining in Phases 1 and 2 is scheduled to be completed over this time frame, while

Phase 3 and 4 would see some activity, with an average of 23.8 Mt of total material mined with an

overall strip ratio of 1.6:1. The average gold grade is expected to be 0.33 g/t Au while average copper

grade is 0.29 % Cu.

Years 10 to 14: Mining is planned to be concentrated in Phases 3 and 4 over the final years of the LOM

plan. The average annual strip ratio is expected to decrease to 0.5:1 with an average of 6 Mt of waste

and 8.4 Mt of process plant feed planned to be mined in each period. Average mining rate is scheduled

to be 36,000 t/d.

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.

16 MINERAL RECOVERY METHODS16.1 Process Plant Design Considerations

The Halilağa process plant and associated service facilities are designed to process 25,000 t/d of ROM

material, to produce copper concentrate, gold doré and tailings. The proposed process includes

crushing and grinding of the ROM material, rougher and cleaner flotation, regrinding, cyanide leaching,

cyanide detoxification, gold room and dewatering of copper sulphides and is amenable to the

mineralization type at Halilağa. The flotation and cyanide destruction tailings would be thickened before

placement in the TSF.

Process Plant Design Criteria

The key process criteria for the plant design and OPEX are provided in Table 16.1.

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.

Table 16.1: Summary of the Process Design Criteria

Criteria Description Units Design Source

Plant Throughputkt/d 25 Pilot Gold Inc.

Mt/a 9.125 Pilot Gold Inc.

Crusher Availability % 75 JDS

Crusher Throughput t/h 1389 Calculation

Crusher SelectionSize 54” X 75” JDS

Number 1 JDS

Mill/Flotation Availability % 92 JDS

Mill Throughput t/h 1132 JDS

Physical Characteristics BWI kWh/t 12.2 Test work

Primary Grind Size P80 µm 150 Test work

Concentrate Grind Size P80 µm 20 Test work

Head Grade (Ave/Max)% Cu 0.34/0.9 JDS

g/t Au 0.34/0.55 JDS

Flotation Recovery Copper % 88.2 JDS /Test Work

(Average Au recovery in CuConcentrate

Gold % 58.4 JDS /Test Work

Concentrate Grade Copper % 30.4 JDS/ Test Work

Leaching Recovery Gold % 15 JDS/ Test Work

Cu Circuit Residence time

Roughers mins 24 JDS/ Test Work

Cleaner 1 & Scav. mins 20 JDS /Test workCleaner 2 mins 10 JDS /Test work

Cleaner 3 mins 10 JDS /Test work

Cu Concentrate Filtration Rate kg/m2 /h 120 JDS

Concentrates Thickening Flux t/m2 /h 0.2 JDS

Tailings Thickening Flux t/m2 /h 0.8 JDS

Tailings Thickener Underflow Density % w/w 60 JDS

Leach Tanks Feed Thickener Flux t/m2 /h 0.25 JDS/Test Work

Leach Tanks Retention Time hrs 12 JDS/Test Work

Cyanide Detox Tank Retention Time min 120 JDS/Test Work

Source: JDS 2014

Detailed Process Design Criteria incorporating the process mass balance, engineering design criteria

and key design, are derived from the results of the metallurgical test work program. Sections 13.2.3 and

13.3 and Figures outline the flowsheet development and describe the planned plant processes and

layout.

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16.2 Processing Plant Configuration

Primary Crushing

The gyratory crusher is proposed as a permanent installation that would take ROM material and

produce a product of 80% passing 150 mm. Haul trucks are planned to supply ROM material to the

primary crusher dump pocket, where they would unload into one of two dump aprons. The dump pocket

would have a hydraulic rock breaker to reduce any oversize rocks that may clog the crusher feed. The

gyratory crusher would process the ROM mill feed rock at a rate of 1,390 t/h. The crushed material

would discharge from the underside of the crusher hopper onto the sacrificial primary crusher discharge

belt conveyor. The material would then feed into the coarse mill feed stockpile belt conveyor which

would elevate the material to deposit onto the coarse mill feed stockpile.

A dust collection and suppression system would be installed to control fugitive dust generated at thecrusher, material transfer points and other relative operations.

The primary crushing installation is proposed to include the following key equipment:

One gyratory crusher - 1,370 x 1,900mm (54” x 75”- 600 hp);

One sacrificial conveyor ;

One hydraulic rock breaker ;

One stockpile feed conveyor ; and

One dust collection/suppression system.

Stockpile and Reclaim

The coarse mill feed stockpile would hold one day of live storage of the crushed material, or 25,000 t.

Two apron feeders would reclaim the material with one operating, and one on standby, during normal

operation. The apron feeders would meter the flow onto the SAG mill feed conveyor equipped with a

belt scale, at a controlled rate.

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Primary Grinding and Classification

The primary grinding circuit is proposed to incorporate a SAG mill and one ball mill. The process rate

would be 1,132 t/h (25,000 t/d).

The SAG mill would be fed at a controlled rate by the reclaim apron feeders under the coarse mill feed

stockpile. Lime would be added to the SAG mill feed belt conveyor to raise the pH of the slurry, which

would aid copper flotation. A SAG mill ball bin and feeder would feed fresh grinding media onto the

SAG mill feed belt conveyor to maintain the grinding charge.

The SAG mill discharge containing 70% solids by weight would pass over a screen to remove oversize

pebbles. The pebbles would be conveyed outside the building to a discharge pile for manual re-entry

into the process, storage for future processing, or disposal, depending on the mineralized rock

characteristics.

The SAG mill screen underflow would combine with ball mill discharge into one common pump box. The

ball mill would be in closed circuit with cyclone cluster and slurry underflow stream. The overflow slurry

stream would feed the copper rougher/scavenger flotation circuit. The cyclone overflow particle size is

proposed to be P80 150μm and contain approximately 32% solids by weight. Cyclone underflow to the

ball mills would be approximately 72% solids by weight, and the circulating load would be approximately

250% of new mill feed. Ball charge systems would add grinding media as required for maintaining

grinding charge.

The grinding circuit would include the following key equipment:

One SAG Mill – 30’ Ø x 18’ EGL, 10,000 hp;

One vibrating screen;

One ball mill – 22Ø x 36 ft., 10,000 hp ;

Two cyclone feed slurry pumps (one operating, one standby) ; and

One cyclone cluster – 14 x 26”.

Flotation and Regrind

Rougher Conditioning

Flotation feed is planned to be conditioned in a rougher conditioner tank where flotation reagents wouldbe added. Frother would be added to the conditioner tanks overflow pipes feeding the first rougher

flotation cell.

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Rougher Flotation

The rougher conditioner tank would overflow to the rougher flotation cells connected in series. Five

300 m3 forced air tank cells have been selected to provide the required residence time for the roughing

flotation duty. The cells would be arranged with a step in level between each pair of cells.

Concentrate from the rougher cells would flow by gravity to a rougher concentrate launder and pumped

to the regrind circuit.

Vertical spindle sump pumps would be provided in the rougher flotation area to facilitate clean up.

Regrind

Concentrate from the rougher cells would be pumped to a regrind mill circuit to achieve the fine regrind

size of P80 20 µm. Regrinding will be achieved in a 18’ dia x 37’ EGL – 7,500 hp ball mill (type of the

regrind mill will be finalized after performing regrind test).

Cleaner Flotation

Reground concentrate is designed to be mixed with flotation reagents: collector and frother, before

being pumped to the cleaner 1 flotation cells. Concentrate from the cleaner 1 flotation cells would flow

via gravity to the cleaner 2 feed pump. Tailings from the cleaner 1 flotation cells are planned to flow by

gravity to the cleaner scavenger flotation cells.

Concentrate from the cleaner scavenger cells would flow by gravity to the cleaner scavenger

concentrate launder and would be pumped to the first cleaner feed box. Tailings from the cleaner

scavenger cells are designed to flow via gravity to a leach thickener pumpbox.

Concentrate from the cleaner 2 cells are planned to be pumped by the cleaner 3 feed pump to the

cleaner 3 flotation cells. Tailings from the cleaner 2 cells would gravity flow to the cleaner 1 distributor

box.

Concentrate from the cleaner 3 cells would be pumped to the concentrate thickener feed box. Tailings

from the cleaner 3 cells would report to the cleaner 2 flotation circuit.

The cleaner flotation circuit is designed to include the following equipment:

One bank of four (4) 50 m3 first cleaner and one bank of two (2) 50 m3 scavenger flotation cells

One bank of four (4) 20 m3 second cleaner cells and;

One bank of four (4) 10 m3 third cleaner cells.

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On Stream Analysis & Particle Size Analysis

The On Stream Analysis (OSA) system would provide online copper assay analysis.

In addition, on-line Particle Size Analysis (PSA) would be performed on the rougher feed slurry sample.

Analysed samples, exiting the OSA and PSA systems, would be discharged into sample return hoppers

and pumped back to the relevant part of the process.

The OSA and PSA systems would be located at an optimal position which would permit the maximum

use of gravity flow of sample feed and sample reject slurries and minimize requirements for sample

pumps and hoppers.

Thickening and Concentrate Filtration

Copper concentrate from cleaner 3 would be pumped to a high-rate concentrate thickener via the

concentrate thickener feed box. Thickener overflow would be pumped to the process water tank for

storage and re-use in the circuit, while thickener underflow would be pumped to the copper concentrate

filter feed tank. The copper concentrate filter feed tank would have approximately 12-hours storage

capacity.

Thickened copper concentrate slurry would be delivered to the copper concentrate filter. The filter press

would reduce the moisture content of the concentrate prior to transport. Filter cake would discharge

through the floor of the filter building into a concrete area underneath the press. Filter cake would be

removed from the bunker by front-end loader (“FEL”) and stored in the covered concentrate storage

shed.

The concentrate filter press and concentrate storage stockpile are planned to be housed in a building,

which would be fully sheeted on all sides for protection from wind and rain. The concentrate would be

stockpiled in the shed by FEL to provide covered on-site storage capacity. Concentrate would be

reclaimed by FEL from the stockpile and loaded into concentrate trucks for transport off-site.

The moisture limit has been assumed to be 12% w/w in the filter cake for the filter duty.

Gold Leaching and Recovery Circuit

The first copper cleaner tailings would be pumped to a dedicated thickener and thickened to about 50%

solids, giving a density at which the carbon should have neutral buoyancy. Approximately 20% of the

feed gold would be in the leach thickener underflow. Leaching is planned to take place in six carbon-in-leach (CIL) tanks. All tanks would be arranged in a pattern to minimize the footprint, and would sit on a

series of descending steps. Any one tank could be taken off line for maintenance. Average residence

time for the CIL circuit would be 12-hours.

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At the end of the CIL circuit, the pulp would enter into a two tank cyanide destruct module with 2-hours

residence time using SO2 and copper sulfate as the detox reagents to treat leaching circuit tails prior todischarge to the tailings storage facility.

Loaded carbon from the leach circuit would be sent to a carbon plant where gold would be recovered.

Water overflowing the leach feed thickener would be pumped to process water tank.

Loaded carbon would be sent to an acid wash vessel and treated by hydrochloric acid solution to

remove scale and other impurities. After neutralization the carbon would be pumped to a Zadra Strip

vessel. Gold would be stripped from the carbon by circulating a hot caustic solution through the vessel

at about 135°C and a pressure of 345-480 kPa. The strip solution would be heated using a combination

of plate and frame heat exchangers and an electric hot water heater. After reaching stripping

temperature, the solution would flow upward through the strip vessel. The gold-laden solution would exit

the top of the strip vessel, flow through the cool down heat exchanger, and flow by gravity toelectrowinning cells.

Gold would plate onto stainless steel cathodes or fall to the tank bottom as a fine sludge. Strip solution

from the electrowinning cells would gravity flow to a barren solution tank. Gold laden stainless steel

cathodes would be taken to a cathode wash tank then cleaned by a high pressure spray. The resulting

gold sludge would be separated from the wash solution by a plate and frame filter press. Sludge would

be collected, mixed with fluxes, and then melted in an electric induction furnace to produce a doré bar

suitable for shipping to a refinery.

After every second strip, carbon from the strip vessel would be transferred to a rotary regeneration kiln

at approximately 700°C. It would then be quenched in water, screened to remove fines, and stored in a

carbon storage tank to be re-used in the leach circuits. New carbon would be periodically added to the

circuit to make up for the fines taken out of the circuit by a sizing screen. The fines would be dried and

stored in sacks or barrels for off-site treatment or sale.

Raw and Process Water

The mill operation is planned to be supplied with two separate water supply systems.

A raw water system for the mill, flotation and leaching areas (mainly reagent mixing, gland seal

and spray water); and

A process water system for the entire process plant.

The majority of the plant water requirements are to be met from the plant process water system,

which would be composed of recycled water streams, i.e. thickener overflows and TSF return

water, supplemented with raw water as required. Raw water would be harvested from site

runoff. Raw water would be sourced from the water storage pits located 5 km northeast of the

process plant, pumped to the plant and used for unit processes that require clean water, such

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as reagent mixing and gland water. The water storage pits would provide to ensure sufficient

capacity for continued plant operations throughout the year.

Plant process water would be contained in the plant process water tank, which would provide

approximately two hours of storage capacity for plant process water requirements at nominal

flow rates.

Tailings

The gold circuit tailings are designed to enter a cyanide destruct module as previously noted. Detoxified

tailings would then be combined with the copper rougher tails and pumped to a high rate tailings

thickener prior to pumping to the tailings storage facility. The reclaim water pumps would be housed on

a reclaim barge at the TSF. The pumps would send reclaim water back to the process water tank.

Process water from the thickener overflow would flow by gravity to the plant process water tank.

Reagents

Reagents used in the process would include:

Hydrated Lime;

Copper Sulfate;

AP404;

Frother (MIBC);

Potassium Amyl Xanthate;

Flocculant(s);

Caustic Soda;

Sodium Cyanide;

Elemental sulfur (to make SO2); and

Hydrochloric Acid.

Each reagent is proposed to have its own preparation system which includes a bulk handling system, a

mixing tank if required, and a storage tank. Raw water would be used for reagent preparation. Themixing and holding tanks would be equipped with level indicators and instrumentation to ensure that

spills do not occur during normal operation. Each reagent solution would be reticulated in a ring main

system to the various dosing points using a ring main pump (one duty and one standby). Line pressure

would be maintained by an automated diaphragm valve. Dosing to each point would be by flow control

valve and flow meter.

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The reagent preparation areas would be equipped with appropriate ventilation, eye‐wash stations,

safety showers, fire and safety protection, and Material Safety Data Sheets (MSDS).

Hydrated lime is planned to be added to the SAG mill feed belt. Lime would be delivered in bulk and

pneumatically unloaded into a 200 m3 silo. The lime powder would also be fed by a variable speed

rotary feeder to a mixing tank where water is added. Mixed “milk-of-lime” slurry would be transferred to

the lime slurry storage tank with 24-hour capacity. The lime slurry would be reticulated in a ring main

system to various user points using the lime ring main pump.

Grinding Media

Forged carbon steel grinding media are delivered to site in 20 t containers. The balls are unloaded into

a storage hopper via a vendor-supplied hydraulically operated container unloader.

Ball charging mechanism in the primary milling and overhead crane in the regrind areas would be used

to charge steel balls into the SAG, ball and regrind mills.

Ancillary Services

Low Pressure Air

Low pressure air at a suitable pressure would be supplied to the flotation cells from the flotation

blowers. There would be two blowers (one duty and one standby) installed to meet the flotation air

requirements. Pressure control valves would be installed in the air distribution lines to cater for different

air pressure requirements of different flotation cells. The blowers would be multiple stage centrifugal

type equipped with variable speed drives to cater for fluctuations in flotation air demand.

The blowers would be housed inside an acoustic enclosure to reduce noise to an acceptable level. The

enclosure will have ventilation for cooling.

High Pressure Air

High pressure air for leaching tanks, plant and instrument air requirements would be provided by four

rotary screw air compressors. There would be three duty compressors and one standby compressor

operating in lead-lag mode. Plant air would be stored in the plant air receiver prior to being reticulated

throughout the plant.

Plant BuildingsA number of plant buildings would be required for operation and maintenance. These include

administration office, laboratory, plant workshop, warehouse, reagent stores, guard house, and plant

control room and ablutions.

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The administration building is planned to contain a number of internal offices and meeting rooms. A first

aid facility would be provided near to this building. The warehouse would house mechanical, electrical,instrumentation and general items in discrete areas. The warehouse structure would be adjacent to the

plant maintenance workshop. Internal offices would be supplied for warehouse and maintenance staff.

The plant control room would be located at an elevated position adjacent to the hydro-cyclone clusters

such that the stockpile, grinding, flotation and thickening areas can be easily viewed from the control

room. Any external windows shall be double glazed with tinted glass.

Process Control Philosophy

The control philosophy to be implemented for the Halilağa Project is planned to be typical of those used

in modern mineral processing operations.

Field instruments would provide inputs to a set of Programmable Logic Controllers (PLCs). Process

control cubicles are envisioned to be located in the Motor Control Centres (MCCs) and contain the PLC

hardware, power supplies and I/O cards for instrument monitoring and loop control.

The PLCs are planned to perform the control functions by:

Collecting status information of drives, instruments and packaged equipment;

Providing drive control and process interlocking; and

Providing PID (proportional-integral-derivative) control for process control loops.

Standard Personal Computers (SPC) would be located in the Main Control Room (MCR) and the

Crusher Control Room (CCR). The PCs would be networked to the PLCs and operate a Supervisory

Control and Data Acquisition (SCADA) program that provides an interface to the PLCs for control and

monitoring of the plant.

The SCADA is planned to provide outputs to alarms, monitoring, and control the function of the process.

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17 PROJECT INFRASTRUCTURE

17.1 General

The project envisions construction of the following key infrastructure items:

9 km of 6 m wide access roads;



154 kV to 6.6 kV 50 MVA substation;

2 MW transformer for construction power;

34 km of new 154 kV power line;

15 km of new 6.6 kV power line;

30 MW substation;

Mineral processing plant and associated infrastructure;

Haul truck parking facility/laydown;

Warehouse and maintenance shop;

Mine dry and administration building;

ANFO storage and transfer pad;

Dual purpose fresh/firewater tank;

Process water tank;

Potable water skid and distribution;

Sewage treatment plant;

Tailings storage facility (TSF);

Rock storage facility (RSF); and

Water treatment plant.

17.2 General Site Arrangement

The overall proposed site arrangement is shown in Figure 17.1.

The site is configured for optimum construction access and operational efficiency. Primary buildings

would be located to allow easy access from the existing mine access road and utilize existing

topography to minimize bulk earthworks volumes. The process plant is planned to be located as close

and as safely as possible to the pit and at an elevation that facilitates mill feed rock conveying. Existing

roads are envisioned to be upgraded and reused wherever possible.

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17.3 Site Access Road

The existing 2 km mine access road from Izmir-Ҫanakkale Highway to the process plant location would

receive significant upgrades to accommodate increased traffic. The road is proposed to be widened to

10 m with new gravel, grading and compaction. It would be suitable for transportation of mining

equipment, fuel trucks, mobilization of construction equipment, concentrate transportation and ongoing

operational requirements.

17.4 Light Vehicle Roads

The existing 7.5 km road to the water storage pits 1, 2 and 3 would be upgraded to a 6 m wide gravel

road. A 1 m wide bench would be constructed along the water storage pit access road to allow for

installation and maintenance of the makeup water pipeline. A new 1.2 km long, 6 m wide light vehicle

road is planned to be constructed adjacent to both stockpile and mill feed conveyors. The tailings facilitywould be accessed using a new 6.5 km long, 10 m wide gravel road. A short side road off the tailings

access road would be used as an entrance to the explosives storage facility.

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17.5 Power Supply and Transmission Line

JDS estimated that the project would require about 30 MW of power at full production. There are

excellent potential sources of electrical power in the vicinity of the project.

There is an existing 154 kV power transmission line owned by the Turkish Electricity Transmission

Corporation (TEIAS) that runs through the project site and crosses the planned open pit. This high

voltage line transmits electrical power from the Ҫan coal-powered plant to the towns of Ezine, Ayvacık

and Bayramiç, which are 30 km to 60 km to the west-southwest of the project site. The transmission line

would require re-routing of about 9 km prior to commencement of mining to avoid the planned pit and

waste facilities.

A private company, Uludağ Elektrik Dağıtım A.Ş (UEDAS) owns the Çan power plant and is the sole

authorized provider allowed to sell power in Çanakkale Province. The Çan power plant has a generationcapacity of 50 MW of which 39 MW is being used. Make-up power for the project is assumed to be

sourced from either an expansion of the Çan plant or a tie-in to the national grid. It is assumed that a

new 154 kV, 25 km long transmission line from the Ҫan area would be required including step-down

transformers.

Further investigation whether the existing 154 kV power line could be used for construction, operations

or both as well as the ultimate source of power would need to be done at the next level of study. The

input of TEIAS and UEDAS will be critical in this future analysis.

Sub-Station and Distribution

The proposed new 154 kV overhead power line would tie into an on-site substation that would reducevoltage to 6.6 kV at a peak of 50 MVA. The sub-station is planned to be located as close as possible to

the process plant and adjacent to the primary e-house. Power delivery to the process plant would be via

buried teck cables.

A 6 km overhead line from the mine substation is proposed to transfer power at 6.6 kV to a small

transformer located at the water storage (coal) pits. This transformer would have associated switchgear

that would supply power at 380 V to the make-up water pumps in the water storage pits.

Electrical power to the TSF reclaim water barge would be via a 7.5 km, 6.6 kV from the main substation.

A small transformer and electrical building (sea container) would be located near the barge. Further

study is required to refine supply voltage.

A dedicated power line would transmit 6.6 kV from the substation to the reclaim tunnel, primary crusher

and truck shop. Approximately 1.5 km of overhead line would be installed with associated transformers

and electrical buildings (sea containers).

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17.5.1 Construction Power

Stand-alone diesel generators are envisioned to supply 2 MW of power during plant and substationconstruction. These would be rented to reduce project capital costs. It may be possible to tap into the

existing power line for construction needs and this option should be investigated further in the next level

of study.

