Amulsar Resource Update and Heap Leach ... - The Gold Report

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Amulsar Resource Update and Heap Leach ... - The Gold Report

LYDIAN INTERNATIONAL LTD.

Amulsar Resource Update

and

Heap Leach Feasibility Study

PREPARED FOR:

LYDIAN INTERNATIONAL, LTD.

Ground Floor, Charles House

Charles Street

St Helier, Jersey JE2 4SF

Channel Islands

PREPARED BY:

Mr. Joseph M. Keane, P.E.

Mr. Richard Kiel, P.E.

Mr. Pete Lemke, P.E.

Mr. Herb Welhener, MMSA-QPM

Mr. John Eyre, FRICS MIMMM MIQ CEnv

K D Engineering

7701 N. Business Park Drive

Tucson, Arizona 85743

Document No. Q439-04-028-01

Project No. 439-04

3 September 2012

KDE FORM No. A263a-7/12/99


September 5, 2012

113-81597FS

Ms. Kim Batorson

KD Engineering & Metcon Research

7701 North Business Park Drive

Tucson, Arizona 85743-9622

RE:

CERTIFICATE OF AUTHOR - RICHARD E. KIEL

Dear Ms. Batorson:

As a co-author of this Technical Report (Amulsar Resource Update and Heap Leach Feasibility Study) on

the Amulsar Project for Lydian International, Ltd., Toronto, Quebec, Canada, I, Richard E. Kiel, do hereby

certify that:

1. I am an Associate, and carried out this assignment, for Golder Associates Inc., 44 Union

Boulevard, Suite 300, Lakewood, Colorado 80228, USA, tel. (303) 980-0540, fax

(303) 985-2080, e-mail rkiel@golder.com.

2. I hold the following academic qualifications:

A. B.Sc. (Geological Engineering), South Dakota School of Mines & Technology, USA,

1976-1979.

3. I am a registered Member of the Society for Mining, Metallurgy, and Exploration (SME).

4. I am a registered professional civil engineer in California, Nevada, Colorado, and

Wyoming.

5. I have worked as a civil and geological engineer in the minerals industry for 21 years.

6. I am familiar with NI 43-101 and, by reason of education, experience, and professional

registration; I fulfill the requirements of a Qualified Person as defined in NI 43-101. My

work experience includes 19 years as a consulting engineer on precious metals, base

metals, and rare earth oxides, and 2 years as a geologist and engineer on an operating

uranium mine. I have an additional 10 years of experience in a related industry (e.g.,

solid and hazardous waste management). I am qualified to prepare and review the

engineering for the heap leach facility, waste dump facility, mine closure and reclamation,

and for geotechnical engineering aspects of the Amulsar project.

7. I have visited the property three times: in June 2011, September/October, 2011, and

again in May 2012.

8. This is the second Technical Report I have co-authored on the mineral property in

question.

9. As of the date of this certificate, to the best of my knowledge, information, and belief, the

Technical Report contains all scientific and technical information that is required to be

disclosed to make this report not misleading.

I:\11\81597FS\0100\11381597FS LTR CertOfAuthorAmulsar-RKiel_FS 05SEP12.docx

Golder Associates Inc.

44 Union Boulevard, Suite 300

Lakewood, CO 80228 USA

Tel: (303) 980-0540 Fax: (303) 985-2080 www.golder.com

Golder Associates: Operations in Africa, Asia, Australasia, Europe, North America and South America

Golder, Golder Associates and the GA globe design are trademarks of Golder Associates Corporation


Ms. Kim Batorson September 5, 2012

KD Engineering & Metcon Research 2 113-81597FS

Sincerely,

10. I am responsible for the preparation of the Technical Report for the Heap Leach Facility,

Waste Dump Facility, Geotechnical Engineering for Plant Facilities, and the Preliminary

Closure and Rehabilitation Plan, as discussed in Sections 17 (17.2) and 18 (18.2.4,

18.2.5, 18.4 Introduction), the geotechnical portions of Section 21 (21.5 and 21.6),

Section 24 and the updated Section 5.

GOLDER ASSOCIATES INC.

Richard E. Kiel, P.E.

Senior Geological Engineer

cc:

Brad Schwab, KDE

I:\11\81597FS\0100\11381597FS LTR CertOfAuthorAmulsar-RKiel_FS 05SEP12.docx


September 7, 2012

113-81597FS

TO:

British Columbia Securities Commission

Alberta Securities Commission

Autorité des marches financiers

TSX Venture Exchange

RE:

CONSENT OF RICHARD E. KIEL

LYDIAN INTERNATIONAL, LTD.

I, Richard E. Kiel, P.E., Independent Civil and Geological Engineer, do hereby consent to the public filing

of the report titled, “Lydian International, Ltd., Amulsar Resource Update and Heap Leach Feasibility

Study,” dated 3 September 2012 (the “Technical Report”) that supports the disclosure in the press

release of Lydian International, Ltd., dated 5 September 2012 (the “Press Release”). I consent to the use

of extracts from, or a summary of, the Technical Report in the Press Release.

I confirm that I have read the Press Release and that it fairly and accurately represents the information in

the Technical Report.

Dated this 7 th day of September 2012.

GOLDER ASSOCIATES INC.

Richard E. Kiel, P.E.

Senior Geological Engineer

cc:

Brad Schwab, KDE

I:\11\81597FS\0100\11381597FS LTR ConsentToFileAmulsar-RKiel_FS 07SEP12.docx

Golder Associates Inc.

44 Union Boulevard, Suite 300

Lakewood, CO 80228 USA

Tel: (303) 980-0540 Fax: (303) 985-2080 www.golder.com

Golder Associates: Operations in Africa, Asia, Australasia, Europe, North America and South America

Golder, Golder Associates and the GA globe design are trademarks of Golder Associates Corporation


Sepember 5, 2012

113-81597FS

Ms. Kim Batorson

KD Engineering & Metcon Research

7701 North Business Park Drive

Tucson, Arizona 85743-9622

RE:

CERTIFICATE OF AUTHOR - PETER R. LEMKE

Dear Ms. Batorson:

As a co-author of this Technical Report (Amulsar Resource Update and Heap Leach Feasibility Study) on

the Amulsar Project for Lydian International, Ltd., Toronto, Quebec, Canada, I, Peter R. Lemke, do

hereby certify that:

1. I am the Water Treatment Technical Lead, and carried out this assignment, for Golder

Associates Inc., 44 Union Boulevard, Suite 300, Lakewood, Colorado 80228, USA, tel.

(303) 980-0540, fax (303) 985-2080, e-mail peter_lemke@golder.com.

2. I hold the following academic qualifications:

A. B.Sc. (Chemical Engineering), Rose-Hulman Institute of Technology, USA,

1977-1981.

B. M.Sc. (Ecological Engineering), Colorado School of Mines, USA, 1986-1989.

3. I am a registered professional environmental engineer in Colorado.

4. I have worked as an environmental engineer on remediation and industrial

water/wastewater treatment projects for 22 years.

5. I am familiar with NI 43-101 and, by reason of education, experience, and professional

registration; I fulfill the requirements of a Qualified Person as defined in NI 43-101. My

work experience includes 22 years as a consulting engineer for environmental

remediation and industrial wastewater treatment projects. Previous experience includes

laboratory research in alternative fuels, and industrial production. I am qualified to

prepare and review the engineering for the wastewater management and treatment

engineering aspects of the Amulsar project.

6. This is the first Technical Report I have co-authored on the mineral property in question.

7. As of the date of this certificate, to the best of my knowledge, information, and belief, the

Technical Report contains all scientific and technical information that is required to be

disclosed to make this report not misleading.

I:\11\81597FS\0100\11381597FS LTR CertOfAuthorAmulsar-PLemke_FS 05SEP12.docx

Golder Associates Inc.

44 Union Boulevard, Suite 300

Lakewood, CO 80228 USA

Tel: (303) 980-0540 Fax: (303) 985-2080 www.golder.com

Golder Associates: Operations in Africa, Asia, Australasia, Europe, North America and South America

Golder, Golder Associates and the GA globe design are trademarks of Golder Associates Corporation


Ms. Kim Batorson Sepember 5, 2012

KD Engineering & Metcon Research 2 113-81597FS

Sincerely,

8. I am responsible for the preparation of the Technical Report for the Wastewater

Treatment Plant design, Sections 18.2.5 and 21.7.

GOLDER ASSOCIATES INC.

Peter R. Lemke, P.E.

Water Treatment Technical Lead

cc:

Brad Schwab, KDE

I:\11\81597FS\0100\11381597FS LTR CertOfAuthorAmulsar-PLemke_FS 05SEP12.docx


7 September 2012 113-81597FS

TO:

British Columbia Securities Commission

Alberta Securities Commission

Autorité des marches financiers

TSX Venture Exchange

RE:

CONSENT OF PETE LEMKE

LYDIAN INTERNATIONAL, LTD.

I, Peter R. Lemke, P.E., Water Treatment Technical Lead, Golder Associates Inc., do hereby consent to

the public filing of the report titled, “Lydian International, Ltd., Amulsar Resource Update and Heap Leach

Feasibility Study,” dated 3 September 2012 (the “Technical Report”) that supports the disclosure in the

press release of Lydian International, Ltd., dated 5 September 2012 (the “Press Release”). I consent to

the use of extracts from, or a summary of, the Technical Report in the Press Release.

I confirm that I have read the Press Release and that it fairly and accurately represents the information in

the Technical Report.

Dated this 7 th day of September 2012.

GOLDER ASSOCIATES INC.

Pete Lemke, P.E.

Water Treatment Technical Lead

cc:

Brad Schwab, KDE

I:\11\81597FS\0100\11381597FS LTR ConsentToFileAmulsar-PLemke_FS 07SEP12.docx

Golder Associates Inc.

44 Union Boulevard, Suite 300

Lakewood, CO 80228 USA

Tel: (303) 980-0540 Fax: (303) 985-2080 www.golder.com

Golder Associates: Operations in Africa, Asia, Australasia, Europe, North America and South America

Golder, Golder Associates and the GA globe design are trademarks of Golder Associates Corporation


Lydian International - Amulsar Resource Update and Heap Leach Feasibility Study

Section

TABLE OF CONTENTS

Page

1.0 SUMMARY ............................................................................................... 1

2.0 INTRODUCTION ........................................................................................... 16

3.0 RELIANCE ON OTHER EXPERTS ............................................................... 18

4.0 PROPERTY DESCRIPTION AND LOCATION .............................................. 19

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES,

INFRASTRUCTURE AND PHYSIOGRAPHY ................................................ 24

6.0 HISTORY ............................................................................................... 27

7.0 GEOLOGICAL SETTING AND MINERALIZATION ..................................... 29

8.0 DEPOSIT TYPES .......................................................................................... 49

9.0 EXPLORATION ............................................................................................. 50

10.0 DRILLING ............................................................................................... 56

11.0 SAMPLE PREPARATION, ANALYSIS AND SECURITY ............................. 60

12.0 DATA VERIFICATION .................................................................................. 63

13.0 MINERAL PROCESSING AND METALLURGICAL TESTING .................... 85

14.0 MINERAL RESOURCE ESTIMATES ........................................................... 114

15.0 MINERAL RESERVE ESTIMATES ............................................................... 142

16.0 MINING METHODS ....................................................................................... 152

17.0 RECOVERY METHODS ................................................................................ 183

18.0 INFRASTRUCTURE ..................................................................................... 194

19.0 MARKET STUDIES AND CONTRACTS ...................................................... 212

K D Engineering Document No. Q439-04-028-01 3 September 2012

KDE FORM No. A263a-7/12/99


Lydian International - Amulsar Resource Update and Heap Leach Feasibility Study

TABLE OF CONTENTS (Continued)

Section

Page

20.0 ENVIRONMENTAL STUDIES, PERMITTING AND

SOCIAL OR COMMUNITY IMPACT ..................................... 213

21.0 CAPITAL AND OPERATING COSTS ........................................................... 256

22.0 ECONOMIC ANALYSIS ................................................................................ 282

23.0 ADJACENT PROPERTIES ........................................................................... 289

24.0 OTHER RELEVANT DATA AND INFORMATION ........................................ 290

25.0 INTERPRETATION AND CONCLUSIONS ................................................... 294

26.0 RECOMMENDATIONS ................................................................................. 297

27.0 REFERENCES .............................................................................................. 301

28.0 APPENDICES ............................................................................................... 303

Appendix 1 - Design Criteria

Appendix 2 - Equipment List

Appendix 3 - Capital Cost Detail

Appendix 4 - Drawings

K D Engineering Document No. Q439-04-028-01 3 September 2012

KDE FORM No. A263a-7/12/99


Lydian International - Amulsar Resource Update and Heap Leach Feasibility Study

List of Tables

Table 1.1 Summary of Amulsar geochemical samples 3

Table 1.2 Amulsar Mineral Resource 4

Table 1.3 Mineral Reserves Represent the Undiluted Ore Scheduled to the Crusher 4

Table 1.4 KCA - Column Tests 8

Table 1.5 Heap Leach Metal Recovery by Deposit 8

Table 1.6 Total Initial And Future Sustaining Project Costs 13

Table 1.7 Cash Operating Cost 13

Table 1.8 Economic Highlights 14

Table 1.9 Economic Analysis Summary - US$ Pre-Income Tax Cash Flow 14

Table 1.10 Summary of Key Financial Parameters (Sensitivity to Gold Price) 15

Table 9.1 Summary of Amulsar Geochemistry Samples 53

Table 9.2 Summary of Amulsar Geochemical Samples 53

Table 10.1 Selected Down Hole Gold Intercepts for the 2011 Amulsar Drilling Program 58

Table 12.1 2010 Assay QA/QC Summary 64

Table 12.2 ALS Data Base Assays Versus Check Assays for Gold 80

Table 12.3 Summary of Assays on Gold Standards 80

Table 13.1 Screened Metallics Analysis for Gold, Comp 1 85

Table 13.2 Head Analyses of Composite 1 86

Table 13.3 Coarse Bottle Roll Leach Test Summary 87

Table 13.4 Test Work Results Summary 88

Table 13.5 Minus 75 µm Bottle Roll Leach Tests 89

Table 13.6 Minus 2 mm Bottle Roll Leach Tests 89

Table 13.7 Minus 38 mm Column Leach Tests 89

Table 13.8 Minus 19 mm Column Leach Tests 90

Table 13.9 Final Gold Recovery Summary by Test and Composite 91

Table 13.10 Column Leach Test Results Summary 92

Table 13.11 WAI – Bulk Density and Specific Gravity 94

Table 13.12 WAI – Bulk Density and Specific Gravity 94

Table 13.13 Coarse Ore Bottle Roll Leach Test Results (Gold) 95

Table 13.14 Coarse Ore Bottle Roll Leach Test Results (Silver) 96

Table 13.15 WAI – Rock Type Column Tests 99

Table 13.16 WAI - Rock Type Column Tests 99

Table 13.17 KCA - Composite Samples For Head Analyses and Leach Testing 100

Table 13.18 KCA - Composite Samples for Physical Test Work 100

Table 13.19 KCA - Abrasion and Work Indices 101

Table 13.20 KCA – Gold and Silver Head Assays 101

Table 13.21 KCA - Copper, Mercury, Carbon and Sulfur Assays 102

Table 13.22 KCA - Bottle Roll Tests 103

Table 13.23 KCA - Column Tests 104

Table 13.24 KCA – Gold Head Assays 108

Table 13.25 KCA - Silver Head Assays 108

Table 13.26 KCA – Column Detoxification – Cyanide Species 109

Table 13.27 KCA – Column Detoxification – Reagents 109

Table 13.28 Heap Leach Design Parameters 112

Table 13.29 KCA - Column Test Leach Kinetics 112

Table 13.30 Heap Leach Metal Recovery by Deposit 113

Table 14.1 Amulsar Mineral Resource 115

Table 14.2 Drilling and Assaying Statistics 116

Table 14.3 Assay and 10 meter Bench Composite Statistics 118

K D Engineering Document No. Q439-04-028-01 3 September 2012

KDE FORM No. A263a-7/12/99


Lydian International - Amulsar Resource Update and Heap Leach Feasibility Study

List of Tables (Continued)

Table 14.4 Gold Variogram Ranges, meters 119

Table 14.5 10 m Composite Gold Statistics by Lithology Code 126

Table 14.6 IMC Model Contained Resources at a 0.20 g/t Gold Cut-off for

Indicated Class 127

Table 14.7 Amulsar Mineral Resource for All Areas 137

Table 14.8 Amulsar Mineral Resource by Classification and Deposit 138

Table 14.9 Model Contained Mineralization, All Deposits 139

Table 14.10 Model Contained Mineralization in Erato 140

Table 14.11 Model Contained Mineralization in Tigranes 140

Table 14.12 Model Contained Mineralization in Artavasdes 140

Table 14.13 Sensitivity of Resource to Estimation Approach 141

Table 14.14 Sensitivity of Resource to Vertical Search Distance 141

Table 15.1 Floating Cone Inputs 143

Table 15.2 NPV of Floating Cone Geometries Evaluated at $1,200/oz Au and $20/oz Ag 148

Table 15.3 Material Contained within Floating Cone Geometries 149

Table 15.4 Mineral Reserves Represent the Undiluted Ore Scheduled to the Crusher 151