17.6 Camp

No camp facilities are planned to be installed for construction or operational purposed. The multiple

nearby towns and villages are assumed to have sufficient capacity to accommodate construction and

operations personal.

17.7 Process Plant

The process plant is planned to be located on a plateau near the site access road. It contains milling,

flotation, regrind, concentrate thickening, concentrate thickening, filter presses, concentrate

storage/loadout, reagent storage, electrical rooms, oxygen plant and refinery. The building would be a

150 m x 50 m stick built steel structure.

17.8 Truck Shop

A 350 m x 150 m pad would be prepared for the mining contractor to install all necessary infrastructures

such as laydown and shop facilities. Water and a 380 V power line would be supplied to the pad.

17.9 Maintenance Shop, Warehouse, Mine Dry and Administration Building

The shop and warehouse would be contained in one common pre-engineered 46 m x 30 m building.

The shop is designed to house three separate bays. Tire changing and large vehicle assembly would

take place outdoors and utilize rough terrain mobile equipment. A general warehouse would be included

within the same building in a partitioned bay.

The mine dry and administration building is planned to be located beside the maintenance shop to allow

for quick access to and/from each facility. It would be a 24 m x 15 m concrete building constructed

using standard Turkish fabrication methods.

17.10 Communications / IT

The offices would include a wired and wireless computer network and phone system.

A hand-held radio system would be used for voice-communication between personnel in the field.

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17.11 First Aid / Emergency Services

A qualified nurse or first-aid attendant would be provided on-site. The first aid room would be located

besides the administration building. The ambulance and fire truck would be parked at the ready in the

maintenance shop.

Buildings are planned to have smoke, carbon monoxide and heat detectors, overhead sprinklers,

hydrants / hoses and appropriate chemical fire extinguishers.

17.12 Bulk Explosives Storage and Magazines

Explosives would be stored at a secured and monitored site located approximately 800 m from the main

plant and populated, high traffic areas. Access to a 50 m x 50 m prepared pad would be via the tailings

access road. All infrastructure items required for storing, loading and unloading of bulk explosive areassumed to be part of the contract mining provider. This includes a storage silo, powder magazine and

detonator magazine.

17.13 Bulk Fuel Storage and Delivery

Diesel fuel would be stored in an 85,000 L dual-wall fuel tanks located near the maintenance shop. The

tank would have an internal submersible pump capable of delivering 40 usgpm (2.5 L/s) to all site

vehicles. Diesel would be delivered to mobile equipment by the fuel and lube truck. An appropriate spill

containment pad would be installed around the fueling station.

17.14 Fresh/Firewater Tank and System

The fire water tank is deigned to be dual purpose serving as a freshwater and fire water storage tank.

Internal risers on all non-firewater suction lines would ensure a minimum volume of 680,000 L. This

capacity would allow for approximately two hours of firefighting capability.

The buried fire water pipe network would be pressurized by two pumps (one electric, one diesel

stand-by). This network would be connected to all buildings requiring fire protection.

17.15 Process Water Tank

The process water tank is design with a 2-hour retention time. It would be a 12.5 m x 12.5 m steel tank

with a total capacity of 1,500,000 L.

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17.16 Potable Water

A potable water treatment system would be installed near the process plant. It is designed to be

capable of producing potable water for 1,600 people at a rate of 125 liters/person/day. As part of the

sustainable mine plan, 20 km of buried potable water lines may be installed in an effort to supply

potable water to nearby villages. Further investigation pertaining to surrounding right-of-ways and

ground conditions would need to be completed to determine the validity of such a network.

17.17 Sewage Treatment

A sewage water treatment system would be constructed near the administration building. It would be

sized to handle approximately 200 personnel on site.

Contaminated water from the heavy equipment wash bay would pass through an oil water separator.Oily sludge would be stored in a transfer tank and back-hauled off-site for disposal. The remaining

water would be stored in a pond and recycled to the wash bay.

17.18 Water Treatment

Any surplus water generated by the mining operation is planned to be treated and tested prior to being

discharged into the environment. The type and capability of the water treatment plant will be determined

at the next phase of study.

17.19 Freight

Freight would be delivered to site on the all-season access road and offloaded at the warehouse orother designated area.

17.20 Tailings Storage Facilities

Design Criteria

Design criteria adopted to assess potential tailings storage facility (TSF) options, are listed below:

The new mine plan assumes that 120 Mt of tailings will report to the tailings storage facility over

a period of 13 years at an average production rate of 25,000t/d;

The assumed settled dry density of the tailings is 1.4 t/m3; this translates into the need to store

86 Mm3 of tailings;

The tailings beach slopes are assumed to be 1% for volume calculation purposes;

The project is in a high seismic area; therefore, dynamic stability is of importance; and

Environmental regulations require a fully lined facility for tailings storage.

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Rationale for Selected Design

The selected TSF is designed to incorporate a rock fill embankment. The embankment would be raised

in stages to allow deferral of CAPEX and the raises would employ the downstream construction

method.

An earth fill embankment was not considered since obtaining large quantities of select earth materials

typically poses a challenge and material availability has not been investigated. A compacted cyclone

sands embankment would require a liner system underneath the embankment based on current

regulations and was, therefore, not considered due to potential stability issues associated with having a

persistent weak layer under the embankment footprint.

Upstream and centerline construction techniques were not considered since the area experiences high

seismicity and neither technique is prudent for design in highly seismic areas.

It was assumed that the TSF would incorporate full geo-synthetic (HDPE or similar) lining.

Tailings Storage Design

Design Description

The TSF is designed to consist of two rock fill embankments with a fully-lined containment area. The

initial starter embankment would be constructed in the larger basin and as more storage capacity is

required, would expand into a smaller side basin, while the final embankment would extend across

these two areas. Seepage collection ponds are planned to be excavated downstream of the

embankments.

The embankments would have an upstream slope of 2.5:1 H:V and a downstream slope of 3:1 H:V. The

crest was assumed to be 10 m in width. The starter embankment would have a length of 560 m with

maximum height and width of 46 m and 270 m respectively. The ultimate embankments are designed to

have a combined crest length of 1,695 m, a maximum height of 97 m and maximum width of 550 m.

Each of the two embankments would have a stability key trench excavated to a depth of 5 m and width

of 50 m along the centerline of the ultimate embankment.

The upstream side of the embankment is planned to be lined with an HDPE liner installed on an

approximate 0.3-0.5 m bedding layer. Beneath the bedding layer would be a clay or transition material

followed by the bulk rock fill. The key trench would be backfilled with rock fill.

The facility is designed as a zero discharge facility until it is closed with a dry cover. The TSF design

includes a spillway to ensure protection of the embankment in the event of a flood. At the PFS stage

after a hazard classification has been undertaken, consideration could be given to designing the facility

to contain the probable maximum flood (PMF) which would negate the need for an operational spillway.

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HALILAĞA

Effective Date:

.

Figure 17.3:

PRO JECT –

December 20,

roposed T

PEA TECHNI

014

F Plan Vie

CAL REPO R

T

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Construction

The TSF facility is envisioned to be constructed in stages with the embankment being constructed in the

downstream direction. The starter embankment, seepage collection facilities and temporary spillway

would all be built prior to project start-up. Construction would be continuous throughout the life of mine

(LOM) thereafter, with a raise completed every year. The typical increase in height is planned to be

between 3 to 5 m. After each raise, the previous spillway would be backfilled and a new side hill

spillway constructed.

Construction material is planned to be sourced from a nearby quarry. All embankment fill would be

compacted in lifts to improve density and stiffness. Liner extensions would be tied into the existing

liners; small benches may be utilized to aid at liner tie-in points.

Clearing and grubbing are scheduled to be completed as needed during the expansion of the facility.The liner would be exposed at surface since new liner sections would be covered within a year or two of

placement and hydraulically placed tailings would not damage the liner system.

Closure

The facility would be re-sloped for positive drainage toward the spillway. A simple infiltration reducing

cover would be constructed that includes a sealing layer of 0.25 m, drainage layer of 0.5 m and top soil

of 0.5 m. The facility would then be re-vegetated and drainage paths will be lined with appropriately

sized riprap. Seepage collection ponds would remain during closure to monitor performance.

17.21 Water Storage

17.21.1 Design Criteria

The total processing water requirement is estimated to be approximately 750 m3 /hr. Of the total

required water, 80 m3 /hr of fresh water would be needed. This translates to roughly 6.5 Mm3 per year,

of which 0.7 Mm3 would be fresh water. It is estimated that 2.6 Mm3 could be recycled from the tailings

stream each year, which leaves the need for an additional 3.9 Mm3 of process make-up water.

-level hydrology analysis, which assumes average yearly values for precipitation and evaporation, with

a runoff coefficient of 0.3. An additional 2.2 Mm3 per year of run-off water can be captured from tailings

catchment area of roughly 20 km2. In order to achieve the total process water requirements, an

additional 1.7 Mm3 of water would be needed annually. In order to meet the additional water

requirement, a new catchment and storage area would need to be considered.

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17.22 Water Storage Design

Several nearby mined out water storage pits, have been identified. Since these pits already exist, they

would require the least amount of earthworks and permitting; therefore, they have been selected as the

water storage option for this PEA. No water storage embankments were considered, since the overall

impact on the project would be much greater if a separate water storage facility were to be constructed.

The water storage pits have a combined catchment of roughly 23.5 km2. The scoping level water

balance for these pits indicates that the 1.7 Mm3 of water can be achieved annually under average

conditions. Since runoff varies significantly from season to season, a volume of water would need to be

stored each year during the wet season to prepare for the dry summer months. Based on the scoping

level water balance, the annual storage requirement is estimated to be approximately 0.7 Mm3. This

does not account for any groundwater recharge into the pits.

Limited information is available for the water storage pits. The volumes have been estimated assuming

that the pits are sloped with an overall slope angle of 45o. Estimated pit volumes are presented below in

Table 17.1.

Table 17.1: Water Storage Pits Volume Estimates

Water Storage Location Estimated Storage Volume (m )

Pit 1 339,000

Pit 2 160,000

Pit 3 421,000

Pit 4 35,000

Pit 5 574,000

Total 1,529,000

Source: SRK 2014

The volume estimates indicate that only two of the larger pits would be needed in order to satisfy the

water storage requirements. Pit 1 is optimally located near the bottom of the catchment and is ideal for

capturing water; therefore, Pit 1 would be required. Of the two remaining larger pits (Pits 3 and 5), Pit 3

has been selected since it is in line with the process plant and Pit 1 has reduced pumping requirements

as compared to Pit 5, which is at a higher elevation. The water collected in Pit 1 would be pumped to

Pit 3 and from Pit 3 to the process plant as needed.

In order to increase confidence that the pits would yield the appropriate volume of water annually, a

small diversion with sluice gate is planned to be constructed to redirect water into Pit 1 from the nearby

natural flow path approximately 100 m southeast of the pit. The design of the diversion would be a

simple channel, 3 m in width across the bottom with 2:1 H:V side slopes. The channel would be lined

with appropriately sized riprap and have a concrete sluice gate to control the inflow of water.

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17.22.1 Estimated Quantities for Tailings and Water Facilities

Primary material volumes were calculated using AutoCAD Civil 3D (2014) and checked manually.Where quantity estimates were made using AutoCAD, additional checks have been completed using

Global Mapper 15.1 (2014) as well as hand calculations. The summary of quantities for the tailings and

water storage facilities are presented in Table 17.2.

Table 17.2: Scoping Level Cost Estimate Quantities

Tailings Storage Facility Unit Estimate

Bedding Material m3 837,900

Transition Material m3 400,200

Run of Quarry Rock m3 12,920,300

Excavations m3

1,262,500Liner m

2 2,985,200

Clear and Grub m2 3,084,900

Deposition Pipe m 7,200

Water Storage Facility

Diversion Excavation m3 2,700

Rip Rap m3 500

Concrete m3 20

Source: SRK 2014

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HALILAĞA

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Figure 17.4:

PRO JECT –

: December 20,

Typical Sec

PEA TECHNI

2014

ion: Tailing

CAL REPO R

Embankm

T

nt

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17.23 Rock Storage Facilities

The Rock Storage Facility is proposed to be located immediately adjacent to the final pit limit for the

Halilağa deposit (see Figure 17.1). Given the deposit configuration and extraction sequence, no

backfilling into previously mined out areas is currently planned for Halilağa.

It is envisioned that the RSF would be built in a series of lifts in a “bottom-up” approach in order to

maximize stability. The RSF would be constructed by placing material at its natural angle of repose

(approximately 1.5H:1V) with safety berms spaced at regular intervals giving an overall operational

slope of 2.5:1.

The east RSF waste tonnage capacity is expected to reach 160 Mt.

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18 MARKET STUDIES AND CONTRACTS18.1 Market Studies

At this time, no market studies have been completed. No contractual arrangements for concentrate

trucking, port usage, shipping, smelting or refining exist at this time. There are no contracts in place for

the sale of copper concentrate or gold doré. It is assumed that the concentrate produced at the Halilağa

mine would be marketed to international smelters in Asia and Europe. No deleterious elements have

identified or considered at this time.

The smelter terms used in the economic analysis are based on recent marketing terms from similar

projects.

Table 18.1: NSR Parameters used in the Economic Analysis

NSR Parameter Unit Value

Cu Concentrate % 30

Cu Payable % 96

Cu Treatment Charge US$/dmt conc 92.00

Cu Refining Charge US$/lb Cu 0.092

Cu Min. Deduction % 0.0

Moisture Content US$/wmt conc 8.0

Truck Freight to Port US$/wmt conc 20.00

Port Charges US$/wmt conc 6.69Marine Transportation Insurance US$/wmt conc 1.00

Ocean Freight US$/wmt conc 30.00

Total Offsite Costs US$/wmt conc 57.69

US$/dmt conc 62.71

Au in Cu Concentrate

Au Payable % 96

Au Refining Charge US$/oz 7.00

Au Min. Deduction g/t in conc 1

Au in Doré

Au Payable in Doré % 99Au Refining Charge US$/oz 7.00

Source: JDS 2014

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HALILAĞA

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HALILAĞA

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Source: JDS

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Table 18.2

Metal

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Source: JDS

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19 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR

COMMUNITY IMPACT

The information presented in this section is provided by SRK Turkey.

19.1 Environmental Baseline Study

An Environmental Baseline Study (EBS) was conducted by. SRK Turkey in 2011-2012 in the Halilağa

exploration program surroundings (SRK, 2012). Due to the early stages of the program, project

information was not sufficient to conduct a comprehensive EBS. Therefore, a limited review of

environmental resources was conducted to obtain a preliminary insight about the existing conditions.

The limitations of the review related to the geographical and temporal extent, and technical detail of the

studies. This early EBS covered the following areas:

Protected areas;

Land use and soils;

Hydrology;

Hydrogeology;

Community water resources;

Water quality;

Geochemical assessment of waste lithologies; and

Biodiversity.

A summary of findings for each area is provided in the following sections.

19.1.1 Protected Areas

In the Turkish regulatory system, there are certain protected areas and land use types where

mining/industrial project development is restricted; while in other certain areas there may be limitations

rather than restrictions. It is important to identify these as early as possible in the project development

to prevent any compliance and permitting related issues. The protected areas around the project site

are shown in Figure 19.1. There are seven protected areas identified in the area. The ones that are

noteworthy are:

Çanakkale-Çan Etili Tepeköy thermal tourism center at 2.5 km distance to the TSF;

Bayramiç potable water reservoir long-range protection area at 2.5 km distance to the TSF;

Kazdağı fir tree nature protection area at 15 km distance to the TSF; and

Kazdağı Nature Park at 16 km distance to the TSF.

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The current footprint of the Halilağa project is outside of these protected areas, and therefore, these

areas are only environmental sensitivities for the Halilağa project at this stage.

Land Use and Soils

The current land use is presented in Figure 19.2. The environmental permits and studies required may

vary based on the land use type of where the project is located. Most of the project footprint is on forest

and scrub land, made up of mainly Turkish pine (Pinus brutia), Black pine (Pinus Nigra), and Oak tree

(Quercus). The forest and scrub lands are under the ownership of Ministry of Forestry and Water

Management (MFWM). MFWM issues the relevant land permits for mine operations on forest lands

and require Forest Rehabilitation Plans as part of their mine closure requirements.

The remaining parts of the project area are on agricultural land. Agricultural lands are likely to be

privately owned and may require Soil Conservation Plans for mine closure and rehabilitation. The LandUse Capability Classes (LUCC) observed in the project area are shown in Figure 19.3. The majority of

the area has LUCC of VI and VII, which are generally unsuitable for cultivation due to steep slopes,

shallow soil layers, or stony, sodic, or salty soil texture. A relatively small area to the northwest of the

open pit has LUCC of III, which is suitable for cultivation and is already utilized by the local

communities.

The soil type in the project area is “Brown Forest Soil without Lime”. The thickness of the top soil in the

area varies from very shallow (0-20 cm) to medium deep (50-90 cm). The medium deep top-soil areas

coincide with the agricultural areas in the northwest, whereas the shallow soils coincide with the

elevated forest areas. In general, the soil erosion in the area is categorized as very severe (4th degree)

in the forested areas to medium severity (2nd

degree) in the agricultural areas. Three soil samples werecollected and analyzed as part of the EBS. All the samples demonstrated elevated levels of Arsenic

(As), Lead (Pb), Selenium (Se), and Antimony (Sb) as compared to typical earth crustal abundances.

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19.1.2 Hydrology

The project area falls on the topographical divide between the two major catchment areas, namelyNorth Aegean catchment and Marmara catchment as shown in Figure 19.4. Karamenderes River and

tributaries drain the North Aegean catchment area; whereas, Kocabaş River and tributaries drain the

Marmara catchment area. There are several sub-catchments in the project area that drain in different

directions (Figure 19.5). The TSF is situated on the northeastern surface drainage system connected to

the Marmara catchment. Whereas, the open pit area is situated on the southwestern surface drainage

system connected to the North Aegean catchment. The area covered by each sub-catchment is given

in Table 19.1.

Table 19.1: Sub-catchment Areas

Sub-catchment Area (ha)

A 29,510

B 755

C 485

D 537

E 475

F 124

G 285

H 2,145

I 186

Source: SRK 2014

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State Hydraulics Works (SHW) operates a surface water flow monitoring station on Koca Stream close

the town of Çan, to the northeast of the project area. Koca Stream is the largest water course in thearea and drains part of the project area. It has a catchment area of 164 km2 and a highly variable

discharge with low-flows of 0.05-0.08 m3 /s in summers and peak-flows of 2-3 m3 /s in winters. Based on

the hydrograph analysis conducted using 1990 Koca Station data, it is estimated that 64% of its base

flow is due to groundwater contribution. A water balance conducted for Koca station’s catchment area

also indicated that 59% of the annual precipitation is lost to evapotranspiration, 25% becomes surface

run-off, and 16% is recharge to groundwater.

A continuous flow measurement was also conducted by SRK by means of pressure transducer on the

downstream of Kapaklı creek between HASW-9 and HASW17 monitoring points (Figure 19.5). It was a

short term measurement covering the period between April 15 – 30 May, 2012. The average flow rates

varied between 1 – 75 L/s.

19.1.3 Hydrogeology

The hydrogeology of the project area was studied through monitoring of springs, developed springs,

water capture points (depots), hand-dug wells, and production wells. A total 183 groundwater points

was identified and recorded (Figure 19.6). Flow rates and field parameters were measured at selected

groundwater points for four seasons. Furthermore, three groundwater wells were installed and tested at

the open pit area as part of the EBS. Calibration test, multiple-rate step-drawdown test, and constant

discharge test were conducted to determine the hydrogeological parameters in the open pit area.

Hydraulic conductivity as well as specific storage was determined for the three groundwater wells.

These are given in Table 19.2. The hydrogeological units can be defined as low-permeable, and

impermeable where higher storage properties can be observed due to presence of clay alteration.

Table 19.2: Hydrogeological Parameters for Groundwater Wells

Monitoring Well Hydraulic Conductivity (m/s) Specific Storage

HAMW-1 1.28 x 10-8

2.44 x 10-6

HAMW-2 1.40 x 10-8

1.20 x 10-3

HAMW-3 1.67 x 10-7

1.00 x 10-3

Source: SRK 2014

Groundwater levels were measured to identify the response of water table to different hydrologicalconditions such as precipitation, evaporation, and resource depletion. Hydraulic gradient, groundwater

flow direction, and depth to groundwater were determined in the open pit area. Depth to groundwater is

shown in Figure 19.7. Groundwater elevations are shown in Figure 19.8. The depth to groundwater

varies between 10 m to 60 m in the open pit area. The larger depths are observed in the southeast

where the topographical elevations are higher.

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19.1.4 Community Water Resources

The project area has scarce water. All the villages obtain their water from springs which are routed towater storage depots for final distribution to the village fountains or networks. The community water

distribution is always by gravity through small pipelines, which keeps the capital and operational costs

minimal for the communities. Some small hand-dug wells are used for small agricultural irrigation as

well. During dry periods the water shortage becomes even more critical for the villages. A hydro-

census conducted identified the water abstraction points, pipelines, and storage depots for the nearby

communities (SRK, 2011). These are shown in Figure 19.9. It should be noted that water supply

pipeline locations and routes are only estimates as there are no existing documentation showing exact

routes and these have not been traced in the field.

As seen in Figure 19.9, all the water abstraction points, water storage depots, and pipelines of the

Muratlar village are within the open pit footprint. These will have to be replaced with the open pit beingdeveloped. Currently, the water supply to the Muratlar village is at a low rate of less than 1 L/s and is

deemed insufficient for the Muratlar village. Hacıbekirler Village obtains its water from springs located

to the south of the open pit and southwest of the Rock Storage Facility. These have the potential of

being impacted in terms of quantity and quality by the cone of drawdown created by the open pit and

potential contamination by the RSF.

The water abstraction points for the Halilağa, Keçiağılı, and Hacıkasım villages are located upstream of

both the TSF and RSF, and are not likely to have potential impacts by these project units. However,

their pipeline and water depots are located between the RSF and TSF, and would need to be carefully

monitored.

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19.1.5 Water Quality

Quarterly water quality monitoring was conducted at selected surface and groundwater locations in the

project area between July 2011 and March 2012. Monitoring was conducted at nine surface water

locations, four groundwater wells, five developed springs, and three water storage depots. Several

surface water locations were observed to be dry in July and September sampling campaigns. The

water samples were analyzed for several parameters including dissolved ions, nutrients, dissolved and

total metals, and several other physical and chemical parameters. The analysis results were compared

with the Turkish and international water quality limits for different uses, including Turkish inland water

quality limits, and WHO, EU, and Ministry of Health (MoH) drinking water limits. A summary of the

findings follows:

Some of the surface water locations exhibit elevated Aluminum (Al), Copper (Cu), and Lead

(Pb) levels;

Nearly all surface water locations were identified as unsuitable for drinking purposes due to

elevated Al, Pb, Manganese (Mn), and iron (Fe) levels;

Some of the groundwater monitoring wells exhibit elevated Arsenic (As), Zinc (Zn), Mn, and

sulfate levels;

Two groundwater wells (HAMW1 and HAMW2) have elevated levels of Na and Cl, which can be

indicative of discharge of either geothermal systems or deep groundwater. Proximity of Etili

Tepeköy thermal center might be a contributing factor;

Developed springs exhibit elevated Pb, Al, Cu, and Fe levels and have pH outside of the normal

limits;

The water depots where water is supplied to the villages all had elevated levels of Al. Some had

elevated Pb, Fe, and Cu levels. pH levels were also found outside of the drinking water limits.

19.1.6 Geochemical Assessment of Waste Lithologies

Screening level geochemical testing of waste lithologies were conducted on 37 samples representing

12 lithologies and eight alteration types found in the Halilağa core samples. The lithologies tested

include the main units such as felsic intermediate intrusives, metamorphics, intermediate volcanics,

volcanic breccia, and fault zone, as well as other minor lithologies such as diorite, diabase,

hydrothermal breccia, clay, dyke, skarn, and siliciclastics.

Static testing on the samples was conducted to determine the:

Potential for producing acid leachates;

Content of significantly elevated elements; and

Potential to release the identified components with elevated concentrations upon leaching.

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The analyses conducted within the static testing program are given in Table 19.3.

Table 19.3: Analyses Conducted in the Static Testing Program

The Main Purpose of Testing Testing Type

Acid Generation/Neutralization Potential

Paste pH (Sobek, 1978)

Neutralization Potential (Modified Sobek Method)

Complete C and S Speciation (Total C, Total Inorganic

C, and Total S analysis by Leco furnace (ASTM

E1915-97), SO4-S (ASTM D2492-02))

Static NAG test (EGI - Environmental GeochemistryInternational, 1986)

Metals and readily available contaminants

Short Term Leach Test (Modified US EPA 1312

Method (3:1, L:S ratio)) followed by ICPMS analysis of

leachate (pH, EC, TDS, alkalinity, cations, anions,

metals and metalloids)

Whole Rock Assay by lithium borate fusion and Trace

Element Analysis with aqua regia digest and ICPMS

finish

Mineralogy Standard XRD method for identification of major

minerals

Source: SRK 2014

The following are the main findings from the Acid Base Accounting (ABA) and Net Acid Generation

(NAG) tests:

Sulphide zone Diorite, Sulphide zone Metasediment and Massive Pyrite are classified as

Potentially Acid Forming (PAF). In addition, short term leach of massive pyrite and a sulphide

zone diorite sample created elevated SO4 levels combined with low pH. This is interpreted as

the indication of the presence of readily available acidity in samples that is formed by the acidicsulfate salts that could cause short-term or long-term water quality issues.

Marble in oxide, sulphide, and transition zones are classified as Non Acid Forming (NAF).

Marble is generally with very high Neutralization Potential (NP). Marble within transition or

sulphide zones might contain elevated sulfur concentrations. This might result in neutral but

saline drainages.

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Oxide and transition zone Diorite and sulfidic Fault Zone lithologies are classified as having

uncertain acid forming potential. Gossan is also classified in uncertain zone with respect to acid forming potential. The NP of the

gossan might be composed of slow reacting minerals. Therefore, the dynamic acid drainage

generation might be possible for some parts of this lithology where the slow reacting buffering

capacity is combined with relatively higher sulfur contents.

Oxide and transition zone Metasediment has low NP. The lithology is classified in uncertain

zone or PAF zone for different samples. Oxide/transition zone metasediment acid generation is

classified as likely considering that local mineralization with relatively higher sulfur content might

create acidic drainages in time, particularly after the consumption of the limited carbonate NP.

The lithology also estimated to contain some easily dissolving acidic salts which might be

released in the case of contact with water.

Potentially Problematic Element (PPE) and metal leaching assessment indicated:

Au, Ag, As, Bi, Ca, Cd, Cu, Fe, Hg, In, Mn, Mo, Pb, Re, S, Sb, Se, Sn, Th, Tl, U, W and Zn

occur at greater than 3 times the crustal abundance and can therefore be regarded as PPE. In

all lithological groups Au, Ag, As, Bi, Cu, Mo, Re, S, Sb, and Se occur at high concentrations.

Massive pyrite and oxide zone diorite contain the highest number of PPE among all lithologies.

Sulphide zone diorite, metasediment and massive pyrite can be classified as having the highest

metal mobility potential in the long term. The major portion of the PPE content in these

lithologies is estimated to be reactive and therefore may contribute to the total metal leach

potential. Total Ficklin metals (copper + cadmium + cobalt + lead + nickel + zinc) leach from marble,

gossan, oxide zone diorite, and fault zone samples is estimated to be generally low. Elevated

concentrations for Ficklin metals are estimated particularly for sulphide zone diorite, sulphide

zone metasediment, and massive pyrite samples.

The dynamic leaching behavior metals should be monitored for these lithologies since there

might be the risk of metal laden drainage exceeding the water quality standards.

19.1.7 Biodiversity

The biodiversity in the project area was studied through literature review as well as a field survey. One

single field survey was conducted in July 2011, mainly in the vicinity of the open pit area.

Dense vegetation is observed at the project site. The site is characterized by forests of Turkish pine

trees (pinus brutia). There are also perennial tree forms such as juniper (juniperus oxycedrus),

Phillyrea laitofolia (Akçakesme in Turkish), and two species of oak (Quercus infectoria and Quercus

pubescens). Vegetal formation under trees is not well developed due to the dense tree cover.

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Forms of bushes and plants having a lifetime of one year are encountered by the road sides or on the

clearings (including the exploration area).

Stony or rocky habitats as well as bare areas are limited within the study area. From this perspective,

the site seems to have a uniform habitat and therefore considered as poor in terms of habitat diversity.

A total 51 plant species, 56 types, six sub-types and two varieties belonging to 25 families have been

identified in the field. Further field surveys in the future extending beyond a single season and a larger

area might yield further species. All the species identified are in the category of Least Concern (LC)

(i.e. exposed to minor threat) in terms of IUCN risk category.

The faunal aspects of the project area were studied through literature reviews, field survey, and

interviews with the locals. The project area is not rich in terms of habitat types which results in limited

nutrition, resting, overnight stays, nesting, and reproduction opportunities affecting the wild lifecomponent. A summary of faunal findings follows:

Only one amphibian species was identified due to limited aquatic habitat. Pelophylax ridibundus

(Water frog) is a cosmopolitan species and has a LC risk category;

Nine species of reptilians are identified in the project area. Similar to amphibians, reptilians also

have limited habitat in the project area. With the exception of one species all reptilians have the

LC risk category. Testudo graeca (ordinary turtle) is listed as Vulnerable (VU) under the

European Red List criteria. Four of the identified species are protected under Annex II of the

Bern Convention (as Species under Absolute Control), while the remaining five species are

protected under Annex III of the Convention (as Species under Control).

Fifty-three bird species were identified in the area. Thirty-six of the species are protected under

Bern Convention Annex II, while 12 of the species are protected under Annex III of the

convention. Coracias garrulus (blue crow) is considered Near Threatened (NT) under the

European Red List. The remaining species have LC rick category. No endemic species were

found in the region.

For the mammal identification facea, footprint, nest entrances, and carcasses were surveyed

and interviews with locals were conducted. Thirteen species were identified including brown

bear (Urcus arctos). Four of the species are under Annex II of the Bern Convention, while five

of them are under Annex III. No endemic species were identified in the project area.

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19.2 Mine Closure Requirements

19.2.1 Legal Requirements

There are several new regulations in Turkey that are being prepared which may change the legal

requirements in the future. However, these are not yet at an effective stage.

The relevant Turkish laws and regulations are described below:

Mining Law Nr. 3213: Article 32 of the Law defines the responsibilities of the license owner in

the event of the mineral license termination. The article stipulates that the license owner is

obligated to take every measure and precaution for the public safety and environmental

protection. The license owner is required to implement these within six months of the license

termination. This period can be extended by only three months, if the absence of any force-

majeur conditions. Should the license owner not implement the measures within the allotted

timeframe, the relevant governmental agencies (forestry departments in forestry areas,

provincial departments in the remaining areas) take over and implement the necessary actions.

However, the legal and financial liability still remains with the license owner. Any financial

expense in excess of the environmental fees already paid is collected from the license owner.

Forestry Law Nr. 6831: Article 17 of the Law regulates the mining activities within forestry areas.

Implementation regulation for Article 17 (Official Gazette 28976, Date: 18.04.2014) defines the

forestry rehabilitation plan requirements and fees.

Regulation for Reclamation of Lands Disturbed by Mining Activities (Official Gazette 27471,

Date: 23.01.2010): This regulation was first enforced in 2007 and then amended in 2010. It is

applicable to all mining projects. In accordance with the regulation, a Mine Reclamation andClosure Plan (MCRP) needs to be prepared for all mining projects. Generally, the MCRP is

prepared during the Environmental Impact Assessment (EIA) permitting process and is

submitted within the appendices of the EIA report. Since it is prepared during the EIA phase,

the nature of the mine closure and rehabilitation plan is “conceptual” and it does not contain any

closure cost estimates. For mining projects on forestry lands, instead of an MCRP a forestry

rehabilitation plan is requested. The competent authority for the forestry rehabilitation plans is

the MFWM. Whereas, the competent authority for MCRPs is the Ministry of Environment and

Urban Planning (MEUP). For mining projects located on absolute agricultural areas, a soil

protection and rehabilitation plan is also required. The competent authority for agricultural and

pasture lands is the Ministry of Food, Agriculture and Livestock (MFAL).

Landfill Regulation (Official Gazette 27533, Date: 26.03.2010): The landfill regulationprescribes the method of disposing solid wastes (domestic and industrial). While the regulation

recognizes mining waste as a “special” waste category, due to the lack of dedicated mining

waste regulation all “processed” mining wastes (i.e. tailings) are currently managed under this

regulation. Therefore, the disposal methods applicable to domestic/industrial wastes are also

applicable to mine tailings. The waste rock dumps are not included under this directive. Article

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16 of the regulation defines the bottom sealing requirements for the waste storage facilities.

Whereas Article 17 of the regulation defines the top sealing requirements at closure. Thesealing requirements vary based on the waste character.

The wastes are characterized into three categories as inert, non-hazardous, and hazardous.

The bottom sealing requirements are:

o Class-I (Hazardous waste): Impermeability of 10-9 m/s with an equivalent thickness of 5 m or

more;

o Class-II (Non-hazardous waste): Impermeability of 10-9 m/s with an equivalent thickness of

1 m or more; and

o Class-III (Inert waste): Impermeability of 10-8 m/s with an equivalent thickness of 1 m or

more.

A clay layer of 50 cm minimum thickness, and where necessary geo-synthetics, are required toachieve the bottom sealing requirements.

At closure, the Class-I wastes are required to be sealed off with geo-synthetics, 50 cm of clay, and

50 cm of top-soil. For Class-II waste geo-synthetics are not required. For Class-III wastes geo-

synthetics and clay layers are not required.

The waste characterization is done based on a CEN leach test similar to TCLP of U.S. EPA. Waste

characterization criteria are established for 22 parameters (including several heavy metals).

Exceedance of the upper limit for any parameter is sufficient to classify the waste in the stricter

waste category.

The following European Union (EU) directives are either being currently drafted or planned to betransposed into Turkish regulations in the future:

Directive on Management of Waste from Extractive Industries (2006/21/EC): This directive was

promulgated in the EU in 2006. It is currently in draft form in Turkey and scheduled for

implementation in the near future. The Directive classifies the TSFs into Category-A and non-

Category-A facilities, and based on this risk categorization require different levels of measures

for construction, operation, closure, and post-closure. The Directive also introduces the concept

of financial guarantees for the closure and post-closure phases.

Directive on Environmental Liability with Regard to the Prevention and Remedying of

Environmental Damage (2004/35/EC): This directive is included in the environmental plans of

the EU acquisition process, but is not yet studied in Turkey. It establishes “the polluter pays”principle and a liability period of up to 30 years.

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Reference Document on Best Available Techniques for the Management of Tailings and Waste

Rock in Mining Activities: This is a document that is used for determining the Best AvailableTechniques (BAT) for the Integrated Pollution Prevention and Control (IPPC) permitting

process. It is informally used in Turkey to seek opinion on technical matters. The reference

document provides the best industry practices observed throughout the European region.

There is one best industrial practice example for a gold mine in Turkey as well.

19.2.2 International Perspective

International perspectives on good practice with respect to mine rehabilitation and closure are

presented in numerous documents, with some of the most important being:

The Planning for Integrated Mine Closure Toolkit, International Council on Mining and Metals

(ICCM) Good Practice Website (2009); Environmental Health and Safety (EHS) Guidelines for Mining, the IFC and World Bank (IFC,

2007);

Mine Rehabilitation: Leading Practice Sustainable Development Program for the Mining Industry

(Commonwealth of Australia, 2006); and

Mining for Closure: Policies, Practices and Guidelines for Sustainable Mining Practices and

Closure of mines (United Nations Environment Program, 2005).

19.2.3 Halilağa Mine Closure and Rehabilitation

At this stage of the project development, a conceptual mine closure and rehabilitation with the following

features is anticipated:

Rock Storage Facility (RSF) will be re-graded to achieve 3H:1V slopes for geotechnical stability.

Given the preliminary geochemical test results, it is likely that the RSF will possess Acid Rock

Drainage (ARD) / Metal Leaching (ML) characteristics which will require special management at

closure. This is usually achieved through a mineral top layer (clay) overlain by topsoil to

promote growth of vegetation. A passive treatment cell might be required at the toe of the dump

to collect and treat any ARD/ML seepage. A forestation fee will be paid to the MFWM during

the operations to account for the re-forestation costs.

Tailings storage facility (TSF) will be closed in accordance with the mandatory specifications.

Experience has shown that tailings of these processes are either Class I or Class II. Given the

sensitivity of the region, it is assumed here that the Halilağa tailings will be of Class I. This willrequire that the top seal at closure is made up of geosynthetics, 50 cm clay, and 50 cm soil. A

forestation fee will be paid to the MFWM before the closure to account for the re-forestation

costs.

The open pit will be-lined with Non-Acid Forming (NAF) waste rock at closure. If there is acid pit

lake forming in the post-closure phase, this may need to be mitigated further. However, without

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doing further geochemical and hydrogeological analysis it is not possible to say whether pit lake

will form and subsequently impact the environment. Administrative and process buildings and other utilities will be dismantled and demolished at

closure. The equipment will be either sold as scrap or second-hand equipment. All equipment

will be decontaminated before dismantling. The demolishing waste will be either buried in-place

or disposed of in the open pit or RSF. The exposed areas will be re-vegetated and converted to

previous land use.

19.3 Environmental Permitting

The mineral licensing system in Turkey is linked to the environmental permitting system. Once an

exploration license is converted into an operation license, the license owner is required to obtain allenvironmental and other permits within three years of the operation license. Once these permits are

obtained, the owner is granted a Mining Operation Permit (MOP). Under the current mining law, failure

to obtain the MOP within three years results in the termination and the loss of the license. The Halilağa

prospect currently comprises 14 mineral licenses. Five of these are operational licenses, two are

exploration licenses (AR), and the remaining seven are in the process of being converted from

exploration to operational licenses. Therefore, the schedule for the five operational licenses is

important for this study. All the Halilağa mineral licenses are shown in Figure 18.10. A list of the five

Halilağa operation licenses and their conditions are given in Table 18.4; the MOP process is still on-

going at the time of this PEA, except for two of the operating licenses. At this stage, it is sufficient to

obtain the MOP for quartz and feldspar at all the Halilağa operation licenses since both quartz and

feldspar are Group iV, as are gold and copper. Minerals under the same license groups are owned by

a single owner. However, at the time of the Halilağa copper and gold project development, the MOP

(hence the environmental permits) will need to be amended for the new project design involving copper

and gold.

Table 19.4: Halilağa Operation Licenses

License Nr. Start Date Expiry Date MOP Notes

81802 21/05/2009 21/05/2019 Yes Obtained for Quartz

82361 6/12/2012 6/12/2022 NoIn the process of getting

business and land permits forfeldspar

51297 7/10/2013 7/10/2023 Yes Obtained for Quartz

20054260 2/3/2012 2/3/2015 No Applied for Quartz

20064172 21/03/2012 21/03/2015 No Applied for Quartz

Source: SRK 2014

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The first step in the Turkish environmental permitting system is the Environmental Impact Assessment

(EIA) permit. It is required for all major mining projects and its ancillary facilities. Turkish EIA regulationwas transposed from European Union (EU) EIA Directive and is similar in structure. Furthermore,

several other regulations and limits, such as air quality, noise, water etc. were also transposed from the

EU directives and are therefore in agreement with the EU requirements. However, the Turkish EIA

does not have a strong Stakeholder Engagement (SE) process as well as a Social Impact Assessment

(SIA) component. In that regard, Turkish EIAs fall short of meeting the Equator Principles and the

World Bank / IFC Environmental and Social Performance Standards (ESPS) which are the international

benchmarks for environmental performance. Additional studies are needed to augment and

complement the Turkish EIA permitting studies. For the Halilağa project described in this PEA, a

comprehensive Environmental and Social Impact Assessment (ESIA) will be conducted that meets both

the requirements of the Turkish regulatory system and the Good International Industry Practice (GIIP)

as described by the IFC ESPS.

The flowchart for the Turkish EIA permitting process with typical timelines is shown in Figure 19.11. It

should be noted that while certain tasks have specific times indicated in the regulation, these are not

necessarily binding for the governmental agencies.

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Figure 19.11: EIA Permitting Process Flowchart

Source: SRK 2014

The Turkish EIA permitting process can be considered in two parts. The first part starts with the

submission of an EIA application document (also referred to as “Project Description Document – PDD”)

to the Ministry of Environment and Urban Planning (MEUP).

Once the submission is made, the PDD is checked by MEUP for compliance with the mandatory report

format and the EIA regulation. If the PDD is found suitable, then it is distributed to the Review and

Assessment Committee (RAC) (made up of representatives of various governmental agencies) and the

date of the official Public Hearing is determined. The date and place of the hearing is then announced

at least in one local and one national newspaper (ten days in advance of the hearing). Following the

public hearing, the RAC members upload their comments about the project to the electronic permitting

system. Then the mandatory Terms of Reference (ToR) for the EIA report format is provided by MEUP

following the payment of the EIA permitting fee by the Project owner (i.e. Client).

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The permitting fee varies according to the estimated capital investment cost required for the project.

Once the mandatory ToR is received from the MEUP, the project owner and its consultant is allowed amaximum of 18 months to complete the draft EIA report ready for RAC’s review.

The second part of the EIA permitting process starts with submission of a draft EIA report to MEUP.

The draft EIA report is checked by MEUP for compliance with the ToR and the regulation. The report is

reproduced and distributed to RAC and the RAC meeting date is decided. At this stage the draft EIA

report is open to written public comment. At the meeting, RAC can request major revision to the EIA

report, if it deems the draft EIA report does not meet its technical or administrative requirements. At

most two revisions can be requested. The permitting process is suspended until the client and its

consultants complete the required revisions. If no or minor revisions are requested, the draft EIA report

is approved in the first RAC meeting. Following completion of minor revisions (if any), the final EIA

report is open to the public review for ten days. Based on the RAC’s decision and consideration of thepublic comments, the EIA permitting process is given either a “Positive” or “Negative” decision (i.e. the

impacts are found acceptable or not). The EIA permitting process is finalized by the Minister’s

signature.

Excluding the times for the EIA application document and draft EIA report preparation, the EIA

permitting process typically takes about:

two to three months for the first part; and

three to five months for the second part excluding the Minister’s Signature.

The time to conduct various technical environmental and social assessments needed to support the EIApermit would take additional time. Based on past experiences, an ESIA study for mine project of similar

size takes not less than three years, including the permitting procedures. However, since the ESIA

process progresses in parallel to the Feasibility Studies (FS), the overall timeframe also depends on the

progress of the FS. Typically the EIA Application is done at the time of pre-feasibility study (PFS)

conclusion. The draft EIA report is generally completed sometime after the FS conclusion or when the

project design freeze is achieved.

The individual studies that are required for the Halilağa ESIA study during the project development

phase are given in Table 19.6 together with their typical costs (including laboratory, drilling, permitting

and other relevant expenses).

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Table 19.5: Studies to be Conducted under the ESIA Framework

StudyCosts(US$)

Waste Geochemical Assessment 500,000

Water Impact Assessment 900,000

Stakeholder Engagement & Social Assessment 350,000

Air Quality Assessment 70,000

Biodiversity Assessment 70,000

Visibility Assessment 30,000

Noise Assessment 30,000

Environmental Management Plans 30,000

Mine Closure and Rehabilitation Plan 70,000

Overall EIA Report & Permitting 350,000

Total (excluding taxes) 2,400,000

Source: SRK 2014

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20 CAPITAL AND OPERATING COSTS

20.1 Capital Cost Estimate

The capital cost (CAPEX) of the project has been estimated based on the scope defined in previous

sections of this report. The following parties have contributed to the preparation of the CAPEX estimate

in specific areas:

JDS:

Process plant;

Plant infrastructure and services, including road diversion, high-voltage transmission

line, and high-voltage substation;

EPCM costs relating to the process plant and infrastructure outlined above;

Owner’s Costs;

Mine working capital;

Mine haul road; and

Initial mine development costs.