Table 16.1 Phase Design Criteria 152

Table 16.2 Comparison of Designed Phase Tonnes Against $900 Cone Tonnes

at a .25 g/t Recovered Gold Cut-off 153

Table 16.3 Phase Tonnages and Grades at a 0.25 g/t Recoverable Gold Cut-off 154

Table 16.4 Material Movements Total Annual Summary 159

Table 16.5 Waste Movement Required for Mine Schedule 162

Table 16.6 Equipment Requirements by Time Period 180

Table 16.7 Salaried Staff Labor Requirements 181

Table 16.8 Mine Hourly Labor Requirements 182

Table 18.1 Summary of Operations Personnel 197

Table 18.2 Mine Power Requirements (by Area) 200

Table 20.1 RoA Permits Required for Development of Amulsar Mine 220

Table 20.2 Average Water Requirements 223

Table 20.3 Summary of findings from the Environmental Impact Assessment 235

Table 20.4 Summary of Social Impacts 241

Table 20.5 Environmental, Community and/or Health & Safety Design

Protection Measures and Best Management Techniques 249

Table 21.1 Summary of Mine Capital Costs ($US x 1000) 256

Table 21.2 Summary of Mine Operating Costs - Total Dollars ($US x 1000) 259

Table 21.3 Summary Process Plant - Initial Capital Cost 260

Table 21.4 Summary Process Plant - Sustaining Capital Cost 261

Table 21.5 Process Plant Operating Cost Estimate Summary 271

Table 21.6 Operating Cost Estimate - Heap Leach Power & Energy 272

Table 21.7 Operating Cost Estimate - Plant Labor (Phase I) 273

Table 21.8 Operating Cost Estimate - Plant Labor (Phase II) 274

Table 21.9 Operating Cost Estimate - Yerevan Office Administration Labor 275

Table 21.10 Operating Cost Estimate - Site Administration Labor 276

Table 21.11 Operating Cost Estimate - Heap Leach Consumables 277

Table 21.12 Maintenance 278

Table 21.13 Water 278

Table 21.14 Waste Dump Facility Cost Estimate, US$ 278

Table 21.15 Heap Leach Pad Cost Estimate, US$ 279

Table 21.16 Water Treatment Plant Cost Estimates 280

Table 21.17 Closure and Reclamation Cost Estimate 280

K D Engineering Document No. Q439-04-028-01 3 September 2012

KDE FORM No. A263a-7/12/99


Lydian International - Amulsar Resource Update and Heap Leach Feasibility Study

List of Tables (Continued)

Table 22.1 Owner Operated Mining - Economic Analysis Summary 282

Table 22.2 Owner Operated Mining - Economic Analysis Summary -

Before Tax Cash Flow and Unit Values 282

Table 22.3 Cash Flow Schedule 283

Table 22.4 Rate of Return Sensitivity 286

Table 22.5 NPV Sensitivity (US$ X 1000) 287

Table 22.6 Summary of Key Financial Parameters (Sensitivity to Gold Price) 288

K D Engineering Document No. Q439-04-028-01 3 September 2012

KDE FORM No. A263a-7/12/99


Lydian International - Amulsar Resource Update and Heap Leach Feasibility Study

List of Figures

Figure 1.1 Mine Plan 5

Figure 4.1 Location of Amulsar Gold Project 20

Figure 5.1 Physiography and Infrastructure in the Vicinity of the Amulsar Licenses 25

Figure 7.1 Eocene to Oligocene Arc Deposits 30

Figure 7.2 Regional Geology at Amulsar 31

Figure 7.3 Amulsar Geology Map 33

Figure 7.4 Classic transition contact of the PA unit DDA 034. 34

Figure 7.5 A, B: Grey and yellow argillaceous PA from DDA036. 35

Figure 7.6 A, B: Example from DDA 003 of silicified Lava flow-breccia facies 36

Figure 7.7 A, B: Polymict volcaniclastic horizon from DDA 150. 37

Figure 7.8 Other Lithologies. 39

Figure 7.9 Vesicular ‘scoriaceous like’ horizons from DDA 056. 40

Figure 7.10

Geological model long section running NE-SW from Tigranes

through to Artavasdes 43

Figure 7.11 Erato Geology Map 45

Figure 7.12 Spatial relationship between the volcaniclastics, faults and grade 47

Figure 7.13 Geological Model vs Gold Grade 48

Figure 9.1 Location map of IP/resistivity surveys at Armulsar 51

Figure 9.2 (15A left above)Plot of Ground Magnetics over the Central Amulsar

Prospecting License, (15B right above) Plot of The Spectrometer data. 52

Figure 9.3 Amulsar Drilling 2007-2011 54

Figure 10.1 Drilling Through End of 2011 59

Figure 12.1 Diamond drilling Laboratory Duplicates 67

Figure 12.2 Laboratory RC drilling Duplicates 68

Figure 12.3 Core Hole Gold Grade Paired with RC Hole Gold Grade within 50 m, 70

Figure 12.4 2011 Gold Grade Paired with Pre-2011 Gold Grade within 50 m 71

Figure 12.5 ALS vs. Alfred H. Knight Check Assays 72

Figure 12.6

ALS vs. Alfred H. Knight Check Assays, Samples Prepared at ALS,

Pre-2011 Drilling 73

Figure 12.7 ALS vs. Alfred H. Knight Check Assays 74

Figure 12.8 ALS Data Base Assays Versus Acme Check Assays, 2011 Drilling 75

Figure 12.9 ALS Data Base Assays Versus Duplicate Assays, All Drilling 76

Figure 12.10 ALS Data Base Assays Versus Check_au, All Holes 77

Figure 12.11 ALS Data Base Assays Versus Check_au, Hole RCA-301 (High-Grade) 78

Figure 12.12 ALS Data Base Assays Versus Check_au, Holes DDA-035, 047, 120, 133

(Low Grade) 79

Figure 12.13 Standard G904-8, Grade 5.53 g/t 81

Figure 12.14 Standard G900-6, Grade 2.56 g/t 82

Figure 12.15 Standard G307-2, Grade 1.08 g/t 82

Figure 12.16 Standard GLG304-1, Grade 0.153 g/t 83

Figure 12.17 Blank Assays 84

Figure 13.1 Gold Leach Curves 87

Figure 13.2 Column Leach Curves (-38 mm) 90

Figure 13.3 Column Leach Curves (-19 mm) 91

Figure 13.4 Column Leach Curves 93

Figure 13.5 Coarse Bottle Roll Leach Recoveries (-12 mm) 97

Figure 13.6 Effect of Head Grade on Gold Leach Recovery 98

K D Engineering Document No. Q439-04-028-01 3 September 2012

KDE FORM No. A263a-7/12/99


Lydian International - Amulsar Resource Update and Heap Leach Feasibility Study

List of Figures (Continued)

Figure 13.7 KCA Columns - Gold Extraction versus Time 105

Figure 13.8 KCA Columns – Gold Extraction versus kl/t 106

Figure 13.9 KCA Columns – Gold Extraction Correlation 107

Figure 13.10 Metallurgical Drill Hole Locations 110

Figure 14.1 Drill Hole Location Map 117

Figure 14.2 Gold and Silver Probability Plots, 10 m Composites 118

Figure 14.3(A) Omnidirectional Covariance Variogram, Gold Erato 120

Figure 14.3(B) Omnidirectional Covariance Variogram, Gold Tigranes 121

Figure 14.3(C) Omnidirectional Covariance Variograms, Gold Artavasdes 122

Figure 14.3(D) Omnidirectional Covariance Variograms, Gold All Data 123

Figure 14.4 10m Drill Hole Composites Exceeding 1 g/t gold, Tigranes & Artavasdes 124

Figure 14.5 Gold Grade-Thickness Product, 0.1 g/t Cut-off, Tigranes & Artavasdes 125

Figure 14.6 Gold Probability Plots, Indicated Blocks Only 127

Figure 14.7 10 m Composite Gold Grades, 2850 Bench, Tigranes & Artavasdes 129

Figure 14.8 IMC Model Gold Grades, 2850 Bench, Tigranes & Artavasdes, Indicated 130

Figure 14.9 10 m Composite Gold Grades, 2800 Bench, Erato 131

Figure 14.10 IMC Model Gold Grades, 2800 Bench, Erato, Indicated 132

Figure 14.11 Kriging Variance Versus Number of Holes in Search Ellipsoid 133

Figure 14.12 Mineral Resource Classification, 2850 Bench, Tigranes & Artavasdes 134

Figure 14.13 Grade-thickness (gram-meters) of Measured-Indicated Blocks 135

Figure 14.14 Grade-thickness (gram-meters) of Inferred Blocks, All Deposits 136

Figure 15.1 $900/oz Floating Cone used for Ultimate Pit Design 145

Figure 15.2 US$ 400-US$ 1200/oz Floating Cones Sliced at 2830m Elevation 146

Figure 15.3 Cross Sections of US$ 400, US$ 600, and US$ 900/.oz Au Cones 147

Figure 15.4 Results of Floating Cone Evaluations 149

Figure 15.5 Ultimate Pit 150

Figure 16.1 Phases Sliced at 2830 m elevation 155

Figure 16.2 Cross Sections of Designed Phases Showing Gold Grade in Block Model 156

Figure 16.3 Graphical Presentation of Mine Schedule 160

Figure 16.4 Pit Backfill at End of Mine Life 161

Figure 16.6 Proposed Stockpiles at the end of Year 10 163

Figure 16.7 End of Pre-Production 165

Figure 16.8 End of Year 1 166

Figure 16.9 End of Year 2 167

Figure 16.10 End of Year 3 168

Figure 16.11 End of Year 4 169

Figure 16.12 End of Year 5 170

Figure 16.13 End of Year 6 171

Figure 16.14 End of Year 7 172

Figure 16.15 End of Year 8 173

Figure 16.16 End of Year 9 174

Figure 16.17 End of Year 10 175

Figure 16.18 End of Year 11 176

Figure 16.19 End of Year 12 177

Figure 17.1 Amulsar Overall Flowsheet 184

Figure 18.1 Proposed Overall Site General Arrangement Layout 203

Figure 20.1 Project Environs showing proposed infrastructure and regional setting

(rivers, villages topography) 216

Figure 20.2 Footprint of Mine Development (throughout the operational life) 222

K D Engineering Document No. Q439-04-028-01 3 September 2012

KDE FORM No. A263a-7/12/99


Lydian International - Amulsar Resource Update and Heap Leach Feasibility Study

List of Figures (Continued)

Figure 20.3

State Sanctuaries and Important Bird Areas in relation to the

Project Exploration License 232

Figure 22.1 Amulsar Gold Project Pre-Tax Sensitivity IRR 286

Figure 22.2 Amulsar Gold Project Pre-Tax Sensitivity NPV@5% 287

K D Engineering Document No. Q439-04-028-01 3 September 2012

KDE FORM No. A263a-7/12/99


Lydian International - Amulsar Resource Update and Heap Leach Feasibility Study

List of Photos

Photo 20.1 Typical Landscape at the Project 215

K D Engineering Document No. Q439-04-028-01 3 September 2012

KDE FORM No. A263a-7/12/99


Lydian International - Amulsar Resource Update and Heap Leach Feasibility Study

Glossary

Bulk Leach Extractable Gold .............................................................. Bleg

Canadian Institute of Mining ............................................................... CIM

Celsius .................................................................................... C

Certified Reference Material ............................................................... CRM

Centimeter .................................................................................... cm

Cubic Meters .................................................................................... m 3

Day .................................................................................... d

Degree .................................................................................... °

Diamond Core Drill Hole ..................................................................... DDH

Feasibility Study ................................................................................. FS

Gram .................................................................................... g

Grams per ton ................................................................................... g/t

High Density Polyethylene ................................................................... HDPE

Greater than .................................................................................... >

Hectare .................................................................................... ha

Hour .................................................................................... h

Hours per day .................................................................................... h/d

Intermediate Leach Solution ................................................................ ILS

Kilogram .................................................................................... kg

Kiloliter .................................................................................... kl

Kilometer .................................................................................... km

Kilovolt .................................................................................... kV

Less than .................................................................................... <

Life of Mine .................................................................................... LOM

Liter .................................................................................... l

Linear Low Density Polyethylene ......................................................... LLDPE

Megawatt .................................................................................... MW

Meter .................................................................................... m

Metric ton (tonne) ................................................................................ t

Millimeter .................................................................................... mm

Micrometer .................................................................................... µm

Million .................................................................................... M

Million tonnes .................................................................................... Mt

Million tonnes per year (annum) .......................................................... Mtpa

Mineral Resource Estimate ................................................................. MRE

Square Kilometers .............................................................................. km 2

Square Meters .................................................................................... m 2

Synthetic Precipitation Leaching Process ............................................ SPLP

Troy Ounce .................................................................................... oz

Parts per million ................................................................................. ppm

Percent .................................................................................... %

Potentially Mineable Mineralization .................................................... PMM

Pregnant Leach Solution ..................................................................... PLS

Volt .................................................................................... V

Year (annum) .................................................................................... a

K D Engineering Document No. Q439-04-028-01 3 September 2012

KDE FORM No. A263a-7/12/99


Lydian International - Amulsar Resource Update and Heap Leach Feasibility Study Page 1

1.0 SUMMARY

Lydian International LTD. (Lydian) has engaged Independent Mining Consultants

(IMC), K D Engineering (KDE) and Golder Associates (Golder), independent

engineering firms, to prepare a Resource Update and Feasibility Study (FS) of the

Amulsar heap leach facility. The study was prepared to standards pursuant to Canada’s

National Instrument 43-101 (NI 43-101) and was based on the following:







KDE site visits conducted in May 2011 and January 2012 and feasibility level

design and cost estimates for the processing plant.

Geological review, mineral resource estimates, mineral reserve estimates and

mine capital and operating cost estimates completed by Independent Mining

Consultants, Inc. (IMC).

Heap leach facility (HLF), waste dump facility (WDF), and wastewater

treatment plant (WWTP) feasibility level designs and cost estimates and a

preliminary mine closure and reclamation (C&R) plan and cost estimate by

Golder.

Integrated water studies (i.e, groundwater, surface water, freshwater ecology,

and water impacts), pit slope stability, seismicity and geochemical studies by

Golder.

Metallurgical testing by SGS Lakefield, SGS Cornwall, Wardell Armstrong

International (WAI), and Kappes, Cassiday & Associates (KCA).

The draft of the Environmental and Social Impact Assessment and a number

of baseline studies performed by WAI, Environmental Resources

Management (ERM), Shape Consulting and other specialized Armenian and

international independent consultants, as well as other focus studies

performed by Golder, Radman Associates, EJ Acoustics, SKM and etc.

1.1 Geology, Exploration and Resource Estimation

Armenia is located in the Lesser Caucasus mountain range which runs in a

northwest-southeast trend. The regions geology is complex due to the accretion of

exotic terranes through plate-tectonic processes, and to ongoing tectonic activity and

volcanism.

The main lithologies of interest at Amulsar are Eocene-Oligocene volcanic and

volcaniclastic rocks which were formed in response to a calc-alkaline magmatic-arc

system that extended NW through southern Georgia and SE into the Alborz-Arc of Iran.

The near-shore continental arc was located between the southern margin of the

Eurasian plate and the northern limit of the Neo-Tethyan Ocean. Subduction ceased in

the early Oligocene when a fragment of the continental crust collided at the trench axis

and accreted with the Eurasian plate (Lydian Geology report).

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The Amulsar deposits are located in a structurally complex zone proximal to an

interpreted caldera structure. Two dominant structure trends are observed running

through the area: NE-SW dominantly normal faults and NW-SE left-lateral strike slip

faults. The locations of the volcanic domes, which are responsible for the Amulsar

deposits, align with these NW regional tectonic structures. The principal host rocks are

related to the formation and collapse of volcanic domes and include flow-banded

andesitic lava flows and lava flow breccias, and volcaniclastic deposits. The high level

of magmatism and fault activity has been responsible for the development of the multistage

phreatic and hydrothermal breccias, hydrothermal alteration and gold

mineralization.

Gold mineralization at Amulsar is a late event in the development of the deposit,

thus the spatial distribution is controlled by the preceding alteration and structural

events. Gold occurs within limonitic zones after the oxidation of sulfide material, and

three dominant controls on gold mineralization have been defined:




The argilic altered (kaolinized) porphyritic andesite formed an impermeable

barrier to mineralizing fluids, and caused ‘ponding’ of higher grade

mineralization along the contacts, often appearing as thick, leached and

‘gossanous’ zones.

Faults and fractures provided conduits for the gold mineralizing fluids,

resulting in high grade ‘gossanous’ veins. These often form broad corridors of

numerous, closely spaced high-grade structures.

Porous and permeable lithological units (hydrothermal breccias, volcaniclastic

breccias, more fractured units and occasional leached/vuggy volcanics)

allowed the lateral migration of mineralizing fluids away from structurally

controlled conduits forming large tonnages of lower grade material.

Since commencing work in 2006, Lydian has carried out diverse exploration

activities at Amulsar including; geological mapping, various geophysical surveys,

surface sampling and drilling. A total of 739 holes have been drilled at Amulsar between

2007 and 2012, including, exploration, geotechnical, metallurgical, and sterilization

holes. Reverse circulation (RC) and diamond drilling (DD) have both been carried out

on the property.

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Table 1.1

Summary of Amulsar geochemical samples

Diamond Drilling RC Drilling Total

Project Year # of

# of

# of

meters

meters

meters

holes

holes

holes

Amulsar 2007 5 593 0 0 5 593

Amulsar 2008 18 2679.3 74 10363 92 13042.3

Amulsar 2009 8 1089.8 101 12446 109 13535.8

Amulsar 2010 52 7575.9 129 17036 181 24611.9

Amulsar 2011 198 16113.2 154 23099 352 39212.2

Total 2007-2011 281 28051.2 458 62944 739 90995.2

The Amulsar deposits of Tigranes, Artavasdes and Erato are comprised of

fracture controlled, discontinuous, higher grade mineralization surrounded by halos of

lower grade mineralization. The nature of this style of mineralization requires tight

control of the higher grade drill intercepts in order to not give them more influence on

the overall grade estimate than they represent.

The mineral resource is based on drilling completed up to and through the 2011

drill season and includes a total of 665 core and reserve circulation (RC) holes. Gold

and silver grades in the Amulsar model were estimated using inverse distance to the

eighth power (ID8), spherical search ellipsoids of 50 meters in Tigranes and Artavasdes

and 100 meters in Erato and no internal domain boundaries.