SRK:

Water reservoir;

Tailings storage facility;

Mine and infrastructure geotechnical engineering; and

Environment, social considerations and permitting.

CH Consultants:

Unit rates for concrete, steel, platework, mechanical and structural installation.

20.1.1 Capital Cost Summary

The estimated costs include mine pre-stripping, mine development, site preparation, process plant, first

fills, infrastructure, buildings, utilities and road works. The estimates are considered to have an overall

accuracy of ±30% and assume the project would be developed on an EPCM basis.

Table 20.1 provides a summary of the CAPEX estimate and Figures 20.1 and 20.2 depict Pre-

production and sustaining CAPEX breakdown.

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Table 20.1: CAPEX Estimate

Capital Cost Pre-Production(US$M)

Sustaining/Closure(US$M)

Total Capital Costs(US$M)

Capitalized Mining Costs 17.9 0.0 17.9

Contractor Mobilization/Demobilization 1.0 1.0 2.0

Mining 0.6 0.0 0.6

Site Development 5.5 0.0 5.5

Process Plant 131.6 0.0 131.6

On-Site Infrastructure 29.6 0.0 29.6

Tailings Storage 25.0 103.3 128.3

Indirects 37.6 0.0 37.6

EPCM 25.3 0.0 25.3Owner's Costs 6.4 0.0 6.4

Sustaining 0.0 15.8 15.8

Closure 0.0 50.2 50.2

Subtotal 280.6 170.3 450.8

Contingency 65.4 42.3 107.7

Total Capital Cost 346.0 212.6 558.5

Source: JDS 2014

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The following parameters and qualifications are made:

VAT (Value Added Tax) assumed to be recoverable;

No escalation;

Estimate was based on Q4 2014 prices and costs; and

No allowance has been made for exchange rate fluctuations over the life of the mine.

Data for these estimates have been obtained from numerous sources, including:

PEA-level engineering design;

Unit rates obtained from local Turkish mining contractors;

Budgetary equipment quotations; QP experience; and

Data from recently completed similar studies and projects.

The following assumptions were used in the CAPEX estimate:

The detail of the design is discussed in the relevant sections of this report;

Mining costs are based on contractor all-inclusive rates and carried in the OPEX;

Benchmarked plant CAPEX estimates are accurate;

Suitably qualified and experienced construction labour would be available at the time of

execution of the project; No allowance for construction camp; construction personnel to be housed in local community;

No exceptional geotechnical and drainage issues during, therefore, no allowance for special

ground preparation was made;

Borrow sources for construction are available from within the mine limits;

A power and water supply capable of supplying the required demand of the processing plant is

assumed to be available;

No extremes in weather would be experienced during the construction phase and as such, no

allowances are included for construction-labour stand-down costs; and

Costs sourced in Turkish Lira (TL) were converted at a rate of 0.44USD:1.00TL. This rate is in

line with the 3–month trailing average exchange rate as at December 2014.

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The following items are excluded from the estimate:

Cost changes due to currency fluctuation;

Force majeure issues;

Sunk costs up to the project go-no go decision point. The costs that are excluded encompass

pre-feasibility study costs, feasibility study costs, resource definition drilling, EIA work,

metallurgical testing, hydrogeological and geotechnical drilling and test work and all other work

associated with a feasibility study and EIA ;

VAT (assumed to be recoverable) - the delay in the recovery has been assumed to be negligible

for PEA purposes;

Future scope changes;

Project insurances;

Project interest and financing cost;

Land acquisition and compensation cost;

Operational insurances such as business interruption insurance and machinery breakdown;

Pebble crushing (will be carried out by local contractor);

Public road maintenance; and

Relocation or preservation costs, delays and redesign work associated with any antiquities and

sacred sites.

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Contractor Mobilization/Demobilization

A $1M mobilization cost was assumed to occur during the pre-production period. In addition, $1M was

assumed as a contractor demobilization cost in Year 14, during the last year of mining. These costs are

based on historical project experience.

Capitalized Mining Costs

Mining costs during the pre-production period are capitalized. Total material mined during the pre-

production period amount to 9.7Mt (including 1.0Mt of mineralized material and 8.7Mt of waste

material). The mining cost was assumed at $1.85/t material mined.

20.1.2 Mine Capital Cost

Open Pit Mobile Equipment

The OP mining activities for the Halilağa pit were assumed for this preliminary study to be undertaken

by a mining contractor. As such, no open pit mining fleet capital costs are included since the contractor

will be responsible for supplying an adequate mining fleet (including both primary and ancillary

equipment), with an estimated maximum capacity of 70,000 t/d total material, which would be sufficient

for the proposed LOM plan. The contractor would also supply the workshop and mobile equipment

maintenance facilities, as based on a budgetary quote received from an in-country mining contractor. A

mining contractor mobilization/demobilization capital cost of US$2.0M has been assumed.

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OP Development

The first year of pre-production predicts that approximately 10Mt of material will be mined (1Mt of mill

feed to stockpile). The cost of this pre-stripping has been added as an initial CAPEX item.

The cost estimate for the clearing and grubbing of the various pit areas and waste dumps cost estimate

was based on total area to be cleared and assumes clearing/grubbing to be undertaken with contractor-

operated fleet prior to full scale production taking place. Topsoil removal and storage (for use in

reclamation) has also been included in these costs. Total CAPEX of $0.7M is estimated.

Site Development

Site development costs include costs for earthworks as well as access and site roads. Cost estimates

are based on historical project experience.

Process Plant Capital Cost

The process plant design for the PEA study incorporates primary gyratory crushing, coarse plant feed

material stockpiling, SAG and ball milling, copper separation and dewatering, gold leach circuit with

gold room, and thickened tailings disposal.

The estimate has been prepared based on new budget quotes for major mechanical equipment and

high level estimates for bulk take-offs on earthworks, concrete, internal steel and major pipelines.

Factors have been applied to cover in-plant electrical distribution, instrumentation, piping, and

allowances for minor mechanical equipment and platework. Estimates for reagent systems, utility

supply (air/water), PLC control, and fire protection have been based on database pricing.

A summary of the process plant capital costs are outlined in Table 20.2.

.

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Table 20.2: Process Plant Capital Costs

Process Plant Capital Costs Unit Value

Primary Crushing & Storage $M 19.3

Grinding Area $M 48.1

Flotation Area & Regrind $M 27.8

Concentrate Dewatering $M 2.2

Tailings Thickening $M 5.7

Carbon In Leach (CIL) & Cyanide Destruction $M 14.8

Acid Wash, Stripping, Electrowinning & Refining $M 4.6

Reagents $M 4.0

Process Plant Utilities $M 5.1

Total $M 131.6

Source: JDS 2014

Earthworks and Civil Works

Earthwork MTO’s were based on AutoCAD models and by using limited topographical survey

information, and thus require further review when detailed topographical data becomes available. Unit

rates carried in the CAPEX were based on benchmarked data from Turkish contractors.

Concrete

Concrete MTO’s were based on preliminary layouts and/or included as estimated allowances based on

similar plants. Unit rates carried in the CAPEX were obtained by CH Consultants, a local contractor

from Turkey.

Mechanical Equipment

The following major process equipment was sized based on the design criteria and budget quotes were

obtained, as detailed below in Table 20.4.

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Table 20.3: Summary of Quoted Equipment

Equipment Description Quote Vendor Estimate (US$M)

Primary Gyratory Crusher Metso 3.6

SAG Mill Metso 10.5

Ball Mill Metso 8.6

Regrind Mill Metso 5.7

Cyclone Feed & Tailings Pumps ITT 0.9

Flotation Cells Outotec 6.3

Concentrate Pressure Filter Outotec 0.7

CN Detox System (Sulphur Burner) Gekko 6.1

Thickeners (Concentrate, Tails & CIL Feed) Outotec 3.9

Total Quoted Equipment 46.4

Source: JDS 2014

The following equipment was sized based on the design criteria and estimates were determined based

on database pricing:

Rock Breaker;

Conveyors;

Reclaim Apron Feeders;

SAG Discharge Screen; Primary & Regrind Cyclopacs;

CIL Screens and Submersible Carbon Pumps (Recessed Impellers);

Flotation Air Blowers;

Compressed Air Systems (CIL Tanks and Plant Utility Air);

Acid Wash, Stripping, Electrowinning & Refining Plant;

Reclaim Water Barge;

Make-up Water Pumps from Water Storage Pits 1 & 3;

Process Water Pumps; and

Fire Water Pump Skid.

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Structural Steelwork

Structural steelwork MTO’s were based on preliminary layouts and/or included as estimated allowances

based on similar plants. Unit rates carried in the CAPEX were obtained by CH Consultants, a local

contractor from Turkey.

These unit rates include supply, shop detailing, fabrication, surface preparation and final painting in the

shop, transport to site, site erection and paint touch-up as required.

Platework

The following equipment was sized based on the design criteria:

CIL Tanks; CIL Tank & Copper Concentrate Stock Tank Agitators; and

Fresh/Fire Water and Process Water Tanks.

The remaining mechanical bulks, such as pumps, vessels and receivers, were factored as a percentage

of overall mechanical costs for each area. Costs were determined based on unit rates obtained by CH

Consultants, a local contractor from Turkey which include supply, shop detailing, fabrication, surface

preparation and final painting in the shop, transport to site, site erection, and paint touch-up.

Piping, Electrical and Instrumentation

Piping, Electrical and Instrumentation costs were factored from mechanical equipment pricing for thecrushing and process plant areas based on actual historical factors for similar plants internationally.

Installation

Due to the contracting strategies in Turkey, the labour costs have not been separated from the unit

rates. Therefore, the unit rates confirmed by local Turkish contractors include the supply and installation

costs.

The mechanical equipment installation costs are based on a cost per weight ($/kg) of each piece of

major mechanical equipment. During the budgetary quotation stage, vendors were requested to provide

equipment weights in order to accurately calculate installation costs. Mechanical equipment installation

has been categorized by the following: “Rough Set” and “Final Installation”. Installation rates are shownbelow in Table 20.5.

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Table 20.4: Installation Rates

Description Units Unit Rate

Rough-Set Install $/kg 0.52

Final Install $/kg 1.2

Source: JDS 2014

In areas that equipment weights were not readily available, JDS has allowed a local labour rate of

$30/hr based on similar projects in the area with a productivity factors of 2.80. Labour rates are based

on a 50-hour work week, which is typical for recent projects executed internationally.

The labour rate includes the following items:

Base rate per hour;

Sick time;

Holiday pay;

Insurance;

Health and welfare;

Small tools and consumables;

Safety gear and clothing;

Site supervision;

Mobilization and demobilization;

Transportation – turnaround; Site and head office overhead; and

Contractor Markup & Profit.

The estimate is based on the majority of the work being carried out under fixed price or re-measurable

unit price contracts under a normal development schedule. No allowance is included for contracts on a

cost plus or fast-track accelerated schedule basis.

The erection of tankage, structural, mechanical, piping, electrical, instrumentation, and civil works would

be performed by experienced contractors, using a mix of national and non-national labour; the project is

allowed by the Turkish government to utilize non-nationals for construction as necessary to achieve the

required quality and meet the project schedule.

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On-Site Infrastructure Capital Cost

On-site infrastructure cost estimates were developed based on historical project experience. A

summary of the on-site infrastructure capital costs are outlined in Table 20.6.

It is noted that port facilities were excluded from the capital cost estimate as it was assumed that the

Port of Bandirma has the facilities to handle both incoming supplies and out-going concentrate.

Table 20.5: On-Site Infrastructure Capital Costs

On-Site Infrastructure Capital Costs Unit Value

Electrical Supply & Distribution $M 17.6

Water Supply & Distribution $M 5.5

Assay Laboratory $M 1.4

Sewage Treatment Plant $M 1.1

Admin Offices & Ancillary Facilities $M 0.6

Bulk Diesel Storage & Distribution $M 0.3

Plant Mobile Fleet $M 3.2

Total $M 29.6

Source: JDS 2014

Site Power Supply

The project requires electrical transmission, sub-transmission, and distribution infrastructure.

High-voltage transmission would be achieved using a 154 kV double-circuit overhead line from theexisting regional grid. The transmission line is proposed to be approximately 25 km long to the main

substation at site.

Medium-voltage power would be transmitted to the different motor control centres of consumption via a

6.6 kV line. Primary locations include:

Tailing storage facility;

Reclaim tunnel;

Primary crusher;

Truck shop; Water storage pits; and

Process plant.

Low-voltage power distribution would be at 380 V for offices and ancillary systems.

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The equipment required includes a 154 kV, 50M VA transformer down to 6.6 kV and all associated

substation components.

Potable Water Supply

There is an opportunity to utilize on-site potable water infrastructure to supply surrounding villages with

potable water. An allowance has been made for surplus supply from the potable water skid. Detailed

layout and design will need to be done during future phases to determine exact pipe lengths,

consumption rates and posting requirements. For purposes of this report, offsite potable water

infrastructure includes 9 km of buried 3” HPDE, one booster station and three 50,000 L water totes.

Pipeline Distribution Systems

High level engineering included sizing of pipelines for tailings, reclaim water and make-up water.

Existing topography information was used to determine optimal routing of lines to avoid farm land and

take advantage of existing infrastructure such as access roads. The pipelines have generally been

designed to be free draining. There is an intermediate low point on the common right of way (ROW) for

the reclaim water and tailings pipelines. Therefore, allowance has been included for a lined emergency

dump pond in this area when the pipelines need to be drained. The pond is sized so that each line can

be drained twice.

Pipeline costs were based North American carbon steel and HDPE pipe supply pricing and Turkey

installation unit rates. An allowance of 20% has been included on pipeline supply costs to allow for

fittings and valves.

Ancillary Buildings

The following ancillary buildings are included in the CAPEX estimate:

Main administration building with medical centre and training room;

Assay Laboratory;

Plan maintenance warehouse; and

Plant truck shop complete with minor equipment.

The costs of ancillary buildings were estimated based on historical unit rates per area for similar

projects. In addition to the building structures, the cost includes the supply of the buildings electrics,

fittings, and furnishings. Earthworks required for the project have been carried in the overall site

development. The total cost was estimated at $0.6M.

The cost to supply power and water services to the buildings form part of the water and electrical supply

and distribution costs. In addition, reagent storage facilities are included in the process plant cost

estimate.

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Mine Services Facilities

Mine service facilities and costs have been built into the Contract Mining unit rate. JDS has accounted

for the earthworks and site preparation for the mine service facilities.

Mobile Equipment

The mobile fleet required to support plant operations is based on used equipment. A 60% value of new

equipment was assumed in the cost estimate was consider for locally sourced, used equipment. Mobile

fleet required to support plant operations shown below in Table 20.7.

Table 20.6: Plant Support Mobile Equipment CAPEX Estimate

Description Number Total Cost (US$)

80T Rough Terrain Crane 1 590,000

Ambulance 1 90,000

Fire Truck 1 200,000

20T Boom Truck 1 210,000

Welding / Service Truck - Ford F550 (Custom) 1 80,000

5 T Fork Lift Zoom-Boom - Terex GTH-5519 1 50,000

F/E Loader - Cat 950K 1 200,000

Motor Grader - Cat 14M 1 390,000

Skid Steer Loader 1 40,000

Snow Plow/Sanding Truck 1 110,000

Stockpile Dozer - CAT D8T 1 500,000

Warehouse Forklift 1 20,000

1 T Diesel Crew Cab Pick-up - Ford F350 10 420,000

1 Ton Passenger Van 1 30,000

44 Passenger Bus 1 800,000

40T Tractor & Trailer 1 150,000

Total 3,180,000

Source: JDS 2014

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20.1.3 Tailings Storage Facility Capital Cost

Quantities

Primary material take offs have been calculated by hand calculations as well as AutoCAD Civil 3D

(2014). Where quantity estimates have been made using AutoCAD, additional checks have been

completed using Global Mapper 15.1 (2014) as well as hand calculations. The summary of quantities

for the tailings and water storage facilities are presented in Table 20.7.

Table 20.7: Scoping Level Cost Estimate Quantities

Tailings Storage Facility Unit Estimate

Bedding Material m3 837,900

Transition Material m3 400,200

Run of Quarry Rock m3 12,920,300

Excavations m3 1,262,500

Liner m2 2,985,200

Clear and Grub m2 3,084,900

Deposition Pipe m 7,200

Water Storage Facility

Diversion Excavation m3 2,700

Rip Rap m3 500

Concrete m3 20

Source: SRK 2014

Tailings and Water Facility Costs

The total scoping level cost excluding contingency is estimated at $120.9M throughout the LOM. This

cost estimate does not include geotechnical investigations or closure costs. The project costs broken

out by year are provided in Table 20.8.

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Table 20.8: Tailings and Water Storage Facility Costs by Year

Tailings & Water StorageFacility

Y-1 Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9

Direct TSF 17,633 12,638 10,568 8,228 8,472 9,038 6,471 10,045 7,587 8,092

Direct Water Storage 29.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Indirect Costs* 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Total 17,662 12,638 10,568 8,228 8,472 9,038 6,471 10,045 7,587 8,092

Notes:

Costs are presented in thousands (US$)

Indirects have been carried in JDS’ CAPEX Summary.

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20.1.4 Indirect Costs

Indirect costs total an estimated $37.6M, equal to 20% of the total direct costs. The various cost centres

that comprise the indirect costs are described in the following sections.

Heavy Construction Equipment

Heavy Construction Equipment costs have been calculated to be $2.9M, which equates to 1.5% of the

direct costs less mining equipment. Costs are intended to cover an 80 t crane and miscellaneous heavy

equipment for the duration of the project to support the construction.

Field Indirect Costs

Field indirect costs have been calculated to be $12.5M, which equates to 6.5% of the direct costs less

mining equipment. Costs are intended to cover the following:

Temporary Construction Facilities: work areas and bays, roads, walks and parking areas,

temporary buildings, temporary utilities for power and sewage, other minor temporary

construction.

Construction Services: general and final clean-up, material handling and warehousing, craft

training and testing, onsite services (soils exploration and soil testing, all labour and material

costs, concrete testing and security), operation and maintenance of temporary facilities,

surveying, pre-operational testing and start-up.

Freight and Logistics

Freight and logistics have been calculated to be $10.2M, which equates to 7.0% of the equipment and

material costs less mining equipment. Costs include ocean freight and Turkish inland freight, this figure

is based on factored historical data for similar projects.

Vendor representatives

Vendor representatives have been calculated to be $1.8M, which equates to 2% of the equipment and

material costs less mining. This figure is based on factored historical data for similar projects.

Start-Up & Commissioning / Capital Spares

Start-Up & Commissioning/Capital Spares have been $7.4M, which equates to 8.0% of the equipmentand material costs less mining equipment. This figure is based on factored historical data for similar

projects.

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First Fills

First fills have been calculated to be $2.8M, which equates to 3% of the equipment and material costs

less mining equipment. This figure is based on factored historical data for similar projects.

EPCM

For the purpose of the PEA estimate, $25.3M or 11% of the direct costs was selected to cover the cost

of EPCM services, which includes detailed engineering, procurement, project management and home

office services as well as construction management. This was calculated on direct costs that excluded

the mine equipment and mine development.

Owner’s Cost

For the purpose of the PEA estimate, $6.4M or 2.5% of the direct and indirect costs were selected to

cover the Owner’s Costs, which includes insurance, owner’s team costs during pre-production, and

project development. This figure is based on factored historical data for similar projects.

Sustaining Capital

A total of $1.3M was estimated as sustaining capital required and the costs are assumed to occurbetween Year 1 and Year 12 (inclusively). Sustaining capital costs were estimated as 1% of processplant capital costs which is in line with previous project experience.

Capital costs estimated during the mine life for the tailings storage amount to $103.3M. These costs are

detailed in section 20.1.8.

Mine Closure

The mine closure and rehabilitation cost has been estimated as $50.2M ($63M with 25% contingency).

The breakdown of the cost estimate is given in Table 20.14 as per the closure assumptions detailed in

Section 19.2.3.

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Table 20.9: Mine Closure and Rehabilitation Cost Estimate

FacilityCostM$

UnitCost$/unit

Dimension Units Comments

Process Plant

Dismantling &Demo

4.0 Similar operations

Re-vegetation 0.1 0.51 97,000 m2 Unit cost as defined by MFWM

TSF

Top Liner(geosynthetics)

20.4 6 3,400,000 m2

Assume Class-I TSF requirements as defined by Landfillregulation

Top Liner (50 cmclay)

7.7 4.5 1,700,000 m3

Assume Class-I TSF requirements as defined by Landfillregulation

Topsoil (fromstripping)

5.8 4 1,451,565 m3 Re-spread topsoil stripped during construction

Revegetation 1.7 0.51 3,400,000 m2

Unit cost as defined by MFWM - to be paid to MFWM atstart-up

WRD

Top Liner (50 cmclay)

4.9 4.5 1,097,500 m3 ARD/ML mitigation - reduce MAP infiltration

Topsoil (fromstripping)

3.8 4 947,333 m3 Re-spread topsoil stripped during construction

Revegetation 1.1 0.51 2,195,000 m2

Unit cost as defined by MFWM - to be paid to MFWM atstart-up

PassiveTreatment

0.1 Similar operations

Open Pit

Bunding (lining) 0.1 6.6 18,000 m3 Bund: 1 m crest + 3H:1V slope on both sides

Environmental 0.5 Consulting, Engineering, Equipment, Supplies, Labs

Total Direct Cost 50.2

Contingency(25%)

12.5

Total ClosureCost

62.7

Source: SRK 2014

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20.1.5 Contingency

A contingency of 25% for all capital was used with the exception of mining equipment. The purpose ofthe contingency provision is to make allowance for uncertain elements (“known unknowns”) of costs to

cover such factors as:

Limited information on site conditions, especially concerning sub-surface conditions and the

engineering properties of excavated materials;

Completeness and accuracy of quantity take-offs and estimate assembly and consolidation

based on the level of engineering and design undertaken at study level;

Accuracy of materials and labour rates (excluding extreme variations that would be covered

under contingency);

Accuracy of productivity expectations; and accuracy of equipment pricing.

Major cost categories (permanent equipment, material purchase, installation, subcontracts, pipelines,

indirect costs and Owner’s costs) were identified and analyzed. An overall contingency of $107.7M was

obtained, representing 25% of the total CAPEX.

20.2 Operating Costs Estimate

The operating costs (OPEX) of the project have been estimated based on the scope defined in previous

sections of this report.

20.2.1 Operating Cost Summary

The OPEX estimate is based on a variety of sources including cost service data, vendor quotes, first

principle calculations, and reference projects. The summary OPEX costs are shown in Table 20.15 and

Figure 20.3.

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Table 20.10: OPEX Estimate Summary

Operating Cost US$/t processed LOM US$M

Mining‡ 4.05 503.7

Re-Handle* 0.01 1

Processing (incl. Tails) 8.35 1,038.20

G&A 0.7 86.6

Total OPEX 13.11 1,629.40

(‡): Excludes capitalized pre-stripping costs(⁰): Based on $1.85/t mined (assuming average LOM 1.3 strip ratio) (*) Re-handle cost amounts to $1/t re-handled. Total material re-handled amounts to 1M tonnes.

Source: JDS 2014

20.2.2 Basis of Estimate

The main OPEX assumptions used in the study are shown in Table 20.9.

Table 20.16: Main OPEX Assumptions

Item Unit Cost Estimate Comment

Diesel Fuel $/litre 1.73 Exclusive of VAT

Electricity $/kWh 0.123 Energy and demand cost estimate

Source: JDS 2014

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Figure 20.

Source: JDS

PRO J ECT

e: December 20

: Life of Mi

2014

PEA TECH

, 2014

e Operating

NICAL REPO

Costs

RT

20-22

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20.2.3 Mining Operating Costs

Costs for OP mining activities for the Halilağa Project, assumed to be undertaken by a miningcontractor, were direct quotes from mining contractors in the area that were checked against build-ups

from first principles, as well as experience of similar-sized OP operations and local conditions. OP

mining costs for both mineralized and waste material take into account variations in haulage profiles

and equipment selection. Equipment efficiency was estimated based on Halilağa conditions (e.g. haul

routes for each phase). Local labour rates and diesel fuel pricing estimates were utilized for estimation

purposes. The OP mining costs encompass pit and dump operations, road maintenance, and mine

supervision and technical services cost. The average OPEX for the LOM plan was estimated to be

$1.85/t material mined or $4.05/t plant feed, for contractor pit and dump operations, road maintenance,

mine supervision, and technical services (Table 20.10). No contingency was added.

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Table 20.11: Mine Average OPEX Estimate by Function

OP FunctionUnit Cost Estimate

($/t mined)

Drill and Blast 0.49

Load 0.53

Haul 0.68

Roads/Dumps/Support Equipment/Maintenance/Supervision 0.15

Total 1.85

Source: JDS 2014

20.2.4 Material Re-handle Costs

Mineralized material re-handle costs are assumed to occur in Year 1 of the project. The costs account

for the material re-handled mined during Year -1. A total of 1Mt of mineralized material is re-handled in

Year 1. A $1/t material re-handle cost has been assumed, based on previous project experience.

20.2.5 Processing Operating Costs

The operating costs for the concentrator are based on estimated direct costs for processing a nominal

annual throughput of 9.1Mtpa (25,000 t/d) of mineralized material. The processing facility availability

has been assumed at 75% for the crushing plant and 92% for the rest of the plant. These availabilities

incorporate both scheduled and unscheduled shutdowns. The overall annual operating cost is

approximately US$76.2M, which equates to US$8.35/tonne processed.

A summary of the average OPEX per tonne of feed treated for the Project is outlined in Table 20.12 andFigure 20.4. The costs have been separated into the key cost components. All costs have been based

on estimates as of Q4 2014.

Table 20.12: Processing Average Unit OPEX Estimate

Summary Estimate ($/t milled)

Labour 0.46

Ancillary Equipment 0.08

Reagents & Consumable Materials 2.55

Expenses (including Power) 3.58

Contingency (25%) 1.67

Total Unit Cost 8.35

Source: JDS 2014

The OPEX are considered to have accuracy in the order of ± 30 %, based on the assumptions listed in

this section of the report.

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Source: JDS

20.2.6

The OPE

based on

characteri

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: Chart of P

2014

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Table 20.13: Derivation of Plant OPEX Estimate

Cost Category Source Of Cost Data

PowerPower rate was based on a quote received from a Turkey

supplier with respect to our load estimate

LabourManning schedules and rates developed by JDS from

similar sized projects

ReagentsConsumptions from test work and benchmarking data; unit

prices were obtained by CH consultant.

ConsumablesConsumptions predicted from test work and experience and

unit prices from suppliers and JDS database.

Maintenance Materials Calculated from industry experience and JDS database

Source: JDS 2014

Major items which are included the Processing OPEX Estimate are:

Labour for supervision, management and reporting of onsite organizational and technical

activities directly associated with the processing plant;

Labour for operating and maintaining plant mobile equipment and light vehicles, process plant

and supporting infrastructure;

Costs associated with direct operation of the processing plant, including all fuels, reagents,

consumables and maintenance materials;

Fuels, lubricants, tires and maintenance materials used in operating and maintaining the plant

mobile equipment and light vehicles based on similar projects and the JDS database;

Tailings discharge pumping and return water, including management of the TSF;

Cost of power as supplied from the local power grid;

Operation of raw water supply facility; and

Allowance for assaying of production samples based on similar projects.

Labour

JDS developed a workforce plan for the process plant based on applicable schedules for similar plants.

Annual salary levels for the various labour classifications have been taken from the SRK 2012 - PEA

report (escalated by CPI). A standardised, 12 hour shift schedule based on four days “on”, four days“off” has been adopted for all hourly workers. The processing plant will operate on a 24 hour/day, 365-

days/year basis. Table 20.15 shows the plant labour complement.

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Table 20.14: Site Labour Cost Summary

Labour Quantity

Staff

Mill Manager 1

Chief Metallurgist 1

Metallurgist 2

Operations Foreman 4

Maintenance Foreman 4

Maintenance Planner 1

Training Officer 1

Labour

Control Room/Plant Operator 4Crusher Operator 4

Grinding Operator 4

Flotation Operator 4

Filtration and concentrate handling operator 4

CIL Circuit Operator 4

Cyanide Detox, Reagents & Carbon Handling 4

Elution & Refinery Operator 4

Tailings & Water supply operator 4

Assayer/Sample bucker 6

Maintenance Personnel. Mech. Elect. Inst. 12

Labourers/Helpers 8

Warehousemen/ laborer 6

Source: JDS 2014

Power

Power would be supplied to the mine site from the local power grid. Unit power costs are estimated to

be approximately $0.123/kWh. (Based from 2014 Turkish Tariff (base price) and includes fees for

energy loss, retail sales cost, distribution fees, system fee, energy fund fee, Turkish Broadcasting

Corporation Share, consumption fee and Reading fee.)

The power requirements for the plant have been developed from the major mechanical equipment

power consumption. The calculated load on which the power costs have been based is determined by

applying a factor to the specific power draw for the major installed equipment (excluding installed

standby equipment) and a utilization factor.

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A summary of power costs for the process plant, TSF, return water and raw water pumping is given in

Table 20.21.

Table 20.15: Process Plant Power Cost Summary (M$)

Area Annual Cost Estimate

Total Annual Cost ($M) 32.7

Unit Cost ($/t) 3.58

Source: JDS 2014

Reagents and Consumable Materials

Reagent consumptions have been estimated from metallurgical test work or comparable operations.Although, reagent consumptions will vary according to metallurgical and production parameters,

average predicted consumptions have been used for this exercise. Generally, these values are in line

with the average consumptions from the test work to date.

The unit prices for the reagents were obtained from CH Consultants, Unit costs include an allowance for

delivery to site but do not include duties, brokerage, handling charges or applicable taxes. The average

LOM consumptions and the unit costs are presented in Table 20.22.

Table 20.16: Reagent Consumptions and Unit Costs

Unit Cost($/kg)

Consumption($/t milled)

Annual Estimate(M$)

Hydrated lime 0.1 0.06 0.6

PAX 3.2 0.08 0.7

AP404 3.5 0.01 0.1

MIBC 4.1 0.08 0.8

Flocculant 4.0 0.08 0.7

Cyanide 3.0 0.59 5.4

Sodium Hydroxide 0.5 0 0.0

Copper Sulfate 2.7 0.03 0.2

Metabisulfite 0.2 0.03 0.2

HCl 0.4 0 0.0

Hydrated lime for cyanide leach and detox 0.1 0.1 0.9

Antiscalent 3.9 0.02 0.2

Carbon 3.3 0.02 0.2

TOTAL 1.09 10.0

Source: JDS 2014

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Plant and Maintenance Consumables

Consumables include major items, such as crusher and mill liners and grinding media and minor items

such as spare parts, site fuel and distribution piping. Expected consumptions for each material type

have been estimated based on factors associated with the overall throughput and industry experience

to calculate LOM averages.

Water Reservoir and Tailings Storage Facility Operating Cost

OPEX for the water supply and tailings discharge to the tailings storage facility consists of the power

cost for the ongoing pumping and associated maintenance costing. The cost was captured in the

processing costs.

20.2.7 General and Administration Operating Costs

G&A Summary

The average G&A OPEX for the supporting facilities and administration for a typical year are estimated

to be $6.44M million per annum or $0.70 /t milled. These costs are assumed to consist of both fixed

costs, independent of plant throughput or mining rate and partially variable, changing in direct

proportion to the plant throughput rate. The G&A costs are summarized below in Table 20.23.

Table 20.17: General and Administration Costs Summary

G&A Item M$/year $/t processed

G&A Labour 1.4 0.16

G&A Equipment 2.1 0.23

Materials 0.2 0.02

Expenses 2.7 0.30

TOTAL $6.4 $0.70

Source: JDS 2014

G&A Labour Costs

Estimated G&A labour costs totalled $1.4M per annum as summarized in Table 20.24.

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Table 20.18: G&A Labour Cost Summary

G&A Labour Cost Summary Labour Quantity

Surface Infrastructure & Maintenance 24

First Aid 2

Environment 1

Administration 8

Health and Safety 1

Human Resources 1

IT & Communications 1

Security 4

TOTAL 42Source: JDS 2014

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21 ECONOMIC ANALYSIS

An engineering economic model was developed to estimate annual cash flows and sensitivities of the

project. Pre-tax estimates of project values were prepared for comparative purposes, while after-tax

estimates were developed and are likely to approximate true investment value. It must be noted,

however, that tax estimates involve many complex variables that can only be accurately calculated

during operations and, as such, the after-tax results are only approximations.

Sensitivity analyses were performed for variations in metal prices, head grades, operating costs, capital

costs, US$:Turkish Lira exchange rate and discount rates to determine their relative importance as

project value drivers.

This technical report contains forward-looking information regarding projected mine production rates,

construction schedules and forecasts of resulting cash flows as part of this study. The mill head gradesare based on sufficient sampling that is reasonably expected to be representative of the realized grades

from actual mining operations. Factors such as the ability to obtain permits to construct and operate a

mine, or to obtain major equipment of skilled labour on a timely basis, to achieve the assumed mine

production rates at the assumed grades, may cause actual results to differ materially from those

presented in this economic analysis.

The estimates of capital and operating costs have been developed specifically for this project and are

summarized in Section 18 of this report (presented in 2014 dollars). The economic analysis has been

run with no inflation (constant dollar basis).

21.1 Assumptions

One metal price scenario was utilized to prepare the economic analysis. However, a sensitivity analysis

on the metal prices was completed and is outlined in Section 18.

All costs, metal prices and economic results are reported in US dollars (US$ or $) unless statedotherwise. LOM plan tonnage and grade estimates are demonstrated in Table 21.1.

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Table 21.1: Life of Mine Plan Summary


Mine Life Years 13.6

Resource Mined M tonnes 124.3

Waste Mined M tonnes 157.6

Total Mined M tonnes 281.9

Strip Ratio w:o 1.3

Plant Throughput Rate tpd 25,000

Average Head Grade

Cu % 0.34%

Au g/t 0.34

Source: JDS 2014

Other economic assumptions used in the economic analysis include the following:

Discount Rate of 7% (sensitivities using other discount rates have been calculated – Refer to

Section 21.6, Table 21.4 and Figure 21.7);

Closure cost of $63.7M (which includes a 25% contingency) was considered;

Revenues, costs and taxes are calculated for each period in which they occur rather than actual

outgoing/incoming payment;

Working capital was calculated as three months of operating costs in (mining, processing,

tailings storage, environment, and G&A) in Year 1 (assumed to be required in Year -1). Theworking capital is recuperated during the last year of production (Year 14). Total working capital

considered in Year -1 amounts to $29M;

Depreciation for CAPEX has been considered based on Turkish regulations and asset class;

Results are presented on a 100% equity basis; and

No management fees or financing costs have been considered.

These assumptions are typical and appropriate for a PEA-level study.

The economic analysis excludes all pre-development and sunk costs up to the start of detailed

engineering (i.e. exploration and resource definition costs, engineering fieldwork and studies costs,environmental baseline studies costs, etc.).

Table 21.2 outlines the metal price assumption used in the economic analysis. Metal pricing is based on

approximate average spot metal prices during December 2014 published by the London Metal

Exchange.

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Table 21.2

Metal Price

Copper Pric

Gold Price

Source: JDS

21.2 R

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Figure 21.

Source: JDS

PRO J ECT

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Figure 21.

Source: JDS

Figure 21.

Source: JDS

PRO J ECT

e: December 20

: Flotation

2014

: Au Recov

2014

PEA TECH

, 2014

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Figure 21.

Source: JDS

21.3 R

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2.6

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Figure 21.

$54.6M ov

PRO J ECT

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: Au Recov

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Source: JDS

21.4 T

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.

Successful designation for tax purposes of the Halilağa project as a Strategic Investment

relating to gold production, and as a Priority Investment as related to copper production. Withholding taxes on repatriation to shareholders and their respective parent companies were

not considered.

21.5 Economic Results

The reader is cautioned that this PEA is preliminary in nature and includes the use of inferredmineral resources that are considered too speculative geologically to have the economicconsiderations applied to them that would enable them to be categorized as mineral reservesand, as such, there is no certainty that the PEA economics will be realized. The PEA uses 31%inferred mineralized material.

The project is economically viable with an after-tax internal rate of return (IRR) of 43.1% and a netpresent value using a 7% discount rate (NPV7%) of $473.8M using the Base Case metal prices. Table

21.5 summarizes the economic results of the project.

The break-even copper price for the project (using the Base Case metal prices) is approximately

$1.83/lb, based on LOM presented herein and a gold price of US$1,200/oz. Table 21.3 demonstrates

the projected cash flows for the project.

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Table 21.3: Summary of Results


Cu Payable LOM M lbs 779.4

Au Payable LOM k oz 924.2

Operating Costs US$/t processed 13.11

Total Capital Costs Incl. Contingency US$M 558.5

Discount Rate % 7.0

Pre-Tax NPV US$M 510.9

Pre-Tax IRR % 45.8

Pre-Tax Payback Years 1.2

After-Tax NPV US$M 473.8

After-Tax IRR % 43.1

After-Tax Payback Years 1.3

Cu Cash Cost‡ US$/Cu lb 2.50

Cu Cash Cost (Net of By-Products)* US$/Cu lb 1.08

Cu Cash Cost (incl. Sustaining Capital)** US$/Cu lb 2.78

Cu Cash Cost (Incl. Sustaining Capital) Net of By-Products⁰ US$/Cu lb 1.35

‡ Cash Cost = (Treatment Charge + Refining Charges + Royalties + Operating Costs) / Payable Cu lbs* Cash Cost (Net of By Products) = ((Treatment Charge + Refining Charges + Royalties + Operating Costs)- (Payable Au* Au Price))/Payable Cu lbs** Cash Cost (incl. Sustaining Capital) = (Treatment Charge + Refining Charges + Royalties + Operating Costs +Sustaining Capital Costs) / Payable Cu lbs⁰ Cash Cost (incl. Sustaining Capital) Net of By-Products = ((Treatment Charge + Refining Charges + Royalties +Operating Costs + Sustaining Capital)- (Payable Au * Au Price))/Payable Cu lbs

Source: JDS 2014

7/24/2019 2015 Halilaga


HALILAĞA

Effective Da

.

Figure 21.

Source: JDS

PRO J ECT

e: December 20

: Annual Af

2014

PEA TECH

, 2014

er-Tax Cas

NICAL REPO

Flows

RT

21-9

7/24/2019 2015 Halilaga


Economic Model Pilot Gold - Halilaga - Final Economic Model - 2015-01-27

HalilagaHalilaga PreliminaryEconomic AssessmentAnnualCashFlows 0 1 2 3 4 5 6 7 8 9 10

Unit Source LOM -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15

Cu Price $/lb link 2.90 2.90 2.90 2.90 2.90 2.90 2.90 2.90 2.90 2.90 2.90 2.90 2.90 2.90 2.90 2.90 2.90 2.90 2.90

Au Price $/oz link 1,200.00 1,200 1,200 1,200 1,200 1,200 1,200 1,200 1,200 1,200 1,200 1,200 1,200 1,200 1,200 1,200 1,200 1,200 1,200

ProductionSc hedule

Ore Mined tonnes link 124,273,978 0 0 999,944 8,134,667 9,116,058 9,127,593 9,130,920 9,126,264 9,121,560 9,123,576 9,127,062 9,123,660 9,121,098 9,123,366 9,128,060 9,124,300 5,645,850 0

Waste Mined tonnes link 157,644,541 0 0 8,667,570 9,280,372 13,848,915 14,349,435 14,709,632 14,675,324 14,738,878 14,701,114 14,816,009 14,686,596 14,688,222 7,443,154 880,420 158,900 0 0

Total Mined tonnes calc 281,918,519 0 0 9,667,514 17,415,039 22,964,973 23,477,028 23,840,552 23,801,588 23,860,438 23,824,690 23,943,071 23,810,256 23,809,320 16,566,520 10,008,480 9,283,200 5,645,850 0

Mining Rate tpd calc 57,793 0 0 26,486 47,712 62,918 64,321 65,317 65,210 65,371 65,273 65,597 65,234 65,231 45,388 27,420 25,433 15,468 0Strip Ratio w:o calc 1.3 0.0 0.0 8.7 1.1 1.5 1.6 1.6 1.6 1.6 1.6 1.6 1.6 1.6 0.8 0.1 0.0 0.0 0.0

Head Grades

Cu Head Grade % link 0.34% 0.00% 0.00% 0.95% 0.73% 0.57% 0.38% 0.31% 0.28% 0.28% 0.31% 0.26% 0.29% 0.23% 0.25% 0.24% 0.24% 0.26% 0.00%

Au Head Grade g/t link 0.34 0.00 0.00 0.64 0.46 0.46 0.41 0.36 0.33 0.32 0.35 0.27 0.33 0.28 0.31 0.28 0.26 0.31 0.00

Processing Schedule

Total Ore Milled tonnes link 124,273,978 0 0 0 9,134,611 9,116,058 9,127,593 9,130,920 9,126,264 9,121,560 9,123,576 9,127,062 9,123,660 9,121,098 9,123,366 9,128,060 9,124,300 5,645,850 0

Plant Throughput tpd calc 25,001 0 0 0 25,026 24,976 25,007 25,016 25,003 24,991 24,996 25,006 24,996 24,989 24,996 25,008 24,998 24,998 0

Head Grades

Cu % link 0.34% 0.00% 0.00% 0.00% 0.76% 0.57% 0.38% 0.31% 0.28% 0.28% 0.31% 0.26% 0.29% 0.23% 0.25% 0.24% 0.24% 0.26% 0.00%

Au g/t link 0.34 0.00 0.00 0.00 0.48 0.46 0.41 0.36 0.33 0.32 0.35 0.27 0.33 0.28 0.31 0.28 0.26 0.31 0.00

MetalContained

Cu lbs calc 920,486,151 0 0 0 1 52,3 11 ,582 11 5,5 18,3 48 76 ,210 ,99 1 6 2,89 9,3 71 5 5,98 3,0 08 5 6,73 1,6 15 6 2,65 8,66 8 5 2,48 1,8 76 5 8,86 6,7 34 46 ,55 2,92 2 49 ,83 1,45 2 49 ,22 7,54 7 48 ,54 8,94 0 32 ,66 3,09 6 0

Au oz calc 1,357,281 0 0 0 141,450 134,029 120,393 106,891 96,057 94,586 102,515 79,718 96,590 82,832 90,580 80,770 75,442 55,428 0

FlotationCircuit

% Cu calc 88.2% 0.0% 0.0% 0.0% 94.1% 91.8% 88.5% 87.0% 86.1% 86.2% 87.0% 85.6% 86.5% 84.7% 85.2% 85.1% 85.0% 85.6% 0.0%

% Au calc 58.4% 57.8% 57.8% 57.8% 58.6% 58.5% 58.5% 58.4% 58.3% 58.3% 58.4% 58.2% 58.3% 58.3% 58.3% 58.2% 58.2% 58.3% 57.8%

Cu tonnes calc 368,280 0 0 0 65,002 48,112 30,585 24,815 21,860 22,179 24,713 20,375 23,089 17,883 19,258 19,003 18,719 12,688 0

Cu lbs calc 811,917,014 0 0 0 1 43 ,305 ,02 1 10 6,0 67,5 80 67 ,428 ,29 4 5 4,70 7,2 75 4 8,19 2,2 55 4 8,89 6,6 29 5 4,48 3,2 46 4 4,91 9,3 67 5 0,90 2,7 46 3 9,42 4,6 74 4 2,45 6,35 3 41 ,89 4,47 2 41 ,26 7,51 0 27 ,97 1,59 2 0