The mineral resource is summarized in Table 1.2 at several cutoff grades, with

the resource at a 0.40 g/t gold cutoff being the tonnage and grade that was selected to

be reported. Higher and lower cutoff grades are presented to show the distribution of

tonnage and grades. The mineral resource is within a floating cone geometry based on

US$ 1300/oz gold price (no credit for silver) and preliminary estimates of gold recovery

and operating costs provided in January 2012. Additional definition of these estimates

were subsequently done and used for the definition of the mineral reserve. The

reported mineral resource represents about 98 percent of the model contained

mineralization.

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Au

Cutoff

g/t

Mt

Measured

Au

g/t

Ag

g/t

Table 1.2

Amulsar Mineral Resource

Sum of

Indicated

Measured &

Indicated

Au Ag

Au Ag

Mt

Mt

g/t g/t

g/t g/t

Mt

Inferred

0.75 15.4 1.64 4.97 13.0 1.57 4.78 28.4 1.61 4.88 15.0 1.45 4.56

0.50 27.7 1.18 4.18 23.9 1.13 4.14 51.6 1.16 4.16 26.8 1.08 4.39

0.40 36.5 1.00 3.82 32.2 0.95 3.84 68.8 0.98 3.83 35.5 0.92 4.01

0.30 50.2 0.83 3.45 46.4 0.77 3.48 96.6 0.80 3.47 48.1 0.77 3.65

0.20 70.7 0.66 3.11 71.1 0.59 3.13 141.9 0.62 3.12 73.9 0.59 3.12

Au

g/t

Ag

g/t

The mineral reserve was established by tabulating the undiluted tonnes and

grades of proven and probable material within the designed final pit geometry that is

scheduled as ore to the crusher over the mine life. A floating cone algorithm

(independently verified by Whittle optimizations) was used to determine the final pit

design and internal phase designs. The final pit design is based on the shell generated

by the US$ 900/oz floating cone as a result of the evaluation of the discounted net value

at US$ 1200/oz gold and US$ 20/oz silver prices for a suite of cone geometries run from

$400/oz to $1200/oz gold. The cones above US$ 900/oz. showed no increase in

contained value for the additional material mined. This is also a function of the

estimation being data limited as the cone at US$ 900/oz captures ore up to where

drilling is limited and insufficient drill data exists to classify material as either measured

or indicated.

Category

Table 1.3

Mineral Reserves Represent the Undiluted Ore Scheduled to the Crusher

Ore

kt

Contained Recoverable Contained Recoverable

Gold

g/t

Silver

g/t

Gold

g/t

Silver

g/t

Gold

oz

Silver

oz

Gold

oz

Silver

oz

Proven 51,143 0.801 3.37 0.713 1.31 1,317,000 5,541,000 1,172,000 2,154,000

Probable 37,106 0.789 3.43 0.694 1.17 941,000 4,092,000 828,000 1,393,000

Proven plus

Probable

88,249 0.796 3.40 0.705 1.25 2,258,000 9,633,000 2,000,000 3,547,000

1.2 Mining

Mining of the Amulsar deposit is planned to be accomplished with conventional

open pit mining methods. Over 12 years, 7 phases covering the Artavasdes, Tigranes

and Erato ore bodies are sequenced to arrive at an ultimate pit geometry containing the

project’s reserve. Mineralization extends to the surface in the Tigranes ore body where

initial mining begins; as a result, minimal pre-stripping of 729,000 tonnes is required to

have adequate ore feed to the crusher. Artavasdes and Tigranes areas are mined

ahead of the Erato area which requires more waste stripping to expose the ore.

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During the initial 3 years of mining, ore is scheduled from the pit as direct feed to

the crusher at a rate of 5 million tonnes of ore per year. In Year 3, crusher capacity is

doubled with a crusher expansion and 10 million tonnes of ore per year are sent to the

crusher starting in Year 4. The average stripping ratio in the first 3 years of mining is

1.8:1 waste:ore. Beginning in Year 4, the stripping ratio increases to 2.35:1 and

continues at that ratio to Year 10.

A graphical representation of the mine plan is given in Figure 1.1. It depicts

waste tonnes mined, ore tonnes sent to the crusher and recoverable ounces of gold

placed on the heap.

Figure 1.1 Mine Plan

As exploration is not as advanced at the Erato deposit a larger proportion of

material in the pit is classified as inferred. This contributes to the lower ounce profile in

Years 7 through 9 as the schedule is based on measured and indicated resources.

A small low grade stockpile is generated near the crusher in Year 2 of mining.

This is scheduled to be fed to the crusher in Years 10 and 12 of the mine life. The low

grade stockpile contains 655 ktonnes of ore which amounts to a little over 3 weeks of

ore at a crushing rate of 10 million tonnes per year.

During the first 8 years of mine life, waste is hauled to the waste dump facility

which is about 4.5 kilometers north of the mining area. Starting in Year 9, the mine plan

takes advantage of the opportunity to backfill completed pits. This has two benefits of

decreasing the haul distance and reducing the cost of reclamation.

Mine mobile equipment has been selected to meet the production requirements

of the mine schedule generated for Amulsar. Blast holes will be drilled with Sandvik

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DP1500i drills. Loading and hauling of material from the pit will be accomplished with a

mixed fleet of Cat 6018’s (RH90) and 6030’s (RH120) hydraulic excavators and Cat 777

haul trucks. An auxiliary fleet of D10 track dozers, 834 wheel dozers, a 16m Grader,

CAT 777 water truck, 992 front end loader and CAT 336 back-hoe are also required for

mining. These machines are planned for dump construction, road construction and

maintenance, pit cleanup and miscellaneous jobs.

The WDF will consist of the waste dump (WD), and an influent equalization basin

(IEB), wastewater treatment plant (WWTP), and evaporation pond (EP), located

downgradient of the WD and utilized for the collection and treatment of mine-influenced

water draining from the WD. A detailed discussion of the design of the WDF is provided

in the Feasibility Design Report (Golder, 2012g).

The WD will be constructed in three phases. The WD phase areas will be

465,500 m 2 , 506,800 m 2 and 360,100 m 2 for Phases 1, 2 and 3, respectively, for a total

WD area of 1,332,400 m 2 . Waste material will be deposited on the WD in nominal 8 m

thick lifts. The WD may be constructed in sub-phases to minimize initial capital costs

and to optimize water management within the IEB and flows to the WWTP.

The WD will be lined with a 0.45-m minimum thickness compacted lowpermeability

soil liner. An underdrain system will be constructed within the WD footprint

beneath the soil liner to drain groundwater/subsurface seepage to the IEB and prevent

the seepage from entering the waste pile above the WD base liner. Rainfall and

snowmelt water within the WD (contact water) will be collected by an overdrain system

constructed above the WD base liner and routed to the IEB.

The IEB was sized in accordance with the project design criteria to store the WD

underdrain and overdrain flows, and to provide flow control to the WWTP. The IEB will

have a composite liner system comprised of High Density Polyethylene (HDPE)

geomembrane underlain by a 0.3-m minimum thickness compacted low-permeability

soil liner.

The WWTP will receive water from the IEB. Treatment processes have been

developed based on the projected water quality characterization of the combined flows

from the WD underdrains and overdrains. The IEB and WWTP capacities have been

designed to accommodate high flows associated with snowmelt, with operation of the

WWTP at a constant rate for about eight months per year. Final treated effluent water

quality targets are to be determined. The WWTP effluent is projected to comply with

Armenian maximum allowable concentration (MAC) Category II standards. Category III

standards (more lenient) have been considered, but the conceptual design and cost

estimation for the WWTP is conservatively based on the more stringent Category II

effluent targets.

Reverse osmosis brine from the WWTP will drain by gravity to the EP for

evaporation. The EP was sized to meet the brine storage requirements. The final unit

operation in the wastewater treatment process is the spray-enhanced solar EP. Use of

the EP limits the operational season for water treatment. The EP will have a composite

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liner system comprised of HDPE geomembrane underlain by a 0.3-m minimum

thickness compacted low-permeability soil liner.

Treated water will be discharged to the Vorotan River. Secondary waste sludge

from chemical precipitation will be disposed on site. Solids accumulated in the EP may

be removed for disposal or disposed in-place at the end of the WWTP life. A detailed

discussion of the design of the WWTP is provided in the Wastewater Treatment

Feasibility Evaluation report (Golder, 2012f).

1.3 Metallurgy

Metallurgical testing on representative samples from the Amulsar project has

been on-going over the last five years with the most recent phase of testing conducted

at Kappes, Cassiday & Associates (KCA) of Reno, Nevada in late 2011 and early 2012.

This test program confirmed previous findings from four phases of SGS and WAI

metallurgical test work that indicate that the Amulsar ore types are amenable to

precious metal recovery by heap leaching.

The KCA test work studied fifteen representative samples from both the Tigranes

and Artavasdes deposits comprising bulk samples from surface outcrops and both

whole and half split core. The samples were used for physical tests including abrasion

and crushing work index. Head analyses included total and cyanide soluble gold, silver

and copper. Mercury, carbon, sulphur speciation and multi-element analyses were also

included. Coarse bottle roll tests were initially conducted to establish optimal test

parameters for the column leach tests.

Column tests were conducted at a crush size of 100 percent minus 12.5 mm

without agglomeration, at varying column heights from 2 to 4 meters, and for leach

cycles up to 75 days. The final column leach test results are presented in Table 1.4.

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Table 1.4

KCA - Column Tests

Deposit

Calculated Head

Reagent Consumption,

Sample – KCA

Extraction, %

Assay, g/t

kg/t

Sample Number

*

Au Ag Au Ag NaCN Ca(OH) 2

Tigranes Bulk - 61723 4.50 8.72 91 5 0.34 1.5

Artavasdes Bulk - 61724 0.67 4.11 89 48 0.12 1.0

Tigranes Split Core - 61730 0.56 0.52 96 34 0.18 2.0

Artavasdes Split Core - 61731 0.50 1.21 92 43 0.17 2.0

Artavasdes Split Core - 61732 0.95 0.76 93 22 0.15 2.5

Artavasdes Split Core - 61733 1.13 13.21 91 37 0.22 2.5

Tigranes Split Core - 61734 1.64 1.30 97 73 0.32 3.1

Tigranes Split Core - 61735 1.18 1.44 96 48 0.18 2.0

Tigranes Split Core - 61736 2.44 0.47 97 30 0.23 1.5

Tigranes Whole Core - 61768 1.27 1.35 92 20 0.14 2.0

Tigranes Whole Core - 61769 1.60 1.16 92 9


Lydian International - Amulsar Resource Update and Heap Leach Feasibility Study Page 9

In the first year of operation 3.75 Mtpa will be processed and 5 Mtpa in year two.

Installation of a duplicate crushing circuit ramps up production to 10 Mtpa in Year 4 till

the end of the life of the project.

Crushed ore will be transported approximately 3.5 kilometers on an overland

conveyor to be distributed along the north side of the leach pad. Pebble lime will be

added to the ore while on the overland conveyor. A tripper conveyor will deliver the ore

from the overland conveyor to a series of twenty four portable conveyors. A stacking

conveyor will place the ore on the leach pad in lifts of a nominal thickness of 8 meters.

The HLF will consist of a leach pad and collection ponds. The leach pad will be

constructed in three phases with the ultimate ore heap amount of 95 Mt stacked in three

stages. The pad phases will be expansions to the north and each phase will be divided

into two cells for a total of six cells for the ultimate pad. The Phase 1 pad area will be

479,690 m 2 and the Stage 1 heap capacity will be 18 Mt, suitable for the first 3.3 years

of operation. The Phase 2 pad area will be 465,000 m 2 and the Stage 2 heap capacity

will be 27 Mt to provide capacity through Year 6. The Phase 3 pad area will be 461,120

m 2 and the Stage 3 heap capacity will be 50 Mt. The ultimate pad area will be 1,405,810

m 2 and will accommodate 95 Mt of ore heap with a nominal maximum heap height of 72

m above the pad liner, suitable for the 11-year operating life. If additional leachable ore

is identified, a fourth pad phase may be constructed to the north of Phase 3 to allow the

stacking of up to 120 Mt of ore heap. The leach pad may be constructed in sub-phases

to further minimize initial capital costs. A detailed discussion of the design of the HLF is

provided in the Feasibility Design Report (Golder, 2012c).

The collection ponds consist of process (pregnant and intermediate) ponds and a

storm event (storm) pond sized in accordance with the project design criteria and

constructed down gradient of the leach pad. Additionally, an overflow pond will be

constructed down gradient of the storm pond. The process ponds were sized to contain

the operational and drain down flows, and the storm pond was sized to contain the

storm runoff from the ultimate pad and ponds.

The leach pad and collection ponds will have composite liner systems

respectively, comprised of Linear Low Density Polyethylene (LLDPE) and HDPE

geomembranes underlain by a 0.3-m minimum thickness compacted low-permeability

soil liner. The process ponds will be double-HDPE-geomembrane lined with an

intermediate leak collection and recovery system (LCRS) layer.

Pregnant leach solution (PLS), intermediate leach solution (ILS) and storm event

ponds will be located south of the leach pad, and a barren leach solution tank will be

located inside the Adsorption-Desorption-Regeneration (ADR) plant near the ponds.

Barren and intermediate leach solutions will be dosed to contain 0.5 gpl sodium cyanide

and will be applied by drip emitters to the top of the ore heap at irrigation rates of 10

l/h/m 2 . The drip emitter application system will operate to reduce evaporation in summer

and allow leaching to continue in winter. These leach solutions will be stacked such that

the barren solution will be used to irrigate the ore in a secondary leach cycle and the

intermediate solution will be used to irrigate fresh ore in a primary leach cycle to

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produce PLS. The primary and secondary leach cycles will be 30 days and t 80 days

respectively. Leaching of precious metal from the ore will continue as the leached ore is

buried by consecutive lifts. After 30 days of buried lift leaching, resulting in 140 total

leach days, the predicted overall recoveries shown in Table 1.2, will be attained with an

overall leach solution to ore ratio exceeding 3 m 3 /t.

The barren and intermediate leach solutions will percolate through the ore and be

collected in a network of perforated drain collection pipes installed within a granular

layer above the pad liner. The solution will gravity flow from the drain pipes via transfer

pipes exiting the pad and draining into the process ponds. The transfer pipes will direct

the solution to either the pregnant or intermediate ponds by valve control.

The PLS will be pumped from the pregnant pond into the ADR plant. Precious

metal will be adsorbed from solution onto activated carbon counter-currently in a train of

five adsorption columns. Carbon desorption and regeneration will occur daily in 7 tonne

batches. A second train of five carbon columns will be installed for the Phase II

expansion. Carbonate scale will be removed from batches of loaded carbon in an acid

wash vessel using dilute hydrochloric acid. Precious metal will be desorbed from the

acid washed carbon in a strip vessel operating under elevated temperature and

pressure. After the carbon is used it will be regenerated in a kiln. The strip solution will

report to an electrowinning circuit where precious metals will be deposited onto steel

mesh cathodes. Weekly, the deposited metals will be washed from the cathodes, dried

in a retort to volatilize and collect elemental mercury, if present, and smelted in a

furnace with fluxes.

The doré, containing roughly equal proportions of gold and silver, will be shipped

off-site for refining and sale.

1.5 Infrastructure

The Amulsar Gold Project covers an area of 130 km 2 , located in south central

Armenia. Currently paved roads are available to the town of Jermuk and a 15 km dirt

road is available from Jermuk to the mine site. Long term accommodation will be

provided to house up to 200 people on site and the remaining personnel will reside in

existing hotels in Jermuk. The contractor will be responsible for his own construction

camp. Currently a small exploration camp is available at site which utilizes a portable

generator.

There is good infrastructure surrounding the Amulsar project. This includes the

paved highway between Yerevan and Iran, high tension power lines and substations, a

gas pipeline from Iran, year round water from the Vorotan River and a fibre optic

internet cable. As a consequence of the project location on the top of a mountain ridge,

a reasonable amount of infrastructure will need to be constructed during project

development. Mobile phones work on most parts of the project area. Out of country

supplies, material and equipment can be shipped to the ports of Poti or Batumi,

Georgia, then trucked through Georgia and Armenia to the Amulsar project site.

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Community relations issues are currently handled by HSEC senior staff, a social

development manager, and a community liaison officer and a good understanding of

local issues and sensitivities has been established.

A detailed strategy for accommodating construction personnel, employees and

security personnel during the construction period will be developed during the detailed

engineering effort.

The Project is located in the catchments of three rivers Vorotan, Arpa and Darb.

The sections of catchments of Vorotan and Arpa rivers fall under the Lake Sevan Law

as the zones of non-immediate impact, where mining and processing are not restricted.

1.6 Environmental and Social Impact

The Environmental and Social Impact Assessment (ESIA) prepared for this

project considers the proposed mining of the Tigranes and Artavasdes deposits at

Amulsar, together with associated mine waste management, mineral handling, heap

leaching, gold extraction and ancillary activities, to produce gold-silver bullion

corresponding to Phases 1, 2 and 3 of the mining operations of a period of

approximately 10 years from the commencement of ore extraction, followed by

reclamation and closure of the mine. Environmental and social studies required with

respect to mining operations at Erato will require full assessment in an ESIA addendum

to be completed at a later stage of the Project.

The broad scope of the ESIA has considered the following:








The policy, legal and administrative framework.

The Project design covering geographical, ecological, environmental, social

and temporal aspects, influences and effects.

Analysis of baseline environment and socio-economic, defined through

detailed scope of work, in consultation with key stakeholders.

Review of alternatives for siting various project facilities, taking into account

regulatory, environmental, biodiversity, cultural properties, social and

community health and safety issues.

Defining the environmental and social impacts associated with Project

construction, operation and mine closure and reclamation.

Incorporating mitigation measures into detailed design to eliminate or

minimize impacts to an acceptable level and consider appropriate

alternatives.

Develop key Framework Management Plans with input to the site specific

Environmental and Social Action Plan (ESAP) for the delivery of the Project.

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Framework Mine Closure Plan and measures for post-mining management.