Au g calc 24,642,137 0 0 0 2,577,559 2,440,640 2,189,394 1,941,282 1,742,694 1,715,776 1,861,027 1,443,960 1,752,454 1,500,832 1,642,453 1,463,159 1,365,943 1,004,963 0

Au oz calc 792,263 0 0 0 82,870 78,468 70,391 62,414 56,029 55,163 59,833 46,424 56,343 48,253 52,806 47,042 43,916 32,310 0

PullFactor calc 101 0 0 0 42 57 90 110 125 123 111 134 119 153 142 144 146 133 0

% link 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0%

g /t A u in C u Co nc c al c 20.1 0.00 0.00 0.00 11.90 15.22 21.48 23.47 23.92 23.21 22.59 21.26 22.77 25.18 25.59 23.10 21.89 23.76 0.00

dmt calc 1,227,599 0 0 0 216,674 160,372 101,950 82,716 72,866 73,931 82,377 67,917 76,964 59,609 64,193 63,343 62,395 42,292 0

wmt calc 1,334,347 0 0 0 235,515 174,317 110,815 89,909 79,202 80,359 89,541 73,823 83,656 64,793 69,775 68,852 67,821 45,970 0

MoistureContent % link 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0%

Cu Payable % link 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0%

Cu Min.Deduction % link 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% 0.0% 0.0% 0.0%

PayableCu basedon %Payable % calc 28.8% 28.8% 28.8% 28.8% 28.8% 28.8% 28.8% 28.8% 28.8% 28.8% 28.8% 28.8% 28.8% 28.8% 28.8% 28.8% 28.8% 28.8% 28.8%

PayableCu basedon min.deduction % calc 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0% 30.0%

lbs calc 779,440,333 0 0 0 1 37 ,572 ,82 0 10 1,8 24,8 76 64 ,731 ,16 2 5 2,51 8,9 84 4 6,26 4,5 64 4 6,94 0,7 64 5 2,30 3,9 17 4 3,12 2,5 92 4 8,86 6,6 37 3 7,84 7,6 87 4 0,75 8,09 9 40 ,21 8,69 3 39 ,61 6,80 9 26 ,85 2,72 8 0

US $ calc 2,260,376,966 0 0 0 3 98 ,9 61 ,1 79 2 95 ,2 92 ,1 42 1 87 ,7 20 ,3 71 1 52 ,3 05 ,0 53 1 34 ,1 67 ,2 37 1 36 ,1 28 ,2 16 1 51 ,6 81 ,3 58 1 25 ,0 55 ,5 17 1 41 ,7 13 ,2 46 1 09 ,7 58 ,2 92 1 18 ,1 98 ,4 87 1 16 ,6 34 ,2 10 1 14 ,8 88 ,7 47 7 7, 87 2, 91 1 0

Au Payable % link 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0% 96.0%

Min Au Deduction g/t in conc link 1.0 1.0 1.0 1.0 1.0 1.0 1.0 1.0 1.0 1.0 1.0 1.0 1.0 1.0 1.0 1.0 1.0 1.0 1.0oz calc 722,683 0 0 0 72,868 70,380 64,428 57,364 51,539 50,675 54,897 42,471 51,714 44,483 48,713 43,205 40,234 29,713 0

US$ calc 867,219,046 0 0 0 87 ,441 ,60 8 8 4,4 55,7 25 77 ,313 ,98 9 6 8,83 6,8 92 6 1,84 6,4 99 6 0,81 0,0 70 6 5,87 6,9 86 5 0,96 5,3 82 6 2,05 6,2 07 5 3,37 9,4 66 5 8,45 5,0 24 5 1,84 5,8 59 4 8,28 0,3 17 3 5,65 5,0 22 0

U S$ /d mt co nc l in k 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00

US$ calc 112,939,139 0 0 0 19,933,990 14,754,194 9,379,399 7,609,882 6,703,631 6,801,610 7,578,719 6,248,366 7,080,665 5,484,044 5,905,756 5,827,598 5,740,386 3,890,899 0

US $/lb link 0.092 0.092 0.092 0.092 0.092 0.092 0.092 0.092 0.092 0.092 0.092 0.092 0.092 0.092 0.092 0.092 0.092 0.092 0.092

US $ calc 71,708,511 0 0 0 12,656,699 9,367,889 5,955,267 4,831,747 4,256,340 4,318,550 4,811,960 3,967,278 4,495,731 3,481,987 3,749,745 3,700,120 3,644,746 2,470,451 0

US $/oz link 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00

US $ calc 5,058,778 0 0 0 510,076 492,658 450,998 401,549 360,771 354,725 384,282 297,298 361,995 311,380 340,988 302,434 281,635 207,988 0

U S$ /d mt co nc l in k 62.71 62.71 62.71 62.71 62.71 62.71 62.71 62.71 62.71 62.71 62.71 62.71 62.71 62.71 62.71 62.71 62.71 62.71 62.71

US $ calc 76,978,484 0 0 0 13,586,861 10,056,350 6,392,929 5,186,839 4,569,145 4,635,928 5,165,599 4,258,840 4,826,129 3,737,884 4,025,320 3,972,048 3,912,605 2,652,008 0

NetReve nues from FlotationCirc uit US $ c alc 2,86 0,9 11,1 00 0 0 0 4 39,7 15 ,160 34 5,0 76,7 76 2 42 ,8 55 ,767 20 3,11 1,9 28 18 0,12 3,8 49 18 0,82 7,47 2 199 ,61 7,78 3 16 1,24 9,1 16 18 7,00 4,93 4 150 ,12 2,46 3 162 ,63 1,70 3 154 ,67 7,87 0 149 ,58 9,69 2 104 ,30 6,58 7 0

Cleaner Tails Leach

Recovery % Au link 15.0% 15.0% 15.0% 15.0% 15.0% 15.0% 15.0% 15.0% 15.0% 15.0% 15.0% 15.0% 15.0% 15.0% 15.0% 15.0% 15.0% 15.0% 15.0%

g calc 6,332,427 0 0 0 659,938 625,316 561,698 498,701 448,156 441,295 478,284 371,926 450,642 386,455 422,603 376,833 351,977 258,602 0

oz calc 203,592 0 0 0 21,217 20,104 18,059 16,034 14,409 14,188 15,377 11,958 14,488 12,425 13,587 12,115 11,316 8,314 0

% Au link 99.0% 99.0% 99.0% 99.0% 99.0% 99.0% 99.0% 99.0% 99.0% 99.0% 99.0% 99.0% 99.0% 99.0% 99.0% 99.0% 99.0% 99.0% 99.0%

oz calc 201,556 0 0 0 21,005 19,903 17,878 15,873 14,264 14,046 15,223 11,838 14,344 12,301 13,451 11,994 11,203 8,231 0

US$ calc 241,867,421 0 0 0 25 ,2 06 ,354 2 3,8 83,9 68 21 ,454 ,08 6 1 9,04 7,9 23 1 7,11 7,3 57 1 6,85 5,2 96 1 8,26 8,1 02 1 4,20 5,7 48 1 7,21 2,2 85 1 4,76 0,6 80 1 6,14 1,33 8 14 ,39 3,16 1 13 ,44 3,80 0 9 ,87 7,32 3 0

Au RefiningCharge US $/oz link 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00 7.00

US $ calc 1,410,893 0 0 0 147,037 139,323 125,149 111,113 99,851 98,323 106,564 82,867 100,405 86,104 94,158 83,960 78,422 57,618 0

NetAu Payable from Cleaner Tails Leach US $ calc 240,456,527 0 0 0 25,059,317 23,744,645 21,328,937 18,936,810 17,017,506 16,756,974 18,161,538 14,122,881 17,111,880 14,674,576 16,047,180 14,309,200 13,365,378 9,819,705 0

US $ calc 3,1 01,3 67 ,628 0 0 0 4 64 ,774 ,47 7 36 8,8 21,4 20 2 64 ,184 ,70 4 22 2,04 8,7 38 19 7,14 1,3 56 19 7,58 4,4 46 21 7,77 9,3 21 17 5,37 1,9 97 20 4,11 6,8 13 16 4,79 7,0 39 17 8,67 8,8 83 16 8,98 7,0 70 16 2,95 5,0 70 11 4,12 6,2 92 0

US$/t milled calc 24.96 0.00 0.00 0.00 50.88 40.46 28.94 24.32 21.60 21.66 23.87 19.21 22.37 18.07 19.58 18.51 17.86 20.21 0.00

Base for AuRoyalty

Au Revenues -Flotation Circuit US$ link 867,219,046 0 0 0 87 ,441 ,60 8 84,4 55 ,7 25 77 ,313 ,98 9 6 8,8 36,8 92 6 1,84 6,4 99 6 0,81 0,0 70 6 5,87 6,9 86 5 0,96 5,3 82 6 2,05 6,2 07 5 3,37 9,4 66 5 8,45 5,0 24 5 1,84 5,8 59 4 8,28 0,3 17 3 5,65 5,0 22 0

Au Revenues -CleanerTails Leach US$ link 241,867,421 0 0 0 25 ,206 ,35 4 2 3,8 83,9 68 21 ,454 ,08 6 1 9,04 7,9 23 1 7,11 7,3 57 1 6,85 5,2 96 1 8,26 8,1 02 1 4,20 5,7 48 1 7,21 2,2 85 1 4,76 0,6 80 1 6,14 1,3 38 1 4,39 3,1 61 13 ,44 3,80 0 9 ,87 7,32 3 0

Total Au Revenues US$ calc 1,109,086,466 0 0 0 112,647,962 108,339,692 98,768,075 87,884,814 78,963,857 77,665,366 84,145,088 65,171,129 79,268,492 68,140,146 74,596,363 66,239,019 61,724,117 45,532,345 0

Au RCs -Flotation Circuit US$ link 5,058,778 0 0 0 510,076 492,658 450,998 401,549 360,771 354,725 384,282 297,298 361,995 311,380 340,988 302,434 281,635 207,988 0

Au RCs -CleanerTails Leach US$ link 1,410,893 0 0 0 147,037 139,323 125,149 111,113 99,851 98,323 106,564 82,867 100,405 86,104 94,158 83,960 78,422 57,618 0

AuRCs US$ calc 6,469,671 0 0 0 657,113 631,982 576,147 512,661 460,622 453,048 490,846 380,165 462,400 397,484 435,145 386,394 360,057 265,605 0

NetAu Revenues (Base for Royalties) US$ calc 1,102,616,795 0 0 0 111,990,849 107,707,711 98,191,928 87,372,153 78,503,234 77,212,318 83,654,241 64,790,964 78,806,092 67,742,662 74,161,217 65,852,625 61,364,060 45,266,740 0

Au Royalty US$ calc 28,668,037 0 0 0 2,911,762 2,800,400 2,552,990 2,271,676 2,041,084 2,007,520 2,175,010 1,684,565 2,048,958 1,761,309 1,928,192 1,712,168 1,595,466 1,176,935 0

Base for CuRoyalty

Cu Revenues -Flotation Circuit US$ link 2,260,376,966 0 0 0 3 98 ,9 61 ,1 79 2 95 ,2 92 ,1 42 1 87 ,7 20 ,3 71 1 52 ,3 05 ,0 53 1 34 ,1 67 ,2 37 1 36 ,1 28 ,2 16 1 51 ,6 81 ,3 58 1 25 ,0 55 ,5 17 1 41 ,7 13 ,2 46 1 09 ,7 58 ,2 92 1 18 ,1 98 ,4 87 1 16 ,6 34 ,2 10 1 14 ,8 88 ,7 47 7 7, 87 2, 91 1 0

TC/RCs/OffsiteCosts US$ calc 261,626,134 0 0 0 46 ,177 ,55 0 3 4,1 78,4 32 21 ,727 ,59 5 1 7,62 8,4 68 1 5,52 9,1 16 1 5,75 6,0 88 1 7,55 6,2 78 1 4,47 4,4 85 1 6,40 2,5 25 1 2,70 3,9 15 1 3,68 0,8 21 1 3,49 9,7 65 1 3,29 7,73 7 9 ,01 3,35 9 0

NetCu Revenues (Base for Royalties) US$ c alc 1,99 8,7 50,8 32 0 0 0 3 52 ,7 83 ,62 8 26 1,1 13,7 09 1 65 ,992 ,77 6 13 4,67 6,5 85 11 8,63 8,1 21 12 0,37 2,1 28 13 4,12 5,0 80 11 0,58 1,0 32 12 5,31 0,7 21 9 7,05 4,3 77 10 4,51 7,6 66 10 3,13 4,44 5 101 ,59 1,01 0 68 ,85 9,55 2 0

Cu Royalty US$ calc 25,983,761 0 0 0 4,586,187 3,394,478 2,157,906 1,750,796 1,542,296 1,564,838 1,743,626 1,437,553 1,629,039 1,261,707 1,358,730 1,340,748 1,320,683 895,174 0

Royalties US $ calc 54,651,797 0 0 0 7,497,949 6,194,879 4,710,896 4,022,472 3,583,380 3,572,358 3,918,636 3,122,118 3,677,998 3,023,016 3,286,921 3,052,916 2,916,149 2,072,109 0

US $ c alc 3,04 6,7 15,8 30 0 0 0 4 57,2 76 ,528 36 2,6 26,5 42 2 59,4 73 ,808 21 8,02 6,2 67 19 3,55 7,97 6 194 ,01 2,08 8 213 ,86 0,68 5 172 ,24 9,87 8 200 ,43 8,81 6 161 ,77 4,02 3 175 ,39 1,96 2 165 ,93 4,15 4 160 ,03 8,92 2 112 ,05 4,18 3 0

US$/t milled calc 24.52 0.00 0.00 0.00 50.06 39.78 28.43 23.88 21.21 21.27 23.44 18.87 21.97 17.74 19.22 18.18 17.54 19.85 0.00

Operating Costs

US$/t min ed calc 1.79 0 0 0 1.85 1.85 1.85 1.85 1.85 1.85 1.85 1.85 1.85 1.85 1.85 1.85 1.85 1.85 0.00

US$ link 503,664,359 0 0 0 32 ,2 17 ,822 4 2,4 85,2 00 43 ,432 ,50 2 4 4,10 5,0 21 4 4,03 2,9 38 4 4,14 1,8 10 4 4,07 5,6 77 4 4,29 4,6 81 4 4,04 8,9 74 4 4,04 7,2 42 3 0,64 8,06 2 18 ,51 5,68 8 17 ,17 3,92 0 10 ,44 4,82 3 0

U S$ /t mi ll ed c al c 0.01 0.00 0.00 0.00 0.11 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00

US$ link 999,944 0 0 0 999,944 0 0 0 0 0 0 0 0 0 0 0 0 0 0

U S$ /t mi ll ed c al c 8.35 0.00 0.00 0.00 8.35 8.35 8.35 8.35 8.35 8.35 8.35 8.35 8.35 8.35 8.35 8.35 8.35 8.35 0.00

US$ link 1,037,687,716 0 0 0 76,2 74 ,0 02 7 6,11 9,0 84 76,2 15 ,402 76 ,24 3,18 2 7 6,2 04,3 04 7 6,16 5,0 26 7 6,18 1,8 60 7 6,21 0,9 68 7 6,18 2,5 61 7 6,16 1,1 68 7 6,18 0,1 06 7 6,21 9,3 01 7 6,18 7,9 05 4 7,14 2,8 48 0

U S$ /t mi ll ed c al c 0.00 0.00 0.00 0.00 0.01 0.01 0.00 0.00 0.00 0.00 0.01 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00

US$ link 479,027 0 0 0 72,304 60,382 39,256 44,338 33,757 26,677 50,273 32,762 33,416 30,633 33,316 21,912 0 0 0

U S$ /t mi ll ed c al c 0.70 0.00 0.00 0.00 0.70 0.70 0.70 0.70 0.70 0.70 0.70 0.70 0.70 0.70 0.70 0.70 0.70 0.70 0.00

US$ link 86,580,572 0 0 0 6,364,002 6,351,076 6,359,113 6,361,431 6,358,187 6,354,910 6,356,314 6,358,743 6,356,373 6,354,588 6,356,168 6,359,438 6,356,818 3,933,413 0

US$/t milled calc 13.11 0.00 0.00 0.00 12.69 13.71 13.81 13.88 13.88 13.89 13.88 13.90 13.88 13.88 12.41 11.08 10.93 10.90 0.00

US$ c alc 1,62 9,4 11,6 18 0 0 0 1 15,9 28 ,074 12 5,01 5,7 42 1 26,0 46 ,272 126 ,75 3,97 2 126 ,62 9,18 6 126 ,68 8,42 3 126 ,66 4,12 3 126 ,89 7,15 3 126 ,62 1,32 3 126 ,59 3,63 1 113 ,21 7,65 1 101 ,11 6,33 9 9 9,71 8,6 43 6 1,52 1,0 83 0

US$ calc 1,417,304,212 0 0 0 3 41 ,348 ,45 4 2 37,6 10 ,799 1 33 ,427 ,53 6 9 1,2 72,2 95 6 6,9 28,7 90 6 7,32 3,6 65 8 7,19 6,5 61 4 5,35 2,7 25 7 3,81 7,4 92 3 5,18 0,3 92 6 2,17 4,3 10 6 4,81 7,8 15 6 0,32 0,2 78 5 0,53 3,1 00 0

U S$ /t mi ll ed c al c 11.40

CapitalCostsCapitalized Mining Costs US$ link 17,884,901 0 0 17,884,901 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0

C on tra ct or Mo bi li za ti on /D emo bi li za ti on U S$ l in k 2,000,000 0 0 1,000,000 0 0 0 0 0 0 0 0 0 0 0 0 0 1,000,000 0

Mining US$ link 588,815 0 135,427 453,387 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0

Site Development US$ link 5,543,748 0 1,275,062 4,268,686 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0

Process Plant US$ link 131,639,423 0 30,277,067 101,362,356 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0

On-Site Infrastructure US$ link 29,615,312 0 6,811,522 22,803,790 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0

Tailings Storage US$ link 128,311,371 0 5,756,479 19,271,690 12,638,601 10,568,327 8,228,895 8,473,113 9,039,084 6,471,931 10,046,018 7,588,012 8,093,160 6,725,686 8,960,750 6,449,626 0 0 0

Indirects US$ link 37,571,501 0 8,641,445 28,930,056 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0

EPCM US$ link 25,298,566 0 5,818,670 19,479,896 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0

Owner's Costs US$ link 6,382,138 0 1,467,892 4,914,247 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0

Sustaining US$ link 15,796,731 0 0 0 1,316,394 1,316,394 1,316,394 1,316,394 1,316,394 1,316,394 1,316,394 1,316,394 1,316,394 1,316,394 1,316,394 1,316,394 0 0 0

Closure US$ link 50,178,430 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 12,544,607 12,544,607

Total Pre-Contingency Total US$ calc 450,810,937 0 60,183,565 220,369,009 13,954,996 11,884,721 9,545,289 9,789,507 10,355,478 7,788,325 11,362,412 8,904,406 9,409,554 8,042,080 10,277,144 7,766,021 0 13,544,607 12,544,607

Contingency US$ calc 107,731,509 0 15 ,04 5,89 1 50 ,37 1,02 7 3 ,4 88 ,749 2,9 71,1 80 2 ,386 ,32 2 2,44 7,3 77 2,58 8,8 69 1,94 7,0 81 2,84 0,6 03 2,22 6,1 02 2,35 2,3 89 2,01 0,5 20 2,56 9,28 6 1 ,94 1,50 5 0 3 ,13 6,15 2 3 ,13 6,15 2

To tal Cap ita lCos ts incl. Continge nc y US$ c alc 55 8,54 2,4 45 0 75 ,22 9,45 6 270 ,74 0,03 6 17,4 43 ,744 1 4,8 55,9 01 11 ,931 ,61 2 1 2,23 6,8 84 1 2,94 4,3 47 9,73 5,4 07 1 4,20 3,0 15 1 1,13 0,5 08 1 1,76 1,9 43 1 0,05 2,6 00 12 ,84 6,43 0 9 ,70 7,52 6 0 16 ,68 0,75 9 15 ,68 0,75 9

Total Pre-Production Costs US$ calc 345,969,491 0 75,229,456 270,740,036 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0

Total Sustaining/Closure US$ calc 212,572,954 0 0 0 17,443,744 14,855,901 11,931,612 12,236,884 12,944,347 9,735,407 14,203,015 11,130,508 11,761,943 10,052,600 12,846,430 9,707,526 0 16,680,759 15,680,759

Working Capital US$ calc 0 0 0 28,982,019 0 0 0 0 0 0 0 0 0 0 0 0 0 -28,982,019 0

Net Pre-Tax Cash Flow C$ calc 858,761,766 0 - 75 ,2 29 ,4 56 - 29 9, 72 2, 05 4 3 23 ,9 04 ,7 09 2 22 ,7 54 ,8 98 1 21 ,4 95 ,9 24 7 9, 03 5, 41 1 5 3, 98 4, 44 3 5 7, 58 8, 25 8 7 2, 99 3, 54 6 3 4, 22 2, 21 7 6 2, 05 5, 55 0 2 5, 12 7, 79 2 4 9, 32 7, 88 0 5 5, 11 0, 29 0 6 0, 32 0, 27 8 6 2, 83 4, 35 9 - 15 ,6 80 ,7 59

C umu la ti ve Pr e-T ax Ca sh F lo w C $ c al c 0 - 75 ,2 29 ,4 56 - 37 4, 95 1, 51 0 - 51 ,0 46 ,8 01 1 71 ,7 08 ,0 97 2 93 ,2 04 ,0 22 3 72 ,2 39 ,4 32 4 26 ,2 23 ,8 75 4 83 ,8 12 ,1 33 5 56 ,8 05 ,6 79 5 91 ,0 27 ,8 96 6 53 ,0 83 ,4 46 6 78 ,2 11 ,2 38 7 27 ,5 39 ,1 18 7 82 ,6 49 ,4 07 8 42 ,9 69 ,6 85 9 05 ,8 04 ,0 44 8 90 ,1 23 ,2 85

Pre-Tax Payback Years calc 1.2 1.0 0.2 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

Pre-Tax IRR % calc 46%

Pre-Tax NPV7% US$ calc 510,881,371 - 37 4 ,9 5 1, 51 0 - 47 , 70 7 ,2 9 0 1 4 9, 9 76 , 50 2 2 3 9, 34 1 ,8 2 0 2 8 3, 9 79 ,6 8 1 3 0 3, 8 91 ,7 3 2 3 2 2, 38 4 ,4 5 3 3 4 6, 75 0 ,5 9 3 3 4 3, 9 83 ,6 1 7 3 5 5, 23 4 ,1 2 3 3 4 4, 76 8 ,2 0 3 7 2 7, 53 9 ,1 1 8 7 8 2, 64 9 ,4 0 7 8 4 2, 96 9 ,6 8 5 9 0 5, 80 4 ,0 4 4 8 9 0, 12 3 ,2 8 5

Au RefiningCharge

Recoveries fromFlotation Circuit

Metalin Concentrate

Cu RefiningCharge

ConcentrateProduced

PayableCu in Cu Conc

Cu TreatmentCharge

Au Payablein Cu Conc

Cu ConcentrateGrade

OffsiteCosts

TotalOpera ting Costs

NetOperating Income

Processing

Au Recovered

Au Payable

NetReve nues After Royalties

NetRevenues

Mining

Tailings Management& Detox

G&A

OreRehandle

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21.6 Sensitivities

A sensitivity analysis was performed on the Base Case metal pricing scenarios to determine which

factors most affect the project economics. The analysis revealed that the project is most sensitive to

metal prices, followed by head grade and operating costs. The project showed least sensitive to capital

costs. Table 21.6 along with Figure 21.7 outline the results of the sensitivity test performed on the after-

tax NPV7% for the Base Case evaluated.