The findings of the draft ESIA concluded the following:



Environmental impacts range from Negligible to Major, in the absence of

mitigation. Through the implementation of detailed mitigation measures,

together with adherence to management plans in the ESAP, it is considered

that any potential residual environmental impacts can be reduced to a range

Negligible to Moderate. Residual environmental impacts include dust

generation, generation of greenhouse gas emissions, conversion of land

reducing habitat available for avian, flora and fauna species and changes to

landscape and topography. If left unmitigated, the project would have the

potential to cause surface and groundwater contamination, however

extensive mitigation measures have been developed which will manage any

potential impacts to an acceptable level.

Social impacts are both positive and negative. Positive impacts relate to

improvements in local livelihoods through direct employment by the project,

as well as knock-on economic growth; and macroeconomic benefits through

taxation, land rent and other revenues paid by Lydian. These positive impacts

range from Minor to Moderate; provided enhancement measures (such as

stakeholder engagement, transparency and governance) are

implemented. Negative impacts relate to economic displacement as a result

of land take; localized inflation driven by higher incomes in the region;

community cohesion issues between mine employees and other local

residents; potential in-migration of job-seekers; managed disturbance of

archaeological sites; and community health impacts, including higher risk

sexual practices. With mitigation, negative impacts range from Negligible to

Moderate.

The effective implementation of the mitigation measures outlined in the ESIA is

essential to ensure that positive benefits of the project outweigh the negative impacts

and the negatives are mitigated or designed out. The development and implementation

of detailed management plans will be incorporated into operational procedures, as well

as Lydian’s ESMS. Lydian’s social strategy and on-going community development

measures are expected to provide additional benefits to local communities over and

above project impacts.

1.7 Capital Cost

Capital costs for the project were estimated by IMC for mining, KDE for the

processing plant/infrastructure and Golder for the leach pad, collection ponds, waste

dump facility, wastewater treatment plant, and for mine closure and rehabilitation.

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The capital expenditures for the Amulsar Project processing facility will occur in

two phases; Years 1 to 3 is Phase I and Years 4 through the remaining years is Phase

II. Besides the Phase II process plant expansion sustaining costs are incurred for the

leach pad, mining fleet and waste dump.

The initial and sustaining capital costs are summarized in Table 1.6.

Table 1.6

Total Initial And Future Sustaining Project Capital Costs

Initial cost, Sustaining cost,

Item

US$

US$ *

Total,

US$

Mining Cost 8,791,700 17,189,800 25,981,500

Process Plant Cost 228,568,063 26,872,254 255,440,318

Waste Water Treatment Plant - 19,078,412 19,078,412

Leach Pads 15,687,450 31,814,488 47,501,938

Waste Dump 16,575,893 14,302,181 30,878,074

Closure and Reclamation 37,221,477 37,221,477

Total Initial and Future Sustaining Project Cost 269,623,106 146,478,612 416,101,718

* Sustaining costs include the majority of the capital costs associated with the Phase II

expansion.

1.8 Operating Costs

Operating costs for the project were estimated with input from KDE, IMC and

Golder. These costs over the life of the mine are summarized in Table 1.7.

Table 1.7

Life-of-Mine Cash Operating Cost

Item

US$/Tonne Ore

Mining 6.29

Processing 2.92

Waste Water Treatment Plant 0.13

G & A 0.47

Cash Operating Cost 9.81

Newmont Payment 0.21

Total Operating Cost (US$/oz) 468.48

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1.9 Economic Analysis

The economic highlights are summarized in Table 1.8.

Table 1.8

Economic Highlights

Average stacked gold grade g/t 0.750

Steady state annual gold production (Yr 1-3) oz 118,341

Steady state annual gold production (Yr 4-12) oz 186,047

Life of Mine from production start yr 12

Planned Steady State Production Rate (Yr 1-3) tpd 15,000

Planned Steady State Production Rate (Yr 4-12) tpd 30,000

IRR Pre tax % 27.7%

NPV Pretax (5% discount rate) US$M 646.0

Payback period from start of production yr 4.0

NPV Pretax (0% discount rate) US$M 1,121.6

Initial Capital Cost US$M 269.6

Total Capital Cost US$M 416.1

Cash Costs US$/oz 468.5

Metallurgical Gold Recovery % 88.6

Total Mined Gold to Leach Pad Moz 2.29

The financials for the base case mining options are summarized in Table 1.9.

Table 1.9

Economic Analysis Summary - US$ Pre-Income Tax Cash Flow

US $ x

1000

US $/t

Resource

US $/oz

Gold

Mine Gate Value of All Resource Net of

Transportation and Refining

2,424,680 25.55 1,194.75

Mining Operating Cost (596,959) (6.29) (294.15)

Processing Cost (277,116) (2.92) (136.55)

Waste Water Treatment Plant (12,276) (0.13) (6.05)

General & Administration (44,407) (0.47) (21.88)

Royalties (Newmont Payment) (20,000) (0.21) (9.85)

Cash Operating Cost (950,757) (10.02) (468.48)

Cash Operating Cash Flow 1,473,923 15.53 726.27

Capital Cost including Pre-Production Development (416,102) (4.38) (205.03)

Pre-Income Tax Cash Flow 1,057,821 11.15 521.24

Metal price scenarios were used in the pre-tax model to evaluate the sensitivity

on NPV, IRR, and payback. The results for the base case mining options are shown in

Table 1.10.

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Table 1.10

Summary of Key Financial Parameters (Sensitivity to Gold Price)

Gold Price, US$/oz 1,100 1,200 1,300 1,400 1,500

NPV(5), (000's) 512,504 645,976 779,448 912,920 1,046,392

IRR, Pre-Taxes 23.8% 27.7% 31.3% 34.8% 38.1%

Payback, Operating Years 4.5 4.0 3.7 3.4 3.1

1.10 Conclusions and Recommendations

Positive results from this feasibility study suggest that the Amulsar project be

advanced towards detailed engineering at an estimated cost of US$ 9.5 million.

Metallurgical testing results and process engineering confirm that the Amulsar

ores are amenable to precious metal recovery by tertiary crushing and cyanide heap

leaching. Additional studies should be conducted into eliminating the Phase 2 ramp up

with initial production at 10 Mtpa, comparing the installation of a single gyratory versus

two jaw crushers, as well as finalizing the crushing and process plant locations.

Mining at Amulsar will be by conventional open pit methods with 90-tonne haul

trucks. A study into improving economics by using 140-tonne trucks should be

conducted as well as project improvements by modeling the ore body on 5-meter blocks

instead of 10-meter blocks. Additional exploration upside at Amulsar has the potential to

increase the mineral resource.

A site wide water balance has been completed with the design management

plans to mitigate the discharge of contaminated water. Further analysis of the balance is

recommended to reduce treatment requirements.

Environmental baseline monitoring programs should be maintained and the

development of the Biodiversity Action Plan should be continued and implemented.

Social engagement should continue towards defining opportunities for local

employment. Preparation should continue towards preparation for implementation of

the Environmental and Social Action Plan.

The Engineering, Procurement and Construction Schedule should be optimized

and work should commence on a Project Execution Plan. Key personnel should be

hired including a construction manager familiar with in-country construction contractors.

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2.0 INTRODUCTION

This report forms an update to the report titled “Development of Amulsar Heap

Leach Facility, Preliminary Economic Assessment (PEA), Lydian International Limited,

43-101 Technical Report” dated 12 August 2011, prepared by Joseph M. Keane, P.E. of

K D Engineering (KDE), and filed on SEDAR, detailing preliminary project engineering,

cost and cash flow estimates. The PEA concluded that development of the Amulsar

resource is economically viable and recommended advancement of the project.

This Technical Report has also been prepared by Mr. Joseph M. Keane, P.E. on

the instruction of various executives of Lydian International Limited. The report includes

details of the updated MRE completed for the Amulsar Gold Project and preliminary

engineering design information for the proposed mine plan and mineral beneficiation

facilities along with capital and operating cost information. The report is written to

comply with the requirements of the National Instrument 43-101, “Standards of

Disclosure for Mineral Properties”, as part of Lydian’s ongoing continuous disclosure

obligations regarding the company’s exploration activities and property development.

The following individuals are Qualified Persons (QP) as defined by the CIM

Definition Standards November 22, 2005 and Section 5.1 of National Instrument 43-101

Standards of Disclosure for Mineral Projects, Form 43-101F1 and Companion Policy 43-

101CP.

Mr. Joseph M. Keane, P.E., of K D Engineering has responsibility for the report

contents and specifically Sections 2, 3, 4, 6, 13, 17, 18, 19 and 23 and the process

portion of 21, 22, and 27 and 28. Mr. Keane visited the property May 2011 and is the

Qualified Person for matters relating to the design and costs of the processing facility.

He has relied upon other experts for specific information in the report as mentioned

subsequently.

Mr. Richard Kiel, P.E., of Golder, visited the property in June, September and

October, 2011 and again in July, 2012, and is the Qualified Person for all matters

relating to the HLF and WDF portions of Sections 17 (17.2) and 18 (18.2.4, 18.2.5, 18.4

Introduction), the geotechnical portions of Section 21 (21.5 and 21.6), Section 24 and

the updated Section 5.

Mr. Pete Lemke, P.E., of Golder, is the Qualified Person for all matters relating to

the Wastewater Treatment Plant design, Sections 18.2.5 and 21.7.

Mr. Herb Welhener, Independent Mining Consultants, is the qualified person for

all matters relating to geology, drilling, open pit and mine design and is responsible for

Sections 7 through 12, 14 through 16, 18.2.1 and 21.1 and 21.2.

Mr. John Eyre, FRICS MIMMM MIQ CEnv visited the property in June 2010 and

2011, and is the Qualified Person for all matters relating to the Environmental and

Social Impact Assessment on behalf of WAI. Mr. Eyre is responsible for Section 20.

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All Qualified Persons contributed to Sections 1, 25 and 26.

This independent Report has been prepared following submittal of the

Preliminary Economic Evaluation (Document No. Q439-03-028-01) dated 12 August

2011. The following information has been utilized during the preparation of this report.




Process engineering designs completed in July 2011 to support this report

by KDE.

Engineering design reports and supplemental studies prepared by Golder

as referenced in Section 27.

Feasibility level capital and operating cost estimates produced by IMC,

KDE, and Golder.

▪ 2012 Mineral Resource Estimate, Amulsar Gold Project, NI 43-101

Technical Report, March 2012, prepared by IMC.

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3.0 RELIANCE ON OTHER EXPERTS

Subject to normal due diligence, KDE has relied on the accuracy of reports and

data supplied by Lydian and other geological and mineral engineering consultants in the

preparation of the Independent Report. KDE has reviewed and analyzed data provided

by Lydian and other geological and mineral engineering consultants, and has drawn its

own conclusions there-from, augmented by its direct field examinations. KDE has not

carried out any independent exploration work, drilled any holes or carried out sampling

or assaying on the property.

Relevant Joint Venture Agreements and exploration permit documents covering

the Lydian Project were viewed by CSA in three previously published NI 43-101

compliant documents, although full legal verification and due diligence of documents

was not undertaken.

The authors acknowledge the full cooperation of Lydian’s management and field

staff, all of whom made any and all data requested available and responded openly and

helpfully to all questions, queries and requests for material. All maps, as well as certain

of the Tables and Figures for this report were either supplied by Lydian or derived from

the documents listed in Section 2.

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4.0 PROPERTY DESCRIPTION AND LOCATION

The following has been extracted and updated from the NI-43-101 Preliminary

Economic Assessment entitled “Development of Amulsar Heap Leach Facility” dated

12 August 2011.

4.1 Summary

The Amulsar Gold Project covers an area of 130km 2 , located in south central

Armenia (Figure 4.1). The project area is covered by two Special Prospecting Licenses

(SPL), the Saravan license (29/043 formerly No. 42) and the Gorayk license (29/042

formerly No. 41). A Mining License (No 14/588) has also been granted, which covers

the Amulsar area. The current SPL’s were granted in August 2007, following the

conversion of the original Amulsar license (No 193) to a mining license. Five years is

allowed for exploration, with two extensions of two years allowed, before a mining

license application must be submitted.

The Armenian Ministry of Energy and Natural Resources approved an initial

C1+C2 reserve of 16.38 tonnes of gold metal in balance and 3.6 tonnes of gold offbalance

(not CIM compliant) for the Tigranes site at Amulsar on February 23, 2009.

Subsequently an application was submitted for a Mining License. Mining License 14/588

was granted over the Tigranes area on the 4th of April 2009.

Following further exploration, reserves were updated and on the 16th of

September 2011 new C1 + C2 reserves of 52.66 tonnes of gold metal and 240.51

tonnes of silver metal (not CIM compliant) were approved for the Tigranes and

Artavasdes sites at Amulsar. On November 22 nd 2011 Mining License No 14/588 was

then extended over the new reserve area and is now current and valid for 25 years at a

proposed mining rate of 2.6Mtpa.

Lydian is yet to conclude a new Mining Agreement for Mining License 14/588.

Progress on the Mining Agreement required EIA approval for the new open pit and

waste dump facility. EIA approval was granted on July 31 st 2012. All relevant documents

have now been submitted and the new Mining Agreement is expected in September

2012.

The licenses are held 100 percent by Geoteam CJSC (Geoteam), an Armenian

registered closed joint stock company. Geoteam is owned 95 percent by Lydian

Resources Armenia (a wholly owned subsidiary of Lydian International Ltd.). Lydian

entered into a put and call arrangement under which the terms for the acquisition of the

remaining 5 percent have been agreed (Lydian press release dated 10-12-2010).

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Figure 4.1 - Location of Amulsar Gold Project

Armenia’s mining law came into force in April of 2003 and was subsequently

amended in December of that year. This law was significantly modified in January 2007

when the National Assembly classified all metallic minerals as strategic and introduced

an auction or tender process for all Prospecting and Mining Licenses. This law was

abolished by the new Mining Code which came into effect on January 1, 2012. It defines

the principles and order of mining throughout the territory of the Republic of Armenia,

governs relations associated with the protection of nature and environment from

deleterious effects, ensures security of works during mining, as well as protection of

rights and legitimate interests of state and individuals during mining. There is also a five

year assurance on foreign investments wherein the investor can choose for that period

to operate under the previous or current legislation. The Gorayk and Saravan Special

Prospecting licenses (No’s 29/043 and 29/042) were awarded to Lydian at auction and

were granted under the 2007 mining law.

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The World Bank advised the Armenian Government on revisions to the mining

laws, including taxes and royalties, and a revised mining code was implemented as at

1st January 2012.

4.2 Armenian Minerals Legislation

4.2.1 License on Geological Explorations for the Purpose of Mining

Geological exploration for the purpose of mining requires a prospecting license

provided by the authority together with a signed geological exploration agreement. A

Prospecting License is an exclusive license granted for a term of three years which may

be extended for 3 consecutive times, each time not exceeding 2 years. Within 10 days

after receiving notification from the Authority on approval of a Prospecting License

application, the applicant shall be invited to execute a Geological Exploration

Agreement.

A Geological Exploration Agreement shall indicate:









The approved purpose of geological exploration,

A schedule of works by phases, including a time-table of geological

exploration, works and the expected period for the submission of obtained

exploration information for state submission,

Border coordinates of the license,

A duration of the contract according to the plan of geological exploration work,

The types of minerals, exploration and style of extraction permitted to the

mining operator,

A provision that where new types of minerals or metals, which were not

indicated before have been identified during exploration, respective

amendments in the mining right for exploration of minerals shall be

introduced,

Environmental measures in the manner prescribed by the legislation,

Terms and conditions of submitting interim and final reports about the course

of the exploration works to the authorized body.

4.2.2 Mining License

Permission for the use of minerals, a mining agreement, mining site allocation

and mine design subject to respective examination together constitute a Mining License

and the right to mine.

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A Mining License shall be provided for the term of mine operation which is based

on a technical review and on the economics of mining the deposit. A Mining License

shall not exceed the maximum term of 50 years. A person holding a Mining License

may at any moment, but not later than six months prior to the end of the period of

mining permission, apply to the authorized body for an extension to the mining term.

A Mining License is accompanied by a Mining Agreement which comes into force

once a new Mining License has received a positive Environmental Impact Assessment.

A Mining Agreement should provide:












4.2.3 Royalty Tax

Information on confirmed reserves of the deposit,

The types of minerals, and style of extraction permitted to the mining

operator,

Coordinates of the allocated mining site,

Duration of the contract according to the project,

An environmental management plan,

Procedure on the submission of reports and works supervision provided by

the Mining Code and other legislation,

Provisions regarding the calculation and payment of fees for the use of

minerals and for the environmental protection fund,

Provisions regarding responsibilities in regards to the socio-economic

development of local communities,

Provisions related to mine closure,

Conditions by which Mining Licenses and Mining Agreements may be

modified in-case of an increase in reserves; and

A mining agreement may contain other conditions regulating relationships of

the parties provided that these conditions shall not contradict legislation.

The Armenian Mining Royalty Tax was amended as from 1 st January 2012 and

comprises two components: a top line 4 percent charge on the revenue from gold and

silver sales; and, what amounts to a 12.5 percent charge on Earnings after Operational

Expenses but before certain other expenses – see below – most notably interest

charges. Both these components of the Royalty Tax are deducted along with interest

costs before corporation tax (20%) is charged.

The calculation of the royalty and payment is done on a quarterly basis.

Profitability is calculated by the following formula (R-C)/R, where;

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R - is Revenue from mined reserves sold

C - is Operational Expenses for the same period

For each 0.8 percent of profitability, the payable royalty is 0.1 percent of the

revenue amount (this equates to a 12.5% charge). Operational Expenses are all

expenses related to production with the exception of:




Capital expenses

Amortization and depreciation calculated for tangible and non-tangible assets

Financing cost including interests for loans etc.

The royalties are payable before 30th of the month following the reportable

period.

4.3 Armenia - Newmont Joint Venture Agreement Royalty

In April 2010 Lydian terminated its 50:50 Joint Venture with Newmont and

acquired full ownership of the Amulsar project via its 95 percent owned Armenian

subsidiary, Geoteam C.J.S.C. Under the agreement Lydian purchased Newmont’s

interests in the Joint Venture for a total consideration of US$ 15 million in pre-production

payments, additional post production payments and the issuance of 3 million ordinary

shares in Lydian on signing the agreement. The preproduction payment is paid in three

US$ 5 million tranches. The first tranche was paid on signing the agreement in April

2010, the second tranche due at the end of 2011 was paid in March 2012 with

US$ 100,000 interest and the final tranche is to be paid either at the end of 2012 or 90

days after completion of a FS and all permissions to move into production are in place,

whichever occurs earliest. Under the Agreement Lydian has a further 12 months to

make this payment after it becomes due with simple interest of 10 percent charged from

the due date.

Regarding the post production payment, Lydian can elect from three options - to

pay Newmont a single payment of US$ 15.6 million on production, to pay an on-going

3 percent NSR, or make 20 quarterly payments of US$ 1 million each.

4.4 Environmental Liabilities and Permits

There are no special environmental restrictions or known past liabilities in respect

to the Amulsar area. Lydian is required to operate under normal environmental terms

and conditions, as set out by the relevant Armenian authorities. Lydian has all the

necessary permits to undertake exploration work at Amulsar. Additional permits will be

required as the project advances from exploration to construction and onwards to

production. For details, refer to Section 20. Conditional EIA’s for the heap leach facility

and waste dump and mine areas were granted in 2012.

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5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES,

INFRASTRUCTURE AND PHYSIOGRAPHY

Following is an update of Section 5 extracted from the NI-43-101 Preliminary

Economic Assessment entitled “Development of Amulsar Heap Leach Facility” dated 12

August 2011.

5.1 Access

The Amulsar area is located some 170 km by sealed road and 15 km by

unsealed track to the southeast of Yerevan. The license area straddles the boundary

between Vayots-Dzor and Syunik provinces and incorporates part of the main highway

south from Yerevan into Iran.

5.2 Climate

The weather is highland continental with generally hot summers and cold winters.

Temperature varies significantly depending on altitude. Mean summer temperatures are

reported as 25ºC, with mean winter temperatures being around minus 4ºC. Annual

precipitation is low, with an average in the order of 700 mm. Snow falls across higher

ground during the winter months and can remain from early November through to late

March. Because of the altitude, Amulsar Mountain is snow covered for the winter

months. Currently access is generally possible only from March to November. Access

for heavy machinery is confined to the period from May to October/November. This

situation will continue until all season roads are constructed, snow plows are purchased,

and routine maintenance is ongoing. To account for adverse weather days during the

winter months mining activities have been restricted to 330 days per annum although

crushing and heap leaching will occur 365 days per year.

5.3 Resources and Infrastructure

Infrastructure in Armenia is generally well developed, with a good road and

power network except in more remote regions. A high-tension power line transects the

southern limits of the license (Figure 5.1). Most areas have mobile phone coverage. The

capital, Yerevan, has a high standard airport, with regular flights to many international

centers.

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Figure 5.1 - Physiography and Infrastructure in the Vicinity of the Amulsar Licenses

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5.4 Physiography

Armenia covers an area of 29,800 km², with much of the country being

mountainous. Elevations are more than 1,500 m in most places, rising to a maximum of

4,090 m. Most of the mountainous areas are covered by scrub vegetation, with some

forested zones. There are a number of fertile river valleys.

The core of the Amulsar area comprises a mountainous terrain, an approximately

7 km long northwest-southeast trending ridge which reaches a maximum altitude of

2,988 m. The mountainous area around Amulsar is undulating and, with vegetation

limited to wild grasses and isolated scrub, access is relatively straightforward outside of

the winter period.

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6.0 HISTORY

The following has been extracted from the NI-43-101 Preliminary Economic

Assessment entitled “Development of Amulsar Heap Leach Facility” dated 12 August

2011.

6.1 Armenian Mining and Exploration - General History

Due to its location within a tectonically active collision zone between the Arabian

and Eurasian plates, Armenia has been endowed with a range of mineral resources,

particularly large, porphyry-style copper-molybdenum deposits, as well as many

polymetallic and gold bearing vein systems.

Large scale metal production began in the early 19th century with the opening of

the Alaverdy and Kapan polymetallic mines. More recently, in the 1950’s, the Zangezur

Copper Molybdenum Combine, developed the world-class Kadjaran deposit in the south

of Armenia, which produces 3 percent of the world’s molybdenum output. The

dissolution of the Soviet Union, coupled with low metal prices, severely disrupted

Armenia’s mining industry in the 1990’s, but a new legislative framework and improved

market conditions led to a significant upturn over the last few years.

Metal production comes from the Kadjaran (Cronimet) and Agarak

(GeoProMining) copper/molybdenum porphyry deposits and the Kapan vein-type

polymetallic deposit (Dundee Precious Metals) in the south and the Shahumyan

polymetallic deposit in the north, although these were been adversely impacted by the

metal price slump of late 2008 and 2009. Gold deposits known to date are primarily

hosted in veins and brittle shear structures - the Zod gold mine (GeoProMining) in

eastern Armenia is an example of the latter, but gold is also present as an accessory

mineral in some polymetallic deposits.

Other foreign exploration and development companies active in Armenia include

Global Gold and Caldera Resources.

6.2 Amulsar History

The Amulsar region was initially identified by the Armenian Soviet Expedition in

1936 to1937 as an area of “secondary quartzite” which was deemed to host potential as

a silica resource. Work aimed at testing the potential of the silica resource commenced

in 1946 with the development of small-scale exploration adits. This work concluded that

the alunite content of the silica was too high (up to 25%) and that, as such, the project

was of no interest as a source of quality silica.

Further work in the early 1960’s identified the “secondary quartzite” as

metasomatic in origin, developed due to the replacement of intermediate composition

volcanic rocks (known regionally as the Amulsar Suite). Some 300 m of tunneling and

640 m 3 of trenching were also completed during this time, mostly on the north-eastern

side of the ridge.

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Testing of a bulk sample concluded that the silica was of sufficient quality for the

production of low grade glasses. Volumetric calculations made during this time estimate

indicated some 360 million tonnes of secondary quartzite rock at Amulsar.

Research work by the Soviet Expedition continued at Amulsar during the period

1979 to 1982. This work was focused principally on understanding and mapping the

alteration zonation across the area. Silica reserves at Amulsar were never entered onto

the Republic of Armenia State Balance and no further exploration or research work has

been conducted by the Soviet Expedition in the area since 1982.

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7.0 GEOLOGIC SETTING AND MINERALIZATION

This section is taken primarily from Amulsar Geological Model report (Dr. Dougal

Jerram-Director and Volcanologist, DougalEARTH Ltd, Dr. Kim Martelli-Volcanologist

and Mentor Demi-Resource Geologist-Lydian International, April 2012). It is worth

noting that a major 3D modeling exercise incorporating all previous lithological and

alteration mapping work and including recent detailed structural analysis is currently

underway and due to be complete in November 2012.

7.1 Regional Geology

Armenia is located in the Lesser Caucasus mountain range which runs in a

northwest-southeast trend. The regions geology is complex due to the accretion of

exotic terranes through plate-tectonic processes, and to ongoing tectonic activity and

volcanism.

The main lithologies of interest at Amulsar are Eocene-Oligocene volcanic and

volcaniclastic rocks which were formed in response to a calc-alkaline magmatic-arc

system that extended NW through southern Georgia and SE into the Alborz-Arc of Iran

(Figure 7.1). The near-shore continental arc was located between the southern margin

of the Eurasian plate and the northern limit of the Neo-Tethyan Ocean. Subduction

ceased in the early Oligocene when a fragment of the continental crust collided at the

trench axis and accreted with the Eurasian plate (Lydian Geology report).

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Figure 7.1 - Eocene to Oligocene Arc Deposits

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Figure 7.2 - Regional Geology at Amulsar

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However, the Eocene magmatic activity occurred after the onset of the collision

and consequently cannot be directly related to the oceanic subduction processes. It has

been suggested that the possibility of a slab retreat and break off allowed the rising of

the asthenospheric mantle, which could have heated the overthickened continental

crust resulting in volcanism (Sosson et al., 2010).

7.2 Property Geology

The Amulsar deposits are located in a structurally complex zone proximal to an

interpreted caldera structure. Two dominant structure trends are observed running

through the area: NE-SW dominantly normal faults and NW-SE left-lateral strike slip

faults. The locations of the volcanic domes, which are responsible for the Amulsar

deposits, align with these NW regional tectonic structures. However, specific features in

dome morphology and structure and the rocks that form them are determined by the

actual tectonomagmatic conditions under which they have been formed.

The principal host rocks are related to the formation and collapse of volcanic

domes and include flow-banded andesitic lava flows and lava flow breccias, and

volcaniclastic deposits. The high level of magmatism and fault activity has been

responsible for the development of the multi-stage phreatic and hydrothermal breccias,

hydrothermal alteration and gold mineralization.

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7.3 Volcanic Lithologies/Facies

Figure 7.3 - Amulsar Geology Map

Three main lithologies/facies were identified during core/RC chips logging and

were incorporated into the 3D geological model. The lithologies included:

1) Coarse porphyry andesite (massive grey/white/yellow - argillaceous),

2) Dome related Lava flow-breccia facies (porphyry andesite lava and lava flow

breccia - pink/red – siliceous) and

3) Volcaniclastic facies (volcaniclastics - primary and reworked). In addition,

some minor lithologies/facies were identified which have not been considered

in the geological model.

The geological complexity has been incorporated where possible, but due to the

model resolution the addition of all the individual units would add more, and perhaps

unnecessary, complexity to the model. Therefore where units are greater than five

meters and continue through more than a few sections they have been included into the

model.

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7.3.1 Massive Porphyritic Andesite Facies: (massive grey/white/yellow -

argillaceous)

Description

This unit is a massive medium- to coarse-grained feldspar-hornblende porphyry

andesite, which is mainly grey but also occurs as white/yellow and reddish coloration

(Figures 7.4 and 7.5). Most commonly the color variation is found in the contact zones.

The crystal size and density varies, but generally the phenocrysts range between 3 and

4mm.

Figure 7.4 - Classic transition contact of the PA unit DDA 034. (from right to left) Grey argillaceous

PA, shows yellow to red argillaceous contact transition into volcaniclastic section.

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Figure 7.5 - A, B: Grey and yellow

argillaceous PA from DDA036. These units

are significantly coarser grained than those

found in the LFB, and are invariably

argillaceous. C: An example of argillicous

porphyry andesite from DDA 40 (40 m from

the top of the hole).

Alteration

The unit is characteristically argillaceous; however some siliceous alteration is

evident which often results in a pitted texture where feldspars have been weathered out.

Formation

The coarse-grained and massive nature of this unit suggests that it is likely to be

a later-stage intrusion, or an unaltered section of the sequence. In addition, smaller and

altered units could represent the initial dome feeders. However these are more likely to

be encountered at depths not encountered yet with drilling. There are smaller and more

variable grey porphyry andesite units associated with the red/pink lava flow/lava flow

breccia unit, however these are inferred to be part of the dome-building facies i.e. lava

flow or dome cores.

7.3.2 Dome Related Lava Flow-Breccia Facies (LFB): Porphyry Andesite Lava

and Lava Flow Breccia (pink/red - siliceous)

Description

This unit is variably porphyritic pink andesite with obvious flow banded lava, and

with or without breccia (Figure 7.6). The unit tends to be finer-grained than the grey

porphyry andesite, with crystal sizes typically 1 to 2mm.

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The breccias vary from being the odd breccia block within the flow banded lava

to a more brecciated unit. The blocky breccias comprise mixed porphyritic material.

Figure 7.6 - A, B: Example from DDA 003 of

silicified Lava flow-breccia facies. Here pinky/red

porphyritic and variably flow banded dome

related material is found. C: Flow banded

porphyritic lava flow from DDA 10. D: Brecciated

flow banded porphyritic lava from DDA 16.

Alteration

The unit is characteristically siliceous; however some argillaceous alteration is

evident particularly at the contact with the grey porphyry andesite unit.

Formation

The flow banded lava is formed during the dome building episodes and the

breccias result from the collapse (i.e. block-and-ash flows) of an unstable dome and/or

proximal debris. Variations of the unit at core-scale show pulses of activity which are

represented by repeated flow banded lava units with brecciated margins. Waning

effusive volcanism could also be responsible for some of the brecciated unit, especially

where there are fewer blocks, i.e. blocky lava flow tops could be become incorporated

within the next pulse of activity. In addition, some more scoriaceous and juvenile

material have been observed at the lava flow tops (see additional units section).

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For the most part it is possible to differentiate between the PA and the LFB due

to the silicified alteration, finer grain size and complex textural features.

7.3.3 Volcaniclastic Facies (VC): Volcaniclastics (primary and reworked)

Description

The volcaniclastic unit comprises of a number of different facies but at corescale,

and due to alteration, it is difficult to further subdivide this unit at present

(Figure 7.7).

The primary volcaniclastic deposit is often represented by grey to grey-yellow

monomict (occasionally porphyritic) angular to sub-rounded clasts with or without

bedding. Clasts can be up to 10-15cm in size and are often 3-7cm. The reworked

volcaniclastic deposit is grey-yellow to yellow-grey with polymict, sub-rounded, variablesized

clasts (up to 7cm). Bedding is more evident in the reworked volcaniclastic deposit.

Figure 7.7 - A, B: Polymict volcaniclastic horizon from DDA 150. A) variation along section at

this horizon. B) Expanded view highlighting textural variation and mineralization in between

grains. C: An example of grey to grey-yellow volcaniclastics from DDA 10. D: Reworked

volcaniclastics from DDA 087.

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Alteration

Alteration of this unit is variable from massive silica where the primary lithology is

not evident and individual clasts are very difficult to see, to slightly siliceous (e.g.

Figure 7) where individual clasts and matrix are easy to define.

Where the alteration has been extreme, so that all that can be seen is grey to

grey/honey yellow, massive to patchy sections, with little textures visible, the lithology

was assigned due to the stratigraphic position of the volcaniclastic unit within the core.

Close inspection of massive sections can also reveal ‘ghost’ crystals, where rectangular

hollows which used to be feldspar phenocrysts, weather out, and faint outlines of grains.

Formation

The primary volcaniclastics are thought to result from mass wasting/erosion and

often represent a period of volcanic quiescence before another dome building event.

However, the more monomict breccias could be associated with dome collapse (blockand-ash

flows) and due to alteration were difficult to assign to the lava flow breccia

facies. Block-and-ash flows are strongly directed and are normally ungraded or

unbedded, with dense clasts. Where they are bedded (on cm or 10 cm scale) they could

be surges, associated with domes from phreatic eruptions. Block-and-ash flows are

characterized by having a matrix that is more or less broken up fragments of the clasts

in the unit.

The reworked volcaniclastics are likely to have resulted from fluvial interactions,

and where the matrix of the unit is full of clay or polylithic, they could relate to lahars

(volcanic flows ranging from debris flow to flood flow) or snow/avalanche melt units. The

polymict volcaniclastics could also be block-and-ash flows but are more likely to be talus

or colluvium around the flanks of the domes or small volcanoes.

Parts of the volcaniclastic unit were assigned to as “tuff” in the previous

logging/modeling (sometimes up to tens of meters thick). However, a thick pyroclastic

unit in this volcanic environment (small-medium scale andesitic dome building and

collapse events) is highly unlikely and is more likely to occur during large-scale Plinian

eruptions (e.g. Mt St Helens, Mt Vesuvius). Pyroclastics (ash, lapilli, and bombs) are to

be expected during the more explosive stages of dome building but at much smaller

scales. Possible lapilli ash units were observed but due to their small scale they were

included within the volcaniclastic unit.

In summary, the logged volcaniclastic sections can therefore be considered as

quite a mixture of actual rock lithologies, which for the most part are made of lapilli size

(2-64mm) angular to sub-rounded, silicified and mineralized units grading away from the

main dome centers, with some coarser breccia horizons. At this stage, and with the

remit of the current scope of work, highly altered and silicified grey porphyritic horizons

and other lithology types which are almost completely obscured in the areas where

original textures are erased with the degree of alteration, have all been mapped under

volcaniclastic units. A further delimitation may be possible of the types of units present

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in the volcaniclastic facies with a more detailed investigation, and could become clearer

in weathered outcrop examples. Much of the rock mass within the volcaniclastic facies

for the most part contains reasonable grade material, and therefore a separation where

possible from the dome related facies is a useful marker.

7.3.4 Additional Units

In addition to the units logged and discussed above some minor units/facies were

identified in the logging and include: 1) juvenile spatter/scoria, 2) glassy sparsely

porphyritic lava, 3) dark fine-grained andesite with pyrite and 4) minor pyroclastics (or

tuff) (Figure 7.8).

Figure 7.8 - Other Lithologies.

A: An example of a scoriaceous flow top from DDA 056 (25.6 m from the top of the hole). B: An

example of glassy sparsely porphyritic lava from DDA 124. C: An example of dark grey finegrained

andesite with pyrite from DDA 09. D and E: An example of lapilli tuff from DDA 21.

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The juvenile spatter/scoriaceous horizons are interesting in terms of volcanology

and represent a more mafic part of the volcanic system. Although they have been

oxidized and altered, the original structure can still be clearly seen, e.g. vesicles,

stretched mafic lava with juvenile fragments (Figure 7.9). These horizons are often

situated at the top of a lava flow unit and highlight a transition from lava flows into

volcaniclastic units. The porous natures of these layers mean they are more susceptible

to mineralization.

Figure 7.9 - Vesicular ‘scoriaceous like’ horizons from DDA 056.

A) streaked and plastically deformed juvenile lava.

B) Close up of A.

C) section showing juvenile fragments mixed with lithic components.

D) close up from C highlighting plastic deformation of juvenile (black) material and undeformed

lithics.