The project was also tested under various discount rates. The results of this sensitivity test are

demonstrated in Table 21.7.

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HALILAĞA

Effective Da

.

Table 21.4

Variable

Metal Price

Cu Price

Au Price

Head Grad

OPEX

CAPEX

Source: JDS

Figure 21.

Source: JDS

PRO J ECT

e: December 20

: Sensitivity

(Combined)

e

2014

: After-Tax

2014

PEA TECH

, 2014

Results for

Sensitivity G

NICAL REPO

ase Case

After-T

-15%

193.2

281.4

385.7

201.9

602.4

523.7

raph for Ba

RT

PV

x NPV7% (US

e Case Res

M)

100%

473.8

473.8

473.8

473.8

473.8

473.8

ults

15%

754.4

666.3

562.0

748.5

345.3

424.0

21-3

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Foreign Exchange Sensitivity

The project was tested for changes in foreign exchange rate to determine the level of the project’s

exposure for changes in USD:TL rate. Although the PEA was completed in USD, an analysis was

completed using assumptions to allow for cost sensitivities with respect to particular costs that would be

realistically incurred in local currency (Turkish Lira (TL)). Table 21.5 outlines the assumptions made

with respect to proportion of costs that would be based in USD (i.e. fuel costs) and TL (i.e. majority of

labour rates).

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Table 21.5: Sensitivity Analysis – Foreign Exchange Assumptions by Cost Area

Cost Currency Split Assumption (Value)

OPEX

Mining% of costs based in USD 60

% of costs based in TL 40

Material Re-Handle% of costs based in USD 60


Processing% of costs based in USD 30


Tailings Storage% of costs based in USD 50


G&A

% of costs based in USD 20


Total OPEX Split% of costs based in USD 40


CAPEX

Capitalized Mining Costs% of costs based in USD 60


Contractor Mobilization/Demobilization% of costs based in USD 0


Mining% of costs based in USD 60


Site Development

% of costs based in USD 30


Process Plant% of costs based in USD 30


On-Site Infrastructure% of costs based in USD 60


Tailings Storage% of costs based in USD 60


Indirects% of costs based in USD 60


EPCM% of costs based in USD 30


Owner’s Costs% of costs based in USD 40


Sustaining Capital% of costs based in USD 40


Total CAPEX Split% of costs based in USD 30


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Based on the assumptions in Table 21.7, the project was tested against the 0.44 USD:TL base case

exchange rate which was used for cost conversion in the PEA cost estimate. Table 21.8 shows theeconomic results of the project using various discount rates and the assumptions of the sensitivity

analysis for pre-tax results. The results show that every +/-0.01 USD:TL change in F/X rate impacts the

project Pre-Tax NPV7% by about +/-$20M.

Table 21.6: Result of Foreign Exchange Rate Sensitivity Analysis

F/X Rate (USD:TL) Pre-Tax NPV7% ($M) Pre-Tax IRR (%)Pre-Tax Payback

(Years)

0.43 530.2 47.3 1.2

0.44 510.9 45.8 1.2

0.45 491.5 44.3 1.3

Source: JDS 2014

Table 21.7: Discount Rate Sensitivity

Discount Rate Pre-Tax NPV (US$M) After-Tax NPV

0% 858.8 802.9

5% 589.7 548.6

7% 510.9 473.8

8% 476.0 440.8

10% 414.0 381.9

12% 360.6 331.3

Source: JDS 2014

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22 ADJACENT PROPERTIES

22.1 Ağı Dağı and Kirazlı

The Biga district hosts numerous high-sulphidation gold systems, as well as porphyry copper-gold

targets. Halilağa is located mid-way between two advanced stage gold projects owned by Alamos Gold

Inc. (“Alamos Gold”). The project is 12 km to the NW of Ağı Dağı, and 12 KM SE of Kirazli, respectively

(seeFigure 22.1). Alamos published a pre-feasibility study on a combined Ağı Dağı – Kirazlıproject in

July 2012, and is planning to begin mine construction at Kirazlıupon receipt of all necessary permits and

approvals. Kirazlıis projected to be the first “modern” gold mine built in Çanakkale State, with Ağı Dağı

expected to follow within 18 months.

Figure 22.1: Location of Halilağa and Adjacent Properties

Source: Alamos 2014

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The current situation with the Ağı Dağı and KirazlıProjects are described as follows from the Alamos

website:“In August 2013, the Turkish Ministry of Environment and Urbanization (the "Ministry") formally

approved the Company's EIA for the Kirazlıproject. However, in January 2014, the Çanakkale

Administrative Court in Turkey (the "Court") granted an injunction order in response to a lawsuit

claiming that the Ministry's approval of the EIA for the Company's Kirazl ıproject failed to assess the

"cumulative impacts" of the Kirazlıproject and other potential mining projects in the region. The Ministry

contested the Court's decision on the basis that there was no applicable regulatory requirement to

include such an assessment in an EIA report at the relevant time. Notwithstanding this factor, in the

third quarter, the Çanakkale Administrative Court, as the first instance court, cancelled the Ministry's

EIA approval in relation to the Kirazlımain project due to the lack of cumulative impact assessment

("CIA"). The Court's basis for the injunction did not relate to concerns with any technical aspect of the

Kirazlıproject.The Ministry and the Company appealed this ruling to the Turkish High Administrative Court. The

appeal decision remains pending, but is expected to be finalized within three to six months. In order to

address the CIA requirements and concerns of the Court, the Company has prepared and submitted a

CIA assessment for the Kirazlıproject, which is currently under review by the Ministry.

In August 2014, the Ministry signed and issued formal approval in the form of an EIA Positive Decision

Certificate for the Agi Dagi project. A new legislative process was recently implemented in Turkey,

whereby any legal challenge to an EIA must be registered within 30 days of the approval by the

Ministry. This deadline has passed and the Company is not aware of any legal challenges filed,

therefore, the Company does not currently anticipate the same legal challenges that have faced the

KirazlıEIA. Obtaining forestry and operating permits are the next steps in the permitting process for the

project.

The Company has budgeted spending of $4.8 million in Turkey in 2014 for permitting, community and

government relations and general administration costs only. Given the continuing delay in receipt of key

permits, the Company reduced its headcount early in 2014 and curtailed spending significantly in

Turkey. A full development budget for Kirazlıand Agi Dagi will be re-initiated once the required permits

are received.”

The Ağı Dağı and KirazlıProjects are high-sulphidation epithermal gold and silver deposits. Table 22.1

summarizes the current stated mineral resource estimates.

Information provided in this section is not necessarily indicative of the mineralization on the Halilağa

property. In addition, the author has not confirmed this publicly available disclosure and has not talked

to Alamos to confirm the data.

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Table 22.1: 2013 Resource for the Ağı Dağı and KirazlıProperties

2013 Ağı Dağı and KirazlıMineral Resources(1),(2),(3),(4),(5),(6)

Measured and Indicated and Inferred Mineral Resources As At December 31, 2013

Resource Area

Measured and Indicated Mineral Resources Inferred Resources

Tonnes Grade Grade Contained Contained Tonnes Grade Grade Contained Contained

(000s)(g/tAu)

(g/tAg)

OuncesAu

OuncesAg

(000s)(g/tAu)

(g/tAg)

Ounces Au Ounces Ag

Ağı Dağı 88,204 0.58 4 1,638,911 11,356,774 15,204 0.41 2.71 202,266 1,322,910

Kirazli 33,917 0.71 8.5 772,470 9,266,615 5,872 0.59 8.78 110,865 1,657,310

Çamyurt 17,730 0.89 6.14 508,877 3,498,222 2,791 0.95 5.8 84,922 520,151

TOTAL 139,851 0.65 6.36 2,920,258 24,121,611 23,867 0.52 4.56 398,053 3,500,371

(1)

The economic cut-off grade is 0.2 g/t Au(2) In-pit measured and indicated mineral resource blocks are exclusive of pit-contained reserves.

(3) Measured and indicated and inferred resources for the Ağı Dağı project, which includes the Baba, Ayitepe,

Deli, and Fire Tower zones, are pit constrained with cut-off determined as a net of process value of $0.10 pertonne, for each model block. The determination was based on a US$1,400 per ounce gold price and aUS$24.00 per ounce silver price, a December 31, 2013 resource model, pit slope angles ranging from 40°to48°, and estimated costs and recoveries based on the pre-feasibility study specifications. The resources werethen tabulated by gold cut-off grade.(4)

Measured and indicated, and inferred resources for the Kirazlıproject, including Rockpile, are pit constrainedwith cut-off determined as a net of process value of $0.10 per tonne, for each model block. The determinationwas based on a US$1,400 per ounce gold price and a US$24.00 per ounce silver price, a December 31, 2013resource model, pit slope angles ranging from 38° to 48°, and estimated costs and recoveries based on the pre-feasibility study specifications. The resources were then tabulated by gold cut-off grade.(5)

Measured and indicated and inferred resources for the Çamyurt project are pit-constrained with cut-offdetermined as a net of process value of $0.10 per tonne, for each model block. The determination was basedon a $1,400 per ounce gold price and a $24/oz silver price, a December 31, 2013 resource model, average pitslope angle of 45°, and estimated costs and recoveries based on the pre-feasibility study specifications. Theresources were then tabulated by gold cut-off grade.(6)

Mineral resources are not mineral reserves and do not have demonstrated economic viability.

Source: Alamos website: ttp://www.alamosgold.com/home/default.aspx ) 2014

The QP of this report cannot verify the Ağı Dağı and Kirazlıresource estimates and they are not

necessarily indicative of the mineralization at Halilağa.

22.1 TV Tower

TV Tower, a large exploration stage project jointly owned by a subsidiary of Pilot Gold (60%) and Teck

Madencilik, a Turkish subsidiary of Teck (40%) is 15 km due west of Halilağa. TV Tower hosts a gold-silver resource defined by Pilot Gold in January 2014, and several recent high sulphidation gold and

copper-gold discoveries believed to be similar to the styles of mineralization present on the Halilağa

tenure.

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Pilot Gold is project operator at TV Tower and can increase its interest in the project to 60% through

sole-funding of a $21-million exploration program over a three-year period. Pilot Gold has completed itssecond-year funding commitment and is nearing completion of the earn-in. See Figure 22.1 for the

location of TV Tower relative to Halilağa.

According to the Pilot Gold Website:

“TV Tower is a high-sulphidation epithermal and porphyry gold-copper property located in the Biga

District of northwestern Turkey, close to established infrastructure. Discovered by Teck Resources and

Pilot Gold (formerly Fronteer Gold), the project hosts numerous gold discoveries to date and abundant

untested targets. A project-first resource estimate on the KCD target returned an Indicated Mineral

Resource of 996,000 AuEq ounces (23.06 Mt at 1.34 g/t AuEq) and an Inferred Mineral Resource of

351,000 AuEq ounces (10.77 Mt at 1.01 g/t AuEq) 2. Recent drilling has focused on the Kayal ı and

Karaay ı targets (collectively “K2”) in the southern part of the tenure. Work at K2 demonstrates thepresence of a 4-km-long silica cap with multiple gold-oxide targets, associated supergene copper zones

and two copper-gold porphyry systems. Drilling by Pilot has returned some of the highest-grade gold,

silver and copper intervals ever reported in northwestern Turkey.”

Subsequent to the publication of the KCD resource, Pilot Gold has focused exploration drilling on the

Valley, Hilltop and Columbaz porphyry targets. Drilling at each of the aforementioned targets returned

long runs of continuous copper-gold mineralization, starting at surface and representing new

discoveries. Further targets of note at TV Tower include the K2 oxide gold trend, and the Gumusluk

and Kestanelik vein hosted gold targets.

Information provided on TV Tower is not necessarily indicative of the mineralization on the Halilağa

property. In addition, the author has not confirmed any information or data in this publicly availabledisclosure.

With common ownership, although with different ownership structure, there are potential synergies

between the Halilağa and TV Tower projects. There is the potential that processing facilities,

infrastructure and personnel could, in some way, be shared. These synergies should be further

explored as the two projects are developed.

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Table 22.2: Küçükdağ (KCD) Resource by Redox State at 0.5 g/t AuEq Cut-off

Redox State Resource ClassTonnes(x10

6)

Au(g/t)

Ag(g/t)

Cu(%)

AuEq(g/t)

Metal (x103)

Au (oz) Ag (oz) Cu (lb)

Total Indicated 23.06 0.63 27.6 0.16 1.34 470 20,479 78,859

Inferred 10.77 0.15 45.7 0.06 1.01 53 15,831 14,883

Oxide Indicated 2.3 0.1 60 0.01 1.12 7 4,447 692

Inferred 0.78 0.13 41.2 0.02 0.85 3 1,028 379

Transition Indicated 3.37 0.26 41.3 0.06 1.04 28 4,470 4,288

Inferred 1.31 0.39 36.1 0.05 1.06 16 1,520 1,324

Sulphide Indicated 17.38 0.78 20.7 0.19 1.43 435 11,563 73,878

Inferred 8.69 0.12 47.6 0.07 1.02 33 13,283 13,179


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HA LILAĞA PROJECT – PEA TECHNICAL REPORT


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23 INTERPRETATION AND CONCLUSIONS

Industry standard mining, process design, construction methods and economic evaluation practices

were used to assess the Halilağa Project. In JDS’s opinion, there is adequate geological and other

pertinent data available to generate a PEA.

Based on current knowledge and assumptions, the results of this study show that the project has

positive economics (within the very preliminary parameters of a PEA) and should be advanced to the

next level of study, a preliminary feasibility study (PFS).

This study achieves its original objective of reducing project CAPEX by reducing the project production

rate from 50 kt/d (SRK 2012) to 25 kt/d while maintaining similar NPV results. The study, when

compared to the previous PEA (SRK 2012), shows improved payback period and IRR and thus

significantly improves the economic robustness of the project.

This PEA demonstrates the inherent advantages presented by the distribution of gold and copper within

the Halilağa resource. The high grade gold and copper zone situated near surface presents the

opportunity for rapid payback of capital in a mining scenario, while the existing infrastructure present in

Çanakkale State mitigates the need for extensive infrastructure development in the form of roads,

power generation and ports.

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HA LILAĞA PROJECT – PEA TECHNICAL REPORT


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24 RISKS AND OPPORTUNITIES

As with almost all mining ventures, there are a large number of risks and opportunities that can

influence the outcome of the Halilağa Project. Most of the risks are based on a lack of scientific

information (test results, drill results, etc.) or the lack of control over external drivers (metal price,

exchange rates, etc.). The following section identifies the most significant potential risks currently

identified for the project, almost all of which are common to mining projects at this early stage of project

development.

Subsequent higher-level engineering studies would be needed to further refine these risks and

opportunities, identify new ones, and define mitigation or opportunity implementation plans. While a

significant amount of information is still required to do a complete assessment, at this point there do not

appear to be any fatal flaws for the project.

24.1 Risks

Table 24.1 identifies what are currently deemed to be the most significant internal project risks, potential

impacts, and possible mitigation approaches.

The most significant potential risks associated with the project are uncontrolled dilution, geotechnical

stability of pit walls and tailings facility, lower metal recoveries than those projected, operating and

capital cost escalation, permitting and environmental compliance, unforeseen schedule delays, changes

in regulatory requirements, ability to raise financing and metal prices. These risks are common to most

mining projects, many of which can be mitigated with adequate engineering, planning and pro-active

management.

External risks are, to a certain extent, beyond the control of the project proponents and are much more

difficult to anticipate and mitigate, although, in many instances, some risk reduction can be achieved.

External risks are things such as the political situation in the project region, metal prices, exchange

rates and government legislation. These external risks are generally applicable to all mining projects.

Negative variance to these items from the assumptions made in the economic model would reduce the

profitability of the mine and the mineral resource estimates.

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Table 24.1: Internal Project Risks

Risk Explanation Potential Impact Possib

RecoveriesFlotation recoveries are largely basedon a limited number of samples andtests.

If life-of-mine recovery of Cu or Au islower than projected, projecteconomics could be negativelyimpacted.

Conducconfirm

Water Supply

The ability to provide water to theprocessing facility is based on the useof the abandoned coal pits for waterstorage.

Inability to utilize the coal pits wouldrequire the design and construction ofadditional water storage facilities. Thiscould increase project CAPEX andfootprint.

Advancand pereliminavolume

Permit Acquisition

The ability to secure a mining permit isof paramount importance as is the

negotiation with current farmers andother stakeholders.A Cumulative Impact Assessment

(CIA) may be required as part of

permitting and may add another layer

of investigation.

EIA litigation and pending changes to

mining law and royalty regime

Failure to secure a mining permitwould stop the project while increasedroyalties may negatively impact theproject economic viability.

The devwith thealong wSocial Idesign considelocal pe

Development Schedule

The development could be delayed fora number of reasons and could impactproject economics and the ability tomine during a period of high metalsprices.

Delays in schedule could alter projecteconomics through lost revenues atpotentially higher metal prices.

If an agfollowedASAP.

Inability to upgrade inferredresources to measured or indicated

The PEA mine plan uses 31% inferredresources which cannot be used at ahigher level of study

If none of the inferred resources canbe upgraded to indicated then themineable tonnage would be reducedof what is presented here and projecteconomics could be negativelyaffected

A well pcampaianalysisto be unamountconvert

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Risk Explanation Potential Impact Possib

TSF Location and Stability

The geotechnical condition of the soilsunder the TSF embankment and rockstorage facilities must be investigatedto confirm the location suitability anddesign adequacy.

A more robust facility design or havingto move the waste storage facilitiescould significantly impact both OPEXand CAPEX.

Conduclevel of

Smelter Terms

Smelter terms used in the study areonly preliminary and could affect theproject economics if the terms(payable %, deductions and/orpenalties, TC/RC’s) change.

A reduction in the net smelter returnwould have a direct effect on projecteconomics.Low concentrate grade and/or thepresence of deleterious elements inthe concentrates could impact thedesirability of the concentrate and theprice smelters are willing to pay.

Conducto identCu con

Power SupplyIt was assumed that sufficient powerwould be available locally from thegrid.

If power is not available, sitegenerated power may have to beinvestigated or an upgrade to theregional grid capacity. These optionswould likely add costs to the projectand potential delays.

Conduca clearepower a

Open PitOpen pit slope angles were preparedbased on limited availability ofstructural and geotechnical data.

Presence of unfavorably orientedstructures, weak rock masses orhydraulic gradients behind pit wallsmay result in shallower slope anglesbeing required.

Conducprogramat the n

Plant Process Copper and GoldRecovery

A reduction in recovery would have anegative impact on the projecteconomics.

A reduction in recovery by 1% Cuwould lower the NPV7% byapproximately $15 M.

Conducto confi

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24-4

.

24.2 Opportunities

Table 24.2: Project Opportunities

Opportunity Explanation Potential Benefit

Exploration Potential

The expansion of the deposit maybe possible with further resourcedrilling. See Sections 8.3.1, 8.3.2and 9 for details.

The expansion of the deposit resources couldpotentially lead to an increased LOM.

Silver and/or MolybdenumCredits

Develop new resource model toinclude silver and investigate thepotential benefit of Mo recovery.

Including silver into the Cu concentrate mayprovide another stream of revenue andimprove project economics. A Mo concentratemay add an alternate base metal stream ofincome.

Geotech Drilling

Slope parameters could be

readjusted and reconfirmed.Reducing the slope by 2 degreescaused a ~$100 million decrease inproject economics.

Supportive geotech drilling could provideinformation to steepen the final pit slopes andreducing the strip ratio for the LOM.

Copper/Gold PriceMetal price has the biggest singleimpact on the project economics.

The impact is shown in the economicsensitivity section.

Power Cost

The cost of power is relatively highat $0.12/kWhr and a reduction inpower costs would decrease overallOPEX of the project.

Electrical power costs are 43% of theprocessing costs. Therefore, a 10% reductionin power costs may potentially achieve asmuch as a 5% reduction in overall processingOPEX and an overall increase in the after-taxNPV7% of +5%.

Metallurgical Recoveries

The potential for optimizedrecoveries vs. concentrate gradecould improve the projecteconomics.