7.4 Geological interpretation

The facies mapped at Amulsar are consistent with a dome related complex, as is

often the case with high sulphidation systems (e.g. table in Sillitoe 1999, and refs

therein). The inferred locations of the domes align with the NW regional tectonic

structures and specific features in dome morphology, structure and lithology have been

determined by the tectomagmatic (e.g. Eocene-Oligocene calc-alkaline magmatic-arc

system) conditions under which they have been formed.

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The logging indicates that there have been several stages of dome building (lava

flows) and collapse (breccias) at the same center, with isolated collapse events (some

volcaniclastic units) that are separate from the main central areas. Periods of

quiescence are characterized by volcaniclastic deposits, which are often followed by

lava flow/lava flow breccias – representing another stage of the dome build up and

collapse.

The volcanological model is complicated by two or more centers being active in

the area at the same time, or at different protracted periods. The age and the relative

timing of the dome building and collapse events is difficult to determine because the

resolution of the cores does not allow one to distinguish events that may be several

thousands of years apart. The activity from more than one center has lead to a complex

facies association pattern as deposits originating from different dome centers are

interbedded. In addition dome growth and collapse is not always radial and depending

on the configuration of the vent, the existing slope angle etc. the deposits can be

strongly directional. This is especially so with block-and-ash flow deposits.

The volcaniclastic deposits which mantle the landscape (refer to the 3D

geological model) are more likely to be talus or colluvium which has formed around the

flanks of the domes or small volcanoes.

It is inferred that the massive porphyritic andesite (PA) is likely to be either a late

stage intrusive (domes, sills and/or dykes) and/or areas of the section which have not

undergone such significant alteration. In places the massive PA has been affected by

faulting, but in others it appears that the intrusion location and/or formation has

controlled by the faulting to some degree.

Since the dome building and collapse events NE-SW normal and NW-SE leftlateral

strike slip faulting has occurred and offset the original stratigraphy making facies

correlation difficult and adding to the complexity of this geological model.

Significant modification to the geomorphology at Amulsar is likely to have

occurred since the Eocene-Oligocene. The area has been glaciated and therefore

would have been affected by glacial processes and post-glacial erosion (e.g. Ollivier et

al., 2010). In more recent times fluvial (snowfall, snowmelt, rainfall) and erosional

processes would have also modified the landscape. Original volcanic domes could have

been eroded exposing only the cores (e.g. some porphyry andesite exposed at the

surface?) or parts of their original structure. Large-scale collapse deposits (debris

avalanche, crater wall collapse?) are evident on the landscape especially when looking

up to Amulsar from the road to the south west.

Overall, the facies mapping demonstrates a geological model with a complex

nature of dome building and collapse facies but also shows that the broad scale

features and distribution are dictated by the volcanology, while the more detailed facies

distribution has some significant fault control.

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7.5 3D Geology Model

The 3D geology model highlights 2 to 3 areas of dome building across the longsection

from Tigranes and Artavasdes (Figure 7.10), with lateral facies variation (from

lava flow breccia to volcaniclastic) moving away from the location of the inferred dome

centers. In addition a small area of porphyritic andesite is encountered at the center

between Tigranes and Artavasdes – this could be what was originally inferred to as an

intrusion, or perhaps an older volcanic core which is now exposed at the surface.

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Figure 7.10 - Geological model long section running NE-SW from Tigranes through to Artavasdes

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At Tigranes there is a possible porphyritic andesite feeder structure within the

lava flow/breccia facies, in addition to some intrusions. The feeder for the dome related

facies in Artavasdes is less evident in long-section and could possibly be encountered

at depth. Some porphyritic andesite is apparent at the base of the section along with

one or two intrusions into the lava flow breccia and volcaniclastic units.

At Erato large areas of porphyritic andesite are exposed at the surface and are

thought to relate to later stage intrusions (Figure 7.11). Dome building and collapse is

also evident in this locality from the lateral facies variation moving away from an inferred

centre of volcanism. Erato differs from Artavasdes and Tigranes as more volcaniclastic

deposits are mapped out at the surface whereas the lava flow breccia unit is more

predominant at the other centres.

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Figure 7.11 - Erato Geology Map

Significantly less porphyry andesite has been included in this model when

compared with previous models. This has important implications for the pit design

because the porphyry andesite is less competent and requires steeper pit wall angles.

Some of what was previously mapped as porphyry andesite has now been classified

into the lava flow/lava flow breccia unit. Elements of the lava flow breccia are still

porphyry andesite but the style of the unit is different and therefore mapped accordingly.

The 3D model includes important elements of the new logging, such as the

separation of facies associated with the dome building from the volcaniclastic units, and

the re-log of areas previously marked down as simply 'TUFF'. This provides a better

overall constraint on the facies relationships and will ultimately lead to a greater

understanding of the alteration and mineralization.

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7.6 Alteration and Mineralization

Alteration mapping was undertaken in parallel with the lithology logging of both

diamond cores and RC chips. Three main alteration facies identified are:

1) massive silica,

2) silica-clay

3) clay

In addition, areas of mineralisation (e.g. hematite, pyrite, alunite, and limonite)

were recorded. (See above detailed description of alteration for different lithology

facies).

Mineralization zones seem to coincide with the main volcaniclastic horizons,

brecciated parts of the lava flow breccia and to the scoriaceous horizons in the lava

flow/breccia facies. Therefore the grade is lithologically controlled to a point, see

Figures 7.12 and 7.13.

It is likely that a combination of fault control into areas where there is a higher

overall permeability (including structural permeability by way of fractures) has helped to

define the mineralization at Amulsar, i.e. the porous scoria/breccia has allowed fluid flow

and therefore mineralization. In addition, the more massive and less porous lava flows

of the dome related facies possibly helped to drive early fluid circulation by providing a

cap, or barrier, which allowed fluid flow into the more porous substrates (i.e. the

volcaniclastics). Grade also occurs in the lava flow breccias along these volcaniclastic

and fault margins.

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Figure 7.12 - Spatial relationship between the volcaniclastics, faults and grade

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Figure 7.13 - Geological Model vs Gold Grade

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8.0 DEPOSIT TYPES

Gold mineralization at Amulsar is a late event in the development of the deposit,

thus the spatial distribution is controlled by the preceding alteration and structural

events. Gold occurs within limonitic zones after the oxidation of sulfide material, and

three dominant controls on gold mineralization have been defined:




The argilic altered (kaolinized) porphyritic andesite formed an impermeable

barrier to mineralizing fluids, and caused ‘ponding’ of higher grade

mineralization along the contacts, often appearing as thick, leached and

‘gossanous’ zones. This is present throughout the deposits, but in different

styles. In the Tigranes area, the contacts are irregular yet generally steeply

dipping. In parts of Artavasdes and at Erato, the argilic altered porphyritic

andesite forms an overlying ‘cap’ which results in the localization of subhorizontal

mineralization within the siliceous units below.

Faults and fractures provided conduits for the gold mineralizing fluids,

resulting in high grade ‘gossanous’ veins. These often form broad corridors of

numerous, closely spaced high-grade structures.

Porous and permeable lithological units (hydrothermal breccias, volcaniclastic

breccias, and occasional leached/vuggy volcanics) allowed the lateral

migration of mineralizing fluids away from structurally controlled conduits

forming large tonnages of lower grade material. This is seen in Erato, where

weaker silicification has preserved the permeability of the original lithologies

and resulted in broad areas of lower grade gold mineralization.

The presence of silver mineralization is poorly understood, and does not appear

to correlate strongly with gold grades. Silver values at Amulsar average 2-5 g/t silver,

although can reach greater than 100 g/t within narrow intervals. A small silver mining

project adjoins the Amulsar license to the northwest, exploiting argentiferous galena

hosted in a structurally controlled vein. This deposit is also located at a lower

stratigraphic elevation to the Amulsar deposit.

Copper and other base metal mineralization are virtually absent in the resource

areas, although minor occurrences are present within carbonate veins from lower

stratigraphic elevations in the west of the Amulsar license area.

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9.0 EXPLORATION

Since commencing work in 2006, Lydian has carried out diverse exploration

activities at Amulsar including; geological mapping, various geophysical surveys,

surface sampling and drilling.

9.1 Geological Mapping

The current surface geology map for Amulsar has been compiled from historic

mapping (2006-2010) and a 2011 review of mapping done by Lydian geologists.

Mapping was conducted at scales varying from small scales, to better understand

structures in detail to large scale encompassing all areas of the concession.

9.2 Geophysical Surveys

Lydian has conducted multiple geophysical surveys from 2006 through 2012,

including Induced Polarization (“IP”), magnetic, and spectrometer geophysical surveys.

An 11 line km orientation dipole-dipole (DPDP) Induced Polarization

(IP)/resistivity survey utilizing 100 m electrode separations on seven lines was

completed at Amulsar between July 10 and 29, 2007.

A subsequent extension of the survey was completed during September /October

2007 where 15.4 line km were read on eleven survey lines using the same survey

parameters, operators, and support.

The objective of the survey was to evaluate the usefulness of the IP/resistivity

method in mapping the sub-surface potential of the gold-bearing high-sulfidation system

identified at surface. Highlights included several high resistivity responses associated

with silicification and corresponding low chargeability representing the oxidized portion

of the system near the main area of interest, and the mapping of a substantial, mostly

covered resistor to the north.

As a follow up to the 2007 campaign, a 28.2 line km program of pole-dipole

(PDP) IP/resistivity utilizing 100 m electrode separations on six survey lines was

completed at Amulsar in May / June of 2008.

The survey indicates that the mineralisation appears to be located on the

margins of the high resistivity volumes.

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Figure 9.1 - Location map of IP/resistivity surveys at Armulsar

The red lines show the 2007 coverage, while the black lines are from 2008. The outer red outline denotes the Armulsar license boundary.

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Two Ground Magnetic surveys were carried out by an Armenian geophysical

contractor in 2007-2008 covering a combined area of 2km x 3.5km over the main

prospect area (Figure 9.2). The survey was focused on delineating a possible porphyry

target on the southwest side of Artavasdes.

Spectrometer geophysics was carried out on the property in 2011. The survey

covers an area of 2.7 km (X) by 3.8 km (Y), in a grid spacing (20m (X) by 40m (Y),

(Figure 9.2).

Figure 9.2 - (15A left above) Plot of Ground Magnetics over the Central Amulsar Prospecting

License, (15B right above) Plot of The Spectrometer data.

9.3 Geochemistry

Bleg, soil, pit, channel, and rock sampling, have been collected and used to

evaluate mineralization potential and generate drilling targets since the beginning of

exploration activities at Amulsar.

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Table 9.1

Summary of Amulsar Geochemistry Samples

Project Type Total No Samples

Amulsar Bleg 12

Amulsar Linear Channel 279

Amulsar Pit 18

Amulsar Rock 921

Amulsar Soil 3105

Amulsar Trench 171

Amulsar Channel 750

Geochemical samples cover an area of 5km x 6km. Detailed soils on a 50m x

50m grid, trench, channel and rock chip sampling have been conducted in the main

area of the project, and in other areas 50m x 100m spaced soil sampling has been

conducted.

9.4 Drilling

A total of 739 holes have been drilled at Amulsar between 2007 and 2012,

including, exploration, geotechnical, metallurgical, and sterilization holes. Reverse

circulation (RC) and diamond drilling (DD) have both been carried out on the property.

Table 9.2

Summary of Amulsar Geochemical Samples

Diamond Drilling RC Drilling Total

Project Year # of

# of

# of

meters

meters

meters

holes

holes

holes

Amulsar 2007 5 593 0 0 5 593

Amulsar 2008 18 2679.3 74 10363 92 13042.3

Amulsar 2009 8 1089.8 101 12446 109 13535.8

Amulsar 2010 52 7575.9 129 17036 181 24611.9

Amulsar 2011 198 16113.2 154 23099 352 39212.2

Total 2007-2011 281 28051.2 458 62944 739 90995.2

Figure 9.3 below provides an overview of the drilling that has been completed on

the Amulsar Project since 2007. Drilling is ongoing at Amulsar but this report references

only drilling completed as at year end 2011.

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Figure 9.3 - Amulsar Drilling 2007-2011

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Although most of the exploration drill holes are drilled in the area of Tigranes-

Artavasdes-Arshak and Erato, where the current resource is delineated, a number of

additional exploration drill holes have been drilled in other areas such as Orontes, to

test geochemistry anomalies and between Erato and Artavasdes/Tigranes to test the

possible structure joining them.

9.5 Database

All Lydian exploration data is stored in the professional geology database,

FUSION (GDMS).DH Logger is the application used to enter all drilling data into the

database (Collars, Surveys, Geology and Alteration, Assays, Recovery/RQD, and

Structures), whilst Sample Station is the application used to store all geochemical

samples and their data (Soils, Rocks etc). The secure location for storage of data is the

Central Database is the Lydian head office in Jersey.

9.6 QA/QC

QA/QC samples are stored in Lydian’s Database FUSION (GDMS). All QA/QC

samples such as duplicates, standards and blanks are submitted to the laboratory with

samples batches at a 1:20 ratio. Certified reference material is purchased by Geostats

Pty Ltd.

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10.0 DRILLING

Pre 2011 drilling results at Amulsar are covered in the previous mineral resource

technical report by CSA Global titled “Mineral Resource Estimate, Lydian International

Limited, Amulsar Gold Project, 43-101 Technical Report, Armenia”, dated May 19,

2011, and in the PEA report prepared by K D Engineering titled “Development of

Amulsar Heap Leach Facility - Preliminary Economic Assessment”, dated 12 August

2011, both are posted on SEDAR. A summary of the drilling discussion from the PEA

report is presented below.

10.1 Pre 2011 Drilling

Drilling in 2008 was focussed on defining the main body of the Amulsar deposit

to satisfy regulatory requirements to gain a mining license. Drilling in 2009 was

undertaken as a step out program, surrounding the previous drilling and to test the

extents of the Amulsar deposit. Drilling in 2010 was focussed on infill drilling at

Artavasdes, Tigranes and Erato, to upgrade some parts of the resource to higher

categories. Lydian carried out two types of drilling during 2010: reverse circulation (RC)

and diamond core (DDH) drilling. A total of 49 core holes for 7,502 meters (m) and 127

RC holes for 17,036 m were completed during 2010.

Drill holes were predominately drilled at -60 degrees, orientated in the northern

area at an azimuth of 120 or 300 and either southwest or southeast orientations in the

southern area to optimize the intersection angle with the steeply dipping mineralized

zones. Holes were laid out on northwest-southeast grid lines. In central areas, at

Tigranes and Artavasdes, drill holes were spaced 40m apart, with holes generally drilled

at 40m intervals along these sections. Hole spacing became larger towards the edges

of the deposit. Drilling at Erato was based on a rough 80 m by 80 m grid.

Hole depth generally ranged from 100-250m but most holes were drilled to

approximately 140m down hole depth. The diamond drill holes were usually drilled at

PQ diameter for the first 20 m - 30 m and then reduced to HQ diameter. Standard

practice during the drill program was to survey all drill collars and to carry out down hole

directional surveys. In addition, during drilling, regular orientation tests were carried out

using the Ezy Mark system to help establish the true dip and strike of the rock units and

vein/fracture systems.

Down hole surveying was conducted using a Globaltech Pathfinder system and

no deviation issues were identified. After logging of rock chips or drill core, samples

were taken, generally at 1m intervals throughout the entire hole, and dispatched for

multi-element analysis and gold assay. A total of 14,613 samples were assayed from

drill core, as well as 21,770 rock chip samples from RC holes.

CSA conducted a suite of comparisons between the diamond core and RC

drilling and found no reasons that both drill types could not be used for resource

estimation.

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10.2 2011 Drilling

Drilling continued during 2011 with an additional 287 holes drilled, comprised of

133 core holes for 15,407m and 154 RC holes for 23,099m. In total, 38,506m of new

drilling was completed during 2011. Drill-Ex International was responsible for all of the

drilling on site during this time as it was for the 2010 drilling.

Drill results, lithological interpretations and assay results have further defined the

two previously identified zones of gold mineralization, Tigranes and Artavasdes.

Increased drilling to the north better defined Erato. Many of the holes show a pattern of

relatively widespread low grade mineralization (generally


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Table 10.1

Selected Down Hole Gold Intercepts for the 2011 Amulsar Drilling Program

Area

Drill Hole

#

From

(m)

To

(m)

Intersection

(m)

Gold

(g/t)

Artavasdes DDA-096 0.0 69.0 69.0 1.0

Artavasdes DDA-109 0.0 101.0 101.0 2.2

Tigranes DDA-111

55.0 124.0 69.0 3.8

68.0 100.0 32.0 7.9

Artavasdes DDA-116 1.0 120.0 119.0 1.3

Artavasdes DDA-119 92.0 136.0 44.0 3.3

Tigranes DDA-147

0.0 20.0 20.0 1.9

48.0 111.0 63.0 2.0

Artavasdes DDA-161 83.0 163.0 80.0 1.0

Tigranes DDA-200 144.0 193.0 49.0 1.2

Tigranes DDA-223 47.0 146 (EOH) 99.0 4.0

Tigranes RCA-306 75.0 130.0 55.0 1.1

Tigranes RCA-308 157.0 185 (EOH) 28.0 2.7

Erato RCA-330 232.0 239.0 7.0 6.6

Tigranes RCA-348 22.0

(EOH)

132.0 110.0 1.0

Tigranes RCA-353 99.0 168.0 68.0 1.2

Tigranes RCA-359 26.0 48.0 22.0 2.2

Tigranes RCA-369 0.0 98.0 98.0 1.5

Tigranes RCA-373 109.0 136.0 27.0 1.4

Artavasdes RCA-388 23.0 179.0 156.0 1.7

Artavasdes RCA-395 75.0 161 (EOH) 86.0 1.0

Artavasdes RCA-396 10.0 149(EOH) 139.0 1.0

Artavasdes RCA-419 188.0 215.0 27.0 4.2

Erato RCA-441 105.0 251(EOH) 146.0 1.6

Artavasdes RCA-445 42.0 180.0 138.0 1.5

The drill holes drilled through the end of the 2011 are shown on Figure 10.1 and

support the mineral resource included in this report.

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Figure 10.1 - Drilling Through End of 2011

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11.0 SAMPLE PREPARATION, ANALYSIS AND SECURITY

The sample collection, preparation and analysis have been reviewed by IMC and

CSA with the drilling through 2010 review documented in the CSA report “Mineral

Resource Estimate, Lydian International Limited, Amulsar Gold Project, 43-101

Technical Report, Armenia” dated May 19, 2011 and posted on SEDAR. A summary

from that report is incorporated below and includes descriptions of sample preparation

procedures.

Herb Welhener of IMC visited the site and the preparation facilities at Gorayk

during his visit of June 20 - 24, 2011. During this time, he observed the drilling and

sample preparation at Amulsar and visited the sample storage and preparation lab at

Gorayk, as well as, viewing core from several holes.

11.1 Sample Collection

11.1.1 Reverse Circulation Drilling

Reverse Circulation drill (RC) samples are taken every one meter down the hole

by collection of rock chips at the cyclone on the RC rig. Lydian state that the hole is

normally air-flushed out after each metre to avoid contamination of the following sample.

At the drill site the samples are weighed, logged, bagged, labelled and sealed, prior to

transporting to the Amulsar Camp for splitting. Samples are split to produce two 1.5 - 2

kg samples, one for assay and the second to be kept as an archive sample. The riffle

splitter is cleaned between samples by brushing and the use of compressed air.

Individual weights for the entire 1 meter sample, and the final sample were recorded in

order to recalculate any over-/under-estimations. These samples are transported to the

Lydian sample preparation facility at Gorayk for insertion of QA/QC samples into the

sample batch and dispatch to the assay laboratory. Archive RC samples are stored at

the Gorayk facility.

11.1.2 Diamond Core Drilling

Geological and geotechnical logging are carried out on the core at the Amulsar

camp. Once logged sample intervals are marked up on core boxes prior to the core

being transported to the sample preparation facility at Gorayk. Sample lengths can vary,

depending on geology and other factors, but are typically in the range of 1 meter

Prior to the sample split at Gorayk, the following activities are completed:

The sample intervals are marked directly on the core.

• Lydian systematically photographs all core prior to sampling.

• Lydian systematically logs an estimate of percent porosity and measures

all faults and fractures.

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11.2 Sample Preparation, Analysis and Security

Lydian’s sample preparation facility is located in the village of Gorayk to the

southeast of the Amulsar project. The facility includes core saws, a core crushing unit,

and storage for both core boxes and archive samples. Prior to the establishment of their

own crushing facility, all Lydian rock and half core samples from the Amulsar project

were bagged, labelled and sealed on site and transported to the ALS Chemex analytical

laboratory in Romania for sample preparation and gold assay.

In mid-September 2008 Lydian installed its own crushing facility at Gorayk. This

comprises an ALS Chemex mobile containerised unit, consisting of a drying oven, jaw

crusher and riffle splitter. The crusher can crush split core to 90 percent less than 2 mm

and also allows Lydian staff to manage on-site, the coarse rejects of the crushed

diamond half-core. The crusher and riffle splitter are cleaned between samples by

brushing and the use of compressed air. Lydian employ an experienced local analytical

laboratory operator to run the unit and uses this facility to carry out preliminary

preparation of its rock and half core samples, allowing it to dispatch 1 kg crushed

samples to the ALS Chemex laboratory in Romania.

Core samples are split by diamond saw in two equal halves - one half is returned

to the core box for future reference, while the other half is bagged, and oven dried prior

to crushing. An assay sample and an archive sample are collected from the splitting of

the crushed core. Diamond core duplicate assays are produced as an additional split

sample. Since the breakdown of the crusher in September 2010 half core has again

been sent to ALS Romania, with quarter core being used for duplicate samples.

On arrival in Romania ALS gives the samples a batch code and weighs all

samples. Each sample is then logged in and ascribed an individual bar code. Rock

samples not already received in crushed form are subject to fine crushing to 70 percent

of material passing less than 2 mm. Samples are split using a riffle splitter and a 1,000

g split is then pulverised to 85 percent passing less than 75 µm.

A 50 gram sub-sample is analysed at the Romanian laboratory for gold by fire

assay, with an AA finish. The remainder of the split sample is sent by ALS Chemex to

its sister laboratory in Perth (pending registration to ISO 9001:2000) or Vancouver

(ISO17025), where the sample is subjected to a four acid digest, followed by analysis

for 33 elements by ICP-MS (ALS ME-ICP61 package). Additional analyses are

undertaken for samples returning assays above the detection limit (typically silver, zinc

and lead).

All ALS Chemex laboratories operate in compliance with ISO17025. ALS

Chemex carries out regular checks by duplicating analyses for several samples, and by

inserting its own blanks and standards as part of its own Quality Management System.

Lydian ships its samples and includes blanks, duplicates, and standards of different

concentrations with random numbering as part of its QA/QC program.

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In mid-September 2010 the crusher at Gorayk had a mechanical breakdown and

was not in operation during CSA’s May 2011 visit. During the IMC visit in June of 2011,

the sample preparation facilities were visited and the core sawing was observed. No

crushing activities were being done during the time of the IMC visit. The half core

samples are shipped to ALS for sample preparation and analysis until the new sample

preparation unit is operational at Gorayk.

Lydian ordered a new containerised sample preparation unit from ALS Chemex

in Johannesburg which is due for delivery in August 2012 and will replace the existing

unit. The new sample preparation laboratory will contain two crushers a rotary splitter

divider and a pulveriser allowing Lydian to reduce the sample size shipped for assay to

200 - 250 grams which will significantly reduce sample shipping costs to Romania of

samples. The addition of this additional equipment was discussed during the IMC visit,

but IMC has not verified first-hand the installation of the new equipment.

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12.0 DATA VERIFICATION

This section presents the combined data verification efforts of CSA during its

2011 due diligence work and the IMC site visit review and QA/QC review. The CSA

work is summarized from its report dated May 19, 2011, referenced in the introduction

of this report and is also included in the PEA report published by K D Engineering,

dated 12 August 2011. During the site visit by Herb Welhener of IMC, no independent

samples were collected for analysis.

12.1 CSA Data Verification

Nerys Walters, Senior Geologist, CSA Global, visited the Lydian core storage

facility at Gorayk village during May 2011 to view drill core from the Amulsar prospect

and to review geology, sampling procedures and assay results. Core from holes DDA-

030, DDA-034, DDA-035, DDA-036, DDA-044 and DDA-047 were examined. From a

study of the assay results for these holes it was clear that the highest gold grade

intercepts were directly related to massive clayey limonite zones, while wider, lowergrade

sections related to pervasively silicified and brecciated zones, sometimes with

limonitic matrix.

Miss Walters also visited the Amulsar site, where she located and measured the

collar positions for holes RCA-223, RCA-280 and DDA-044. They were surveyed in

using handheld GPS along with the survey Trig beacon at the top of Artavasdes. A brief

review of the geology was undertaken but thick snow cover prohibited review in some

areas.

Observation of diamond drill practices were undertaken on the lower slopes of

Amulsar. CSA considers diamond drill practices to be of a high standard with a well

organised and safe work environment. Drilling was undertaken by well-trained

competent drillers and helpers. No further on-site verification has been undertaken, and

it has been assumed that all processes and procedures that were reviewed and found

to be adequate during the previous review remain current.

In preparation for the forthcoming feasibility study, Lydian engaged Wardell

Armstrong Ltd. in 2010 to review all diamond and RC drilling, sampling and assaying

QA/QC procedures for assayed core and metallurgical test work samples as well as

comparing results of twinned RC and core drill holes. Wardell Armstrong concluded the

following:

“With regard to the exploration programme itself, WAI is very satisfied with

all aspects of the work from the planning, execution and results derived

therefrom. Furthermore, Lydian has used international best practice at all

times to undertake the drilling and ancillary exploration works, also made

easier by the selection of a highly competent drilling contractor in Drill-Ex

International” (Newall, 2010).

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12.1.1 CSA QA/QC Review

Industry standard assay QA/QC control was implemented by Lydian during

exploration and resource development activities in 2008, these practices are still

current.

Lydian has submitted approximately 1,200 pulp samples, from 2006-2010 RC

and diamond drilling, to Alfred H Knight in Alaska. Lydian states that these samples

come from holes drilled across the resource area and have been selected at a rate of 1

in 20 (5%) for assay greater than 0.2 grams per tonne gold and 1 in 40 (2.5%) for

assays less than 0.2 grams per tonne gold. A similar program has been undertaken at

the end of the 2011 drill program.

In addition to reviews undertaken as part of previous resource work in 2009, CSA

conducted a review of QA/QC data from RC and DDH sampling undertaken during

2010. A summary of QA/QC data is tabulated below.

QA/QC

Check

Blank

Samples

Duplicate

Samples

Standard

Expected

Grade (PPM)


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12.1.2 CSA Conclusions for RC QA/QC

Results from the analysis of blank reference samples shows less than 1 percent

of samples (2/470) returning grades greater than 3 times the detection limit ( greater

than 0.015ppm gold) which is considered acceptable. Sample assays adjacent to these

outliers were checked for the possibility of contamination during sample preparation.

Adjacent samples were not mineralised. CSA recommended that during future drilling

campaigns the primary laboratory be audited to ensure adequate sample preparation

practises are adhered to.

Results from the analysis of duplicate sample assays suggest a close correlation

between the original assay and the duplicate assay (correlation coefficient = 0.98).

On the whole, results from the submission of standard reference material

suggest assaying accuracy to be within acceptable limits. The control plots show that

99%, 100%, 98%, 100%, 99%, 98% and 93 percent of samples returned values within

an acceptable tolerance of plus or minus 10 percent for CRM’s 300-7,303-2,398-6,900-

6, 904-8, GLG304-1 and 307-2 respectively. Although the majority of returned values

fall within acceptable limits, CRM 300-7 and 398-6 show a slight bias towards underreporting

and CRM 304-1 and 302-2 show a slight tendency towards over reporting.

There appear to be no significant, non-random patterns to any dataset other than weak

to moderate cyclicity within acceptable limits. It is probable that this cyclicity is related to

equipment calibration.

Outliers are present in the control charts, notably for CRM G300-7, G398-6,

G904-8, GLG 304-1 and G307-2. Review of these outliers indicates that they may be

due to incorrect labelling of the CRM, resulting in some confusion in expected values.

CSA Recommends that Lydian reviews systems in place for the insertion of CRM

samples to minimise these errors in future.

The results of standard analysis suggest no significant bias exists and assay

results may be considered reliable. CSA notes that although flagged, outliers have not

been investigated by Lydian at the time of reporting. CSA recommends that Lydian

reviews systems currently in place that manage the insertion of CRM samples to

minimise these errors in future.

Current blank samples are sourced from alluvial sand for RC samples. Although

no significant contamination has been identified CSA recommended that Lydian source

certified blank material to align their procedures with industry best practice. Lydian is in

the process of acquiring certified blank standards.

12.1.3 CSA Conclusions on Core Hole QA/QC

Results from the analysis of blank reference samples show less than 1 percent

(2/267) of samples returning grades greater than 3 times the detection limit (greater

than 0.015 ppm gold). Sample assays adjacent to these outliers were checked for the

possibility of contamination during sample preparation.

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Adjacent samples were not mineralised. CSA recommended that during future

drilling campaigns the primary laboratory be audited to ensure adequate sample

preparation practises are adhered to.

Results from the analysis of duplicate sample assays suggest a reasonable

correlation between the original assay and the duplicate assay (correlation coefficient =

0.83).

On the whole, results from the submission of standard reference material

suggest assaying accuracy to be within acceptable limits. The control plots show that

95%, 100%, 98%, 100%, 97%, 92% and 100 percent of samples returned values within

an acceptable tolerance of plus or minus 10 percent for CRM’s 300-7,303-2,398-6,900-

6, 904-8, GLG304-1 and 307-2 respectively.

Although the majority of returned values fall within acceptable limits, CRM 303-2

and 304-1 show a slight bias towards over-reporting grades, whilst CRM 398-6 and 300-

7 show a slight bias towards under reporting grades.

There appear to be no significant, non-random patterns to any dataset other than

weak to moderate cyclicity within acceptable limits. It is probable that this cyclicity is

related to equipment calibration.

Outliers are present in the control charts, notably for CRM G300-7, GLG 304-1,

G398-6 and G904-8. Review of these outliers indicates that they may be due to

incorrect labelling of the CRM and Blank samples, resulting in some confusion in

values. CSA recommended that Lydian reviews systems currently in place that manage

the insertion of CRM samples to minimise these errors in future.

The results of standard analysis suggest no significant bias exists and assay

results may be considered accurate. CSA notes that although flagged, outliers have not

been investigated by Lydian at the time of reporting. CSA recommends that QA/QC

data should be monitored as drilling progresses, and any erroneous data investigated

as they arise. Lydian has improved QA/QC monitoring and addressed these issues.

CSA considers the assay results collected from RC drilling and diamond drilling

to be reliable for the purposes of resource estimation, being representative of the

sampled material and exhibiting no significant bias. RC and diamond assay datasets

exhibit similar population distributions and are considered compatible for use in

resource estimation.

Current blank samples are sourced from local basalt for diamond samples.

Although no significant contamination has been identified, Lydian is in the process of

sourcing certified blanks.

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12.1.4 Lab QA/QC

Lydian provided results of laboratory QA/QC duplicate analysis to CSA for

review. Results from the analysis of laboratory duplicate sample assays suggest a

reasonable correlation between the original assay and the duplicate assay for both RC

and diamond drilling, with correlation coefficients of 0.99 for both, Figure 12.1 and 12.2.

These figures show that duplicate results exhibit acceptable levels of accuracy and

precision at grade ranges less than 2 ppm gold. At elevated grades, less levels of

precision are observed.

Figure 12.1 - Diamond drilling Laboratory Duplicates

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Figure 12.2 - Laboratory RC drilling Duplicates

CSA considers the results of Lydian’s internal QA/QC and the results detailed

above adequate to assess the accuracy and precision of the ALS Chemex analytical

laboratory in Romania.

12.2 QA/QC Review by IMC

The June 2012 QA/QC review includes observations made in the IMC report for

the 2012 Mineral Resource Update dated January 22, 2012 as well as new information

for the 2011 drill program. A summary of the review results are included below.



The Amulsar deposit has been drilled using both diamond core and reversed

circulation (RC) methods. Paired comparisons of core and RC drillholes show

no significant differences in gold grades or grade distributions between

assays run on core and RC samples. However, it is recommended that RC

drilling be de-emphasized in favor of diamond core drilling as the project

advances.

Check assays run by an independent assay laboratory (Alfred H. Knight) on

pulp samples from the 2010 and earlier drilling compare closely with the gold

assays run by the primary laboratory (ALS), with sample pairs clustering

closely around the 1:1 line and mean grades comparing within 1 percent

when one outlier pair (probably a mislabeled sample) is discarded.

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Check assays run by an independent assay laboratory (Acme) on pulp

samples from the 2011 drilling also compare closely with the gold assays run

by the primary laboratory (ALS), with sample pairs clustering closely around

the 1:1 line and mean grades comparing within 1 percent.

The assays run by ALS on standard and blank samples are acceptable.

Sample preparation procedures could not be checked directly because no

check assays have been run on fresh pulps prepared from rejects by an

independent assay laboratory, but assays run on samples from the pre-2011

drilling prepared entirely at the ALS laboratory in Romania between 2007 and

September 2008 and partly at ALS and partly at Amulsar after September

2008 both compare closely with the Knight check assays, and this acts as a

partial verification of ALS and Amulsar sample preparation procedures over

this period. Paired comparisons of 10 meter gold composites also show no

significant differences in gold grades or grade distributions between assays

run on 2011 and pre-2011 samples.

Check assays run by ALS on pulps prepared by ALS from split core or unsplit

RC samples, however, show very poor individual-assay repeatability when

compared with the original ALS assays run on 1kg splits prepared at Amulsar.

It is recommended that procedures at ALS and Amulsar be reviewed to

determine the cause of this poor repeatability and to take appropriate action

to correct it.

Based on these results IMC considered that the Amulsar gold assays meet NI43-

101 standards for reporting mineral resources. IMC concurs with CSA that Lydian

should do lab audits to assure all industry standard practices are being met. IMC also

recommends that future check assays be run on fresh pulps prepared by an

independent laboratory from rejects; since check assays run on pulps prepared by the

primary laboratory (or the Lydian sample preparation facility at Gorayk) does not check

the sample preparation procedures.

No QA/QC data were supplied for silver. It is recommended that silver check

assays and assays on a silver standard be run since silver is to be given economic

credit in the financials

12.2.1 Paired Comparisons

IMC performed paired comparisons of 10 meter gold composites to investigate

whether there were any significant differences in mean gold grades and/or grade

distributions between diamond core and reverse circulation drilling or between the

assays from the pre-2011 and 2011 drilling.

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Figure 12.3 compares the gold grade distributions in core and RC composites

paired within a limiting separation distance of 50 meters. Grade distributions are

effectively the same and mean grades compare within a percent. These results confirm

that the RC samples are not globally biased high relative to the core samples.

Figure 12.3 - Core Hole Gold Grade Paired with RC Hole Gold Grade within 50 m,

All Drilling Mean RC = 0.415 g/t gold, Mean Core = 0.419 g/t gold.

Total 1,564 pairs (approximately 17,600m of core and 17,600m of RC drilling)

Figure 12.4 compares the gold grade distributions in pre-2011 composites with

those in 2011 composites within a limiting separation distance of 50 meters.

Grade distributions are again effectively the same and mean grades compare

within 5 percent. These results confirm that there are no significant global differences

between the gold grades and grade distributions measured in the 2011 and earlier

drilling campaigns.

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Figure 12.4 - 2011 Gold Grade Paired with Pre-2011 Gold Grade within 50 m

All Drilling Mean 2011 = 0.378 g/t Gold, Mean Pre-2011 = 0.361 g/t Gold.