The NPV of the project may be improved with

optimization of metallurgical recoveries andconcentrate grade. The sensitivity of theproject with respect to changes to processrecovery is similar to the project’s sensitivity tochanges in processed head grades which hasbeen included in the sensitivity analysis of thePEA.

Optimizing CIL Circuit

Gold is associated with pyrite and ismore finely disseminated than thechalcopyrite. Pyrite is being rejectedin the rougher circuit resulting in~25% of the gold reporting directlyto tails. A pyrite float on the roughertails could be applied to maximizegold reporting to the CIL circuit.Applying a regrind of the total CILfeed would assist in liberating themore finely disseminated goldparticles.

Optimizing the feed into the CIL circuit couldresult in increasing Au recovery to the doré.

TSF construction material

There is a possibility that the TSFcould be constructed out of earth fillrather than rock fill. The availabilityof suitable material would need to

The TSF is a major project cost and the aswitch from a rock fill to an earth fill dam couldpotentially reduce initial and sustainingCAPEX.

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24-5

.

Opportunity Explanation Potential Benefit

be investigated.

Cost Synergies with TVTower

Continual development of the sisterproperty TV Tower could result incost synergies between the twoprojects

Site infrastructures and logistics could beshared between the two properties spreadingthe costs and reducing CAPEX and OPEXthrough economies of scale.

Tax and InvestmentIncentives

Additional tax and investmentincentives may be available for theproject that could reduce payabletaxes and improve the after-taxproject economics. The taxcalculations in the PEA considerstandard depreciation, howeveradditional capital write-off

allowances may be available.

Optimizing taxation or providing other forms ofinvestment incentive would improve the after-tax project NPV.

Pipeline Supply CostsPipeline CAPEX estimates werebased on North American supplycosts.

The next project phase should examine localsupply costs to potentially reduce CAPEX forpipelines.

Tailings ThickenerConstruction

CAPEX includes estimate for 46 mdiameter carbon steel thickener.

There is a potential savings if the thickener isinstalled in an earthen basin (compacted fillwith liner).

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25-6

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25 RECOMMENDATIONS

JDS recommends that the project be advanced to the next level of study, a preliminary feasibility

study (PFS), contingent on positive results from a drilling campaign. Prior to undertaking the PFS,

the potentially mineable resource will have to be drilled more extensively in an attempt to convert

inferred to indicated resources. After drilling, sampling and assaying, a new resource model will be

required. A high-level estimate of the resource drilling and re-estimation cost is as follows:

Table 25.1: Cost Estimate for Additional Drilling and an Updated Mineral Resource Estimate to

Support a PFS

Item and Description New Mineral Resource Estimate Cost (M$)

Resource definition drilling (8,000 m x $160/m) 1.28

Assaying ($40/m average) 0.32

Camp operations, trucks, fuel, supplies 2.3

Resource Estimation 0.16

Salaries and staff costs 3.2

Condemnation drilling under surface facilities (2,000 m x$160)

0.32

Mineral Resource Estimate 7.58

Source: JDS 2014

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26-7

.

26 LIST OF ABBREVIATIONS

Units of measurement used in this report conform to the SI (metric) system. A complete list of

abbreviations is shown in Table 26.1.

Table 26.1: Units of Measure & Abbreviations

°C degree Celsius

°F degree Fahrenheit

A ampere

a annum

Ag silver

Au gold

bbl barrelsC$ or CAD Canadian dollars

cal calorie

cfm cubic feet per minute

cm centimetre

cm2 square centimetre

Cu copper

d day

dia. diameter

dmt dry metric tonne

dwt dead-weight ton

ft foot

ft/s foot per secondft

2 square foot

ft3 cubic foot

G giga (billion)

g gram

g/L gram per litre

g/t gram per tonne

Gal Imperial gallon

gpm Imperial gallons per minute

gr/ft3 grain per cubic foot

gr/m3 grain per cubic metre

ha hectare

hp horsepowerhr hour

HRIA Heritage Resource Impact Assessment

in inch

in2 square inch

J joule

k kilo (thousand)

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26-8

.

kcal kilocalorie

kg kilogram

km kilometre

km/h kilometre per hour

km2 square kilometre

kPa kilopascal

kVA kilovolt-amperes

kW kilowatt

kWh kilowatt-hour

L litre

L/s litres per second

LSA Local Study Area

M mega (million)

m metre

micron

m2 square metre

m3 cubic metre

m3 /h cubic metres per hour

MASL metres above sea level

MBCA Migratory Birds Convention Act

microgram

min minute

mm millimetre

MMER Metal Mining Effluent Regulations

MOE Saskatchewan Ministry of the Environment

mph miles per hourMVA megavolt-amperes

MW megawatt

MWh megawatt-hour

OPEX/CAPEX Operating Cost / Capital Cost

opt, oz/st ounce per short ton

oz Troy ounce (31.1035g)

Pb lead

PEA Preliminary Economic Assessment

ppm part per million

psia pound per square inch absolute

psig pound per square inch gauge

RL relative elevations second

st short ton

stpa short ton per year

stpd short ton per day

t metric tonne

TOR Terms of Reference

t/a metric tonne per year

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26-9

.

t/d metric tonne per day

TSF Tailings Storage Facility

US$ United States dollar

USg United States gallon

USgpm US gallon per minute

V volt

VMS Volcanogenic Massive Sulphide

W watt

wmt wet metric tonne

yd3 cubic yard

yr year

Zn zinc

Source: JDS 2014

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27-1

.

27 REFERENCES

ALS Metallurgy Kamloops. Metallurgical Testing on Samples from the Halilağa Deposit KM3897,

26th September 2013

AGN International, Tax Guide: Brief Summary of Turkish Tax System, Updated for 2010

Arancibia, O.N. and Clark, A.H., 1996, Early magnetite-amphibolite-plagioclase alteration-

mineralization in the Island Copper porphyry copper-gold-molybdenum deposit, British Columbia:

Economic Geology, v. 91, p. 402-438.

Bozkurt, E., On the Structural Geology of the Halilağa Project, Biga Peninsula, Western Turkey,

2007.

Ceyhan, N., Kilic, D., and Kose, G., Halilağa Property Exploration 2009-2010 Year End Report,

2011.Cook, S., Evaluation of ALS Chemex Etili Sample Prep Lab, Turkey. Teck Cominco Internal

Memorandum, May 17 2007.

Cook, S., and Houle, M. Analytical, Assay and QA/QC Protocols – Halilağa Cu-Au Project, Western

Turkey. Teck Cominco File Note, October 28 2009.

Cunningham-Dunlop, I.R., NI 43-101 Technical Report on the Halilağa Exploration Property,

Çanakkale, Western Turkey, 2011.

Einaudi, M., Mapping Altered & Mineralized Rocks; The “Anaconda Method”, 1997.

Einaudi, M., Report on Visit to Halilağa District Çanakkale Province (With Emphasis on Kestane

Porphyry Cu-Au Prospect), Western Anatolia, Turkey, December 2007.

Ernst and Young, Concise Tax Guide for Turkey 2009


Halilağa Project KM2020, 11th July 2007;

G&T Metallurgical Services Ltd., Mineralogy Testing on Samples from the Halilağa Project

KM2157, 1st April 2008;

G&T Metallurgical Services Ltd., Metallurgical Testing KM2908, 15th April 2011;

JDS Turkey Site Visit Notes April 7 – 14, 2014; Prepared for Pilot Gold Inc. by JDS Energy &Mining Inc., April 29, 2014.

JDS Halilağa Project Summary of Optimized PEA Case; Prepared for Pilot Gold Inc. by JDS

Energy & Mining Inc., April 30, 2014.

JDS Halilağa Project Summary of High-grade Case; Prepared for Pilot Gold Inc. by JDS Energy &

Mining Inc., April 30, 2014.

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27-2

.

Gray, J., Kirkham, G., Technical Report titled “Resource Estimate for the Halilağa Copper-Gold

Property NI 43-101 Technical Report “ March 23, 2012

Grieve, P., Brief Comment on the Geology and Mineralization of the Halilağa Prospect, Biga

Peninsula, Western Turkey, June 2006.

Grieve, P., NI 43-101 Technical Report on the Pirentepe and Halilağa Exploration Properties,

Çanakkale, Western Anatolia, Turkey, March 2007.


Halilağa Project, July 2007.

G&T Metallurgical Services Ltd., Metallurgical Testing Halilağa Project, August 2011.Gustafson,

L.B. and Hunt, J.P., 1975, The porphyry copper deposit at El Salvador, Chile: Economic Geology,

v. 70, p. 857-912.

Heberlein, D., Halilağa Project; An Audit of Quality Control Procedures and Data for Teck

Resources Inc., July 2012

Hedenquist, J., Observations on the Halilağa porphyry plus lithocap-hosted epithermal prospects

near Etili, Biga peninsula, Republic of Turkey, June 2007.

Hedenquist, J.W., Arribas, A., Jr., and Gonzalez-Urien, E., 2000, Exploration for epithermal gold

deposits: Reviews in Economic Geology, v. 13, p. 245-277.

Kuscu, G., Petrographical Analysis and Interpretation of the samples: 708975, 708977, 708983,

708984, 708985, 708989, 708991, 708997, and 708998, 2008.

Meinert, L.D., 2000, Gold in skarns related to epizonal intrusions: Reviews in Economic Geology, v.

13, p. 347-375.Mesta, B., Ceyhan, S., Başer, O., SRK Danışmanlık ve Mühendislik A.Ş.: Project No: 241001,

Muratlar Water Supply Project, 2011

Ross, K., Petrographic Study of the Halilağa Copper-Gold Porphyry Deposit, Turkey, April 2007.

Sillitoe, R.H., 2010, Porphyry copper systems: Economic Geology, v. 105, p. 3-41.

Preliminary Economic Assessment Technical Report for the Halilağa Project, Turkey; Prepared for

Truva Bakır Maden İşletmeleri A.Ş. and PilotGold Inc. by SRK Consulting (Canada) Inc., October

10, 2012.

Yiğit, O., 2012, A prospective sector in the Tethyan Metallogenic Belt: Geology and geochronology

of mineral deposits in the Biga Peninsula, NW Turkey: Ore Geology Reviews, p.118-148.

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27-1

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APPENDIX A

QP CERTIFICATES

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Jdsmining.ca


T 604.687.7545 f 604.689.5041 t 250.763.6369 f 250.763.6302

#860

–

625

Howe

Street

Vancouver,

BC

V6C

2T6

#200

– 532

Leon

Avenue,

Kelowna,

BC

V1Y

6J6

CERTIFICATE OF AUTHOR

I, Gordon Doerksen do hereby certify that:

1. I am currently employed as V.P. Technical Services with JDS Energy & Mining Inc. with an office atSuite 860 – 625 Howe Street, Vancouver, BC, V6C 2T6;

2. This certificate applies to the technical report titled “Revised Preliminary Economic AssessmentTechnical Report Halilağa Project, Turkey”, with an effective date of December 20, 2014, (the“Technical Report”) prepared for Pilot Gold Inc.(“the Issuer”);

3. I am a Professional Mining Engineer (P.Eng. #32273) registered with the Association of ProfessionalEngineers, Geologists of British Columbia. I am also a registered Professional Mining Engineer inAlaska, Wyoming and Yukon Territory. I am a Member of the Canadian Institute of Mining andMetallurgy and a Registered Member of the Society of Mining Engineers of the AIME.

I am a graduate of Montana Tech with a B.Sc. in Mining Engineering (1990). I have been involved inmining since 1985 and have practiced my profession continuously since 1990. I have held senior mineproduction and mine technical positions in mining operations in Canada, the US and in Africa. I haveworked as a consultant for over eight years and have performed mine planning, project management,cost estimation, scheduling and economic analysis work, as a Qualified Person, for a significantnumber of engineering studies and technical reports many of which were located in Latin America.I have read the definition of "qualified person" set out in National Instrument 43-101 (NI 43-101) andcertify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for thepurposes of NI 43-101.

4. I visited the Halilağa Project site from February 29 to March 1, 2012, April 26 to May 2, 2013 and fromApril 7 to 14, 2014;

5. I am responsible for Section numbers 1 to 6 and 17 to 27 (except 17.20, 17.21 and 17.22) of theTechnical Report;

6. I am independent of the Issuer and related companies applying all of the tests in Section 1.5 of theNI 43-101;

7. I have had prior involvement with the property that is the subject of the Technical Report and was aQualified Person for the 2012 Technical Report titled “Preliminary Economic Assessment TechnicalReport for the Halilağa Project, Turkey” prepared by SRK Consulting (Canada) Inc. with an effectivedate of August 27, 2012 and executed on October 10, 2012;

8. I have read NI 43-101, and the Technical Report has been prepared in compliance with NI 43-101 andForm 43-101F1.

9. As of the effective date of the Technical Report and the date of this certificate, to the best of myknowledge, information and belief, this Technical Report contains all scientific and technicalinformation that is required to be disclosed to make the Technical Report not misleading;

Effective Date: December 20, 2014Signing Date: February 16, 2015

ORIGINAL SIGNED AND SEALED

Gordon Doerksen, P.Eng.

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Jdsmining.ca


T 604.687.7545 f 604.689.5041 t 250.763.6369 f 250.763.6302

#860

–

625

Howe

Street

Vancouver,

BC

V6C

2T6

#200

– 532

Leon

Avenue,

Kelowna,

BC

V1Y

6J6


I, Stacy Freudigmann do hereby certify that:

1. I am currently contracted as a Project Manager with JDS Energy & Mining Inc. who has an office atSuite 860 – 625 Howe Street, Vancouver, BC, V6C 2T6;


3. I am a Professional Engineer (P.Eng. License #33972) registered with the Association of ProfessionalEngineers, Geologists of British Columbia. I am a Member of the Canadian Institute of Mining andMetallurgy and the Australasian Institute of Mining and Metallurgy.

I am a graduate of James Cook University with a B.Sc.(Hons) in Industrial Chemistry (1996) and CurtinUniversity, Western Australia School of Mines with a Grad.Dip. Metallurgy (1999). I have beeninvolved in mining since 1996 and have practiced my profession continuously since 1996. I have heldsenior process and metallurgical production and technical positions in mining operations in Canadaand Australia. I have worked as a consultant for over five years and have performed processmanagment, project management, cost estimation, scheduling and economic analysis work for anumber of engineering studies and technical reports located in Latin America, USA and Canada.I have read the definition of "qualified person" set out in National Instrument 43-101 (NI 43-101) andcertify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for thepurposes of NI 43-101.

4. I have not visited the Halilağa Project;

5. I am responsible for Section number 16 of the Technical Report;


7. I have had prior involvement with the property that is the subject of the Technical Report.





Stacy Freudigmann, P.Eng.

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Jdsmining.ca


T 604.687.7545 f 604.689.5041 t 250.763.6369 f 250.763.6302

#860

–

625

Howe

Street

Vancouver,

BC

V6C

2T6

#200

– 532

Leon

Avenue,

Kelowna,

BC

V1Y

6J6


I, Dino Pilotto do hereby certify that:

1. I am currently employed as Mine Engineering Lead with JDS Energy & Mining Inc. with an office atSuite 860 – 625 Howe Street, Vancouver, BC, V6C 2T6;

2. This certificate applies to the technical report titled “Revised Preliminary Economic AssessmentTechnical Report Halilağa Project, Turkey”, with an effective date of December 20, 2014, (the“Technical Report”) prepared for Pilot Gold Inc.(“the issuer”);

3. I am a Professional Mining Engineer (P.Eng. #14782) registered with the Association of ProfessionalEngineers, Geologists of Saskatchewan. I am also a registered Professional Mining Engineer inAlberta and Northwest Territories.

I am a graduate of the University of British Columbia with a B.Sc. in Mining and Mineral ProcessEngineering (1987). I have practiced my profession continuously since June 1987. I have beeninvolved with mining operations, mine engineering and consulting covering a variety of commodities atlocations in North America, South America, Africa, and Eastern Europe.

I have read the definition of "qualified person" set out in National Instrument 43-101 (NI 43-101) andcertify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for thepurposes of NI 43-101.

4. I visited the Halilağa Project site from February 29 to March 1, 2012;

5. I am responsible for Section 15 (excluding 15.1) of the Techical Report;

6. I am independent of the Issuer and related companies applying all of the tests in Section 1.5 of the Nl43-101;

7. I have had prior involvement with the property that is the subject of the Technical Report and was aQualifed Person for the 2012 Technical Report titled “Preliminary Economic Assessment TechnicalReport for the Halilağa Project, Turkey” prepared by SRK Consulting (Canada) Inc. with an effectivedate of August 27, 2012 and executed on October 10, 2012;





Dino Pilotto, P.Eng.

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Jdsmining.ca


T 604.687.7545 f 604.689.5041 t 250.763.6369 f 250.763.6302

#860

–

625

Howe

Street

Vancouver,

BC

V6C

2T6

#200

– 532

Leon

Avenue,

Kelowna,

BC

V1Y

6J6


I, Maritz Rykaart do hereby certify that:

1. I am currently employed as a Practice Leader and Principal Consultant with SRK Consulting (Canada)Inc. with an office at Suite 2200 – 1066 West Hastings Street, Vancouver, BC, V6E 3X2;


3. I am a Professional Engineer registered with the Association of Professional Engineers &Geoscientists of British Columbia (APEGBC #28531); I am also a registered Professional MiningEngineer in Saskatchewan, Yukon Territory and Nunavut/Northwest Territories. I am member of theCanadian Geotechnical Society and the Canadian Dam Association.

I am a graduate of the Rand Afrikaans University in 1991 and 1993 with B.Eng. (Civil) and M.Eng.(Civil) degrees respectively. In 2001 I graduated with a PhD (Geotechnical Engineering) from theUniversity of Saskatchewan. I have practiced my profession continuously since graduation in 1993,except for a 3 year break from 1998 to 2001 to complete my PhD. My work experience is relatedalmost entirely to the mining industry, specifically the design, construction, monitoring and closure ofmine waste facilities including tailings impoundments.

4. I visited the Halilağa Project site from I have personally inspected the subject project on February 29 toMarch 2, 2012;

5. I am responsible for Section numbers 17.20 to 17.22 of the Technical Report;

6. I am independent of the Issuer and related companies applying all of the tests in Section 1.5 of the

NI 43-101;

7. I have had prior involvement with the property that is the subject of the Technical Report and was aQualified Person for the 2012 Technical Report titled “Preliminary Economic Assessment TechnicalReport for the Halilağa Project, Turkey” prepared by SRK Consulting (Canada) Inc. with an effectivedate of August 27, 2012 and executed on October 10, 2012;





Maritz Rykaart, P.Eng.

7/24/2019 2015 Halilaga


CERTIFIC I, Michael

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7/24/2019 2015 Halilaga



I, Gary Simmons do hereby certify that:

1. I am currently the Owner of GL Simmons Consulting, LLC, with an office at 105 Chapel Road, ClydePark, MT 59018.


3. I am a Qualified Professional (QP) Member with special expertise in Metallurgy, QP No. 01013QP,registered with the Mining and Metallurgical Society of America (MMSA). I am also a RegisteredMember of the Society for Mining, Metallurgy and Exploration of the SME, Member ID 2959300.

I am a graduate of the Colorado School of Mines with a B.Sc. in Metallurgical Engineering (1972). Ihave been invlolved in the Mining business since 1974 and have practiced my profession continuouslysince 1974. I have held senior mine and metallurgical production and corporate level management,technical and development positions for mining companies with opertions in the United States,Canada, Australia, Indonesia, Peru and Mexico. I have worked as an independent consultant since

2008.

I have read the definition of "qualified person" set out in National Instrument 43-101 (NI 43-101) andcertify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for thepurposes of NI 43-101.

4. I visited the Halilağa Project site from April 26 to May 2, 2013;

5. I am responsible for Section numbers 13) of the Technical Report;


7. I have had no prior involvement with the property that is the subject of the Technical Report;





Gary L Simmons, QP

7/24/2019 2015 Halilaga


Advantage Geoservices Limited

1051 Bullmoose Trail

Osoyoos, BC V0H 1V6


I, James N. Gray do hereby certify that:

1. I am a consulting geologist with Advantage Geoservices Limited with an office at Osoyoos, BC;

2. This certificate applies to the technical report titled “Revised Preliminary Economic AssessmentTechnical Report Halilağa Project, Turkey”, with an effective date of December 20, 2014, (the“Technical Report”) prepared for Pilot Gold Inc. (“the Issuer”);

3. I am a Professional Geoscientist, registered and in good standing with the Association of ProfessionalEngineers and Geoscientists of British Columbia (#27022);

I graduated from the University of Waterloo in 1985 where I obtained a B.Sc in Geology. I havepracticed my profession continuously since 1985. My experience includes resource estimation work atoperating mines as well as base and precious metal projects in North and South America, Europe,Asia and Africa;

I have read the definition of "qualified person" set out in National Instrument 43-101 (NI 43-101) andcertify that by virtue of my education, affiliation with a professional association and past relevant workexperience, I fulfill the requirements to be a "qualified person" for the purposes of NI 43-101.

4. I have not personally visited the Halilağa Project site;

5. I am responsible for Section 14 for the Technical Report;

6. I am independent of the Issuer and related companies applying all of the tests in Section 1.5 of NI 43-101;

7. I have had prior involvement with the property that is the subject of the Technical Report and was aQualified Person for the 2012 Technical Report titled “Preliminary Economic Assessment TechnicalReport for the Halilağa Project, Turkey” prepared by SRK Consulting (Canada) Inc. with an effectivedate of August 27, 2012, as well as the 2012 Technical Report titled “Resource Estimate for theHalilağa Copper-Gold Property” prepared by Advantage Geoservices Ltd. and Kirkham GeosystemsLtd. with an effective date of March 23, 2012;


9. As of the effective date of the Technical Report and the date of this certificate, to the best of myknowledge, information and belief, this Technical Report contains all scientific and technicalinformation that is required to be disclosed to make the Technical Report not misleading.


Original Signed and Sealed

James N. Gray, P.Geo.

Documents

2015 Halilaga