Total 1,771 pairs (approximately 20,000m of 2011 and 20,000m of pre-2011 drilling)

12.2.2 Check Assay Comparisons

ALS was the primary laboratory for samples collected during the pre-2011 drilling

campaigns, and 1,215 check assays on ALS pulps (approximately one assay every 36th

sample) were run by the Alfred H. Knight laboratory in Alaska. The Knight check assays

are compared with the ALS data base assays in the Figure 12.5 XY-plot. Except for one

outlier (probably a mislabeled sample or a data entry error) the points cluster closely

around the 1:1 line and mean grades match within a percent:

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Figure 12.5 - ALS vs. Alfred H. Knight Check Assays

All Data, Pre-2011 Drilling

Mean grades are ALS = 0.596, Knight = 0.590 with outlier removed

Initially all half-split core and RC rock chip samples from the Amulsar drilling

were bagged and sent to ALS in Romania for sample preparation, but in September

2008 Lydian installed its own crushing facility at Amulsar and thereafter sent smaller

crushed core and/or split RC samples to ALS. However, this procedural change had no

impact on the ALS assays, which as shown in the Figure 12.6 and 12.7 XY plots

compare equally well with the Alfred H. Knight check assays both before and after the

change.

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Figure 12.6 - ALS vs. Alfred H. Knight Check Assays, Samples Prepared at ALS, Pre-2011 Drilling

Mean grades are ALS = 0.710, Knight = 0.701 with outlier pair removed

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Figure 12.7 - ALS vs. Alfred H. Knight Check Assays

Samples Crushed/Split at Amulsar Before Shipment to ALS, Pre-2011 Drilling

ALS was also the primary laboratory for samples collected during 2011, and

1,610 check assays on ALS pulps (approximately one every 23rd sample) were run by

Acme. The Acme check assays are compared with the ALS data base assays in the

Figure 12.8 XY plot. Again the points cluster closely around the 1:1 line and the means

agree to within a percent.

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Figure 12.8 - ALS Data Base Assays Versus Acme Check Assays, 2011 Drilling

As part of the QA/QC program 1,961 duplicate assays (approximately one every

42nd sample) were run by ALS on second-split pulps prepared from the original reject.

The duplicate assays are compared with the data base assays in the Figure 12.9 XY

plot. There is moderate scatter of point around the 1:1 line but mean grades still

compare within 2 percent.

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Figure 12.9 - ALS Data Base Assays Versus Duplicate Assays, All Drilling

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Figure 12.10 - ALS Data Base Assays Versus Check_au, All Holes

To act as a check on assay precision, 352 “check_au” assays were also run on

samples from selected core and RC holes. The assays were run on fresh pulps

prepared by ALS from half-core or RC rock chip samples sent from Amulsar to ALS,

and the check_au assays are compared with the original ALS data base assays in the

Figure 12.10 through 12.12 XY plots.

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Mean grades are broadly comparable in all three cases but the level of scatter

around the 1:1 line is very large even in the case of lower-grade samples.

Figure 12.11 - ALS Data Base Assays Versus Check_au, Hole RCA-301 (High-Grade)

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Figure 12.12 - ALS Data Base Assays Versus Check_au, Holes DDA-035, 047, 120, 133 (Low Grade)

IMC does not regard these poor comparisons as a matter of serious concern

because many assays are averaged together when resources are calculated and this

has the impact of canceling out random errors in individual assay values. Nevertheless

it is recommended that studies be conducted to determine the cause of the scatter.

12.2.

The results of the check assay cases discussed above are summarized in Table

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Table 12.2

ALS Data Base Assays Versus Check Assays for Gold

Case Number ALS mean Check mean Check/ALS %

Knight before 9/2008 299 0.710 0.701 -1.3

Knight, 9/2008 through 2010 915 0.561 0.551 -1.8

Knight, all data 1,214 0.596 0.590 -1.0

Acme, 2011 samples 1,610 0.320 0.317 -0.9

ALS Duplicates, All Data 1,961 0.330 0.325 -1.5

ALS “Check_au” 352 0.724 0.820 +13.3

12.2.4 Standard and Blank Assays

Lydian supplied the results of 3,811 assays run on 14 Geostat certified gold

standards, representing the insertion of a standard into the ALS sample stream

approximately once every 21st sample. The results of the standard assays were

generally excellent, with only minor scatter around the certified standard grade and with

the means matching within a few percent in all cases except for the very-low-grade

standards. The results are summarized in Table 12.3. (Note that 42 assays where the

standard had been mislabeled are excluded).

However, it is not necessary to use 14 different standards for gold QA/QC, and

IMC believes the number could be lowered to four for the purposes of future QA/QC

work. The four recommended standards are shown highlighted in bold type in Table

12.3.

Standard

Standard Grade

g/t

Table 12.3

Summary of Assays on Gold Standards

Number of

Assays

Years Covered

Mean Assay

Grade

g/t

Assay/ Standard,

G306-3 8.66 210 2010 & 2011 8.80 +1.7

GBMS304-4 5.67 100 2010 & 2011 5.75 +1.5

G904-8 5.53 672 2007 to 2011 5.57 +0.7

G303-2 4.15 501 2008 to 2011 4.23 +1.9

G398-6 2.94 456 2010 & 2011 2.95 0

G900-6 2.59 242 2008 to 2010 2.59 +1.2

G399-6 2.52 46 2008 2.57 +1.9

G302-2 2.50 125 2010 & 2011 2.56 +2.3

GBMS304-5 1.62 99 2011 1.61 -0.4

G307-2 1.08 293 2010 & 2011 1.08 0

OXD57 0.413 53 2008 0.408 -1.2

GLG304-1 0.153 699 2007 to 2011 0.160 +4.3

GLG303-2 0.017 150 2010 & 2011 0.013 -27.3

GLG307-1 0.0029 123 2011 0.0025 -13.8

Grade-versus-time plots for assays run on the four highlighted standards are

shown in Figures 12.13 through 12.16 (The horizontal red lines show the certified

standard grade and the time scale is related to drilling date because assaying dates

were not supplied). The assays run on the other ten standards show comparable

results.

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Assays on all of the standards except for GLG304-1 (Figure 12.16) remain very

constant. The assays on standard GLG304-1 show a tendency to increase with time but

the magnitude of the increase is only on the order of 0.01 g/t gold.

7

6

5

4

3

2

1

0

0 200 400 600 800 1000 1200 1400 1600

Figure 12.13 - Standard G904-8, Grade 5.53 g/t

X-Axis = Days from June 17, 2010 (Plot Ends October 24, 2011)

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3.5

3

2.5

2

1.5

1

0.5

0

0 100 200 300 400 500 600 700 800 900

Figure 12.14 - Standard G900-6, Grade 2.56 g/t

X-Axis = Days from May 8, 2008 (Plot Ends August 10, 2011)

1.4

1.2

1

0.8

0.6

0.4

0.2

0

0 50 100 150 200 250 300 350 400 450 500 550

Figure 12.15 - Standard G307-2, Grade 1.08 g/t

X-Axis = Days from June 10, 2010 (Plot Ends October 24, 2011)

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0.2

0.18

0.16

0.14

0.12

0.1

0.08

0.06

0.04

0.02

0

0 200 400 600 800 1000 1200 1400 1600

Figure 12.16 - Standard GLG304-1, Grade 0.153 g/t

X-Axis = Days from June 26, 2007 (Plot Ends October 24, 2011)

A total of 2,255 assays run on blank samples (approximately one every 36 th

sample) also gave acceptable results, with the vast majority of the assays returning

grades of less than 0.01 g/t gold. The results of the assays run on blanks are

summarized in Figure 12.17.

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0.1

0.09

0.08

0.07

0.06

0.05

0.04

0.03

0.02

0.01

0

0 200 400 600 800 1000 1200 1400 1600

Figure 12.17 - Blank Assays

X-Axis = Days from September 17, 2007 (Plot Ends November 1, 2011)

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13.0 MINERAL PROCESSING AND METALLURGICAL TESTING

Over the last five years metallurgical testing has been conducted on

representative samples from the Amulsar project to develop operating parameters and

metallurgical predictors for a crushed heap leach gold and silver processing facility. The

following review of the metallurgical test work has been extracted from the August 2011

43-101 PEA and expanded upon to include 2011/2012 testing.

13.1 SGS Lakefield Research (2008)

In September 2008 a gold recovery test program was undertaken at SGS

Lakefield in Canada on crushed continuous half drill-core from the entire 146 m length

of hole DDA-004, a scout hole from the 2007 drill program. The metallurgical program

consisted of a qualitative mineralogical evaluation of the composited sample, Bond ball

mill work index determination, gravity concentration, cyanide leaching of the gravity

tailings and the whole ore, and an evaluation of the amenability or the ore to heap

leaching. A summary of this work is contained in the March 2009 report, to which the

reader is referred.

13.1.1 Head Assays

Two 1 kilogram charges were submitted for duplicate screened metallics analysis

for gold at 150 mesh. The minus 150 mesh fraction was assayed for gold in duplicate.

The results indicated that the calculated head grade for the samples submitted for

assay was 1.06 to 1.08 g/t gold. Only 1.5 to 1.9 percent of the gold reported into the

plus 150 mesh fraction indicating the gravity recoverable gold component in this sample

is fairly low. A 100 to 200 gram sample of the minus 10 mesh material was submitted for

sulfur, sulfide, and a multi-element semi-quantitative ICP scan.

The results of these assays are summarized below in Tables 13.1 and 13.2.

Table 13.1

Screened Metallics Analysis for Gold, Comp 1

Composite

Gold Distribution,

Calc. Head +150 mesh Fraction -150 mesh Fraction

%

Grade Au,

Au, Au, Au avg., +150 -150

g/t

Au,g/t

g/t g/t g/t Mesh Mesh

Comp Cut A 1.06 0.78 1.05 1.09 1.07 1.9 98.1

Comp Cut B 1.08 0.80 1.05 1.13 1.09 1.5 98.5

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13.1.2 Mineralogy

Table 13.2

Head Analyses of Composite 1

Element Units Comp 1 Element Units Comp 1

S % 0.08 S 2- % 0.05

ICP Scan

Ag g/t 2 Mo g/t 11

Al g/t 3,700 Na g/t 160

As g/t 88 Ni g/t


%Au Recovery

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13.1.3 Coarse Bottle Roll Leach Test Results

In order to evaluate the amenability of the ore to heap leaching, a series of

coarse ore bottle roll tests were completed at two feed sizes, minus 1/2 inch and minus

1/4 inch. Typically in these tests the bottles are rolled intermittently (1 minute every

hour) in order to maximize solution contact with the ore while minimizing ore attrition.

Results of the coarse bottle roll leach tests are summarized in Table 13.3.

Table 13.3

Coarse Bottle Roll Leach Test Summary

Leach

Calc.

Gold Recovery,

Final

Feed NaCN

CaO

Head

%

Residue

Size Added Cons. Added Cons.

Grade 1 3 7 10 15

g/t Au

kg/t kg/t kg/t kg/t

Au, g/t day days days days days

1/2" 0.75 0.25 1.32 1.32 0.06 1.14 80.1 89.4 90.3 90.7 94.7

1/4" 0.70 0.23 1.36 1.37 0.04 1.00 81.1 92.9 94.2 94.1 96.0

The gold leach curves for the coarse bottle roll leach tests conducted at different

crush sizes are shown in Figure 13.1.

100

Coarse Bottle Roll Leach Tests

Gold Leach Curves

95

90

85

80

75

70

1 day 3 days 7 days 10 days 15 days

Leach Time

1/2in.

1/4in.

Figure 13.1 - Gold Leach Curves

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3.1.4 Test Work Summary

In addition to the coarse bottle roll leach tests, Composite 1 was also subjected

to gravity separation test work and bottle roll tests on ground samples. A summary of all

the test results is shown in Table 13.4.

Table 13.4

Test Work Results Summary

Liberation Size P 80 1/2" 1/4" ~150µm ~100µm ~75µm

Heap Leach Simulation (15 days)

Gold Recovery, % 94.7 96.0

Residue Assay, g/t Au 0.06 0.04

Gravity Separation

Gold Recovery, % 8.5

Gravity Separation & Cyanidation (48h)

Gold Recovery, % 94.7 95.5 94.5

Residue Assay, g/t Au 0.06 0.05 0.06

Whole Ore Cyanidation (48h)

Gold Recovery, % 97.3 94.7 96.5

Residue Assay, g/t Au 0.03 0.06 0.04

The following observations can be made:





For all tests a gold recovery of 90 percent was established after only 8 hours,

and reached 95 percent after 24 hours, both with modest to moderate sodium

cyanide consumptions.

The results suggested that the mineralization is amenable to heap leaching

and conventional whole ore cyanidation. Given the minor difference in the

gold leach recoveries between crushed and ground samples, heap leaching

would be the preferred process option.

The recovery of gold was in the range of 96 to 97 percent, leaving a residue

assay of 0.03 to 0.06 g/t gold.

The reagent consumptions were very low, below 0.1 kg/t sodium cyanide and

0.3 kg/t lime.

13.2 SGS Mineral Services UK Ltd

During 2009 Lydian engaged SGS Mineral Services UK Ltd to conduct further

test work, focusing on coarser fractions and lower cyanide concentration solution

concentrations than previous test work and included column leach tests on large

fraction half-core.

The test work was conducted on three master composites (labeled A, B and C) of

half drill core samples from different parts of the Tigranes and Artavasdes areas. This

half core was from drill hole numbers DDA-001, 003, 006, 007, 008, 009, 015, 016, 017,

019 and 020.

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The composites were differentiated by alteration style and by gold and multielement

distribution. The three composites had head grades ranging from 1.09 to 1.29

g/t gold.

Metallic screen analyses of the composites show that greater than 98 percent of

the gold reported to the minus 106 micron fraction. The results confirm the observations

made in previous work indicating that a gravity concentration step is not warranted with

the absence of a significant coarse gold component.

13.2.1 Bottle Roll Leach Tests

Cyanidation bottle roll leach tests were conducted at two size distributions, minus

75 microns and minus 2 mm, to approximate conventional carbon-in-leach (CIL) and

heap leach technologies. The results are shown in Tables 13.5 and 13.6.

Table 13.5

Minus 75 µm Bottle Roll Leach Tests

Leach Period,

Gold Recovery, (%)

(h) Comp A Comp B Comp C

24 83.7 81.8 80.8

48 96.2 90.2 89.1

Cyanide and lime consumptions at minus 75 µm ranged from 0.05 to 0.10 kg/t

and 1.13 to 1.32 kg/t, respectively.

Table 13.6

Minus 2 mm Bottle Roll Leach Tests

Leach Period,

Gold Recovery, (%)

(d)

Comp A Comp B Comp C

1 89.1 81.2 78.2

14 95.1 91.8 89.2

Cyanide and lime consumptions at minus 2 mm ranged from 0.08 to 0.09 kg/t

and 1.06 to 1.20 kg/t, respectively.

13.2.2 Column Leach Tests

Column leach tests were carried out at crush sizes of minus 38 mm and minus

19 mm for leach cycles of 144 and 72 days, respectively.

The results of the column leach tests at a crush size of minus 38 mm are shown

in Table 13.7.

Table 13.7

Minus 38 mm Column Leach Tests

Leach Period,

Gold Recovery, %

d

Comp A Comp B Comp C

70 56.7 71.0 53.1

144 68.5 80.3 64.4

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%Au Recovery

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Cyanide and lime consumptions ranged from 0.18 to 0.31 kg/t and 0.63 to 0.97

kg/t, respectively. The column leach curves for the different composites at crush size of

minus 38 mm are shown in Figure 13.2.

100

Column Leach Curves (-38mm)

90

80

70

60

50

40

30

20

10

0

0 20 40 60 80 100 120 140 160

Leach Cycle (Days)

Comp A Comp B Comp C

Figure 13.2 - Column Leach Curves (-38 mm)

Results of the column leach tests at a crush size of minus 19 mm are shown in

Table 13.8.

Table 13.8

Minus 19 mm Column Leach Tests

Leach Period,

Gold Recovery, %

d

Comp A Comp B Comp C

35 86.0 85.1 73.0

72 89.1 88.6 76.5

Cyanide and lime consumptions ranged from 0.10 to 0.13 kg/t and 0.90 to 1.14

kg/t, respectively. The column leach curves for the different composites at crush size of

minus 19 mm are shown in Figure 13.3.

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%Au Recovery

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100

Column Leach Curves (-19mm)

90

80

70

60

50

40

30

20

10

0

0 10 20 30 40 50 60 70 80

Leach Cycle (days)

Comp A Comp B Comp C

Figure 13.3 - Column Leach Curves (-19 mm)

For all three composites, higher gold recoveries were attained at a crush size of

minus 19 mm than at minus 38 mm. Leach kinetics were also rapid with 80 percent of

the gold leached after 20 days at the minus 19 mm crush size.

A review of all final gold recovery results for all tests showed that Composite A

produced the highest level of gold recovery in all but the minus 38 mm column leach

test. These results are summarized in Table 13.9.

Table 13.9

Final Gold Recovery Summary by Test and Composite

Sample Fineness

Gold Recovery, %

Size and Test Comp A Comp B Comp C

80% -75µm bottle roll 95.8 95.2 93.2

-2 mm bottle roll 95.1 91.8 89.2

-19 mm column 89.1 88.6 76.5

-38 mm column 68.5 80.3 64.4

These initial scoping test work results suggest attractive processing economics of

the Amulsar project. Bulk mining of low-grade ore with a leach operation requiring three

stages of crushing is feasible.

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13.3 Wardell Armstrong International (2010)

In 2010, Lydian commissioned WAI to undertake a further program of laboratory

testing on the Composite A and Composite B samples originally tested by SGS. The

test work generally focused on leaching at finer crush sizes and using higher cyanide

concentrations than were used in the SGS test work.

13.3.1 Column Leach Tests

Column test work was conducted using cyanide concentrations of 0.075, 0.050

and 0.025 percent. The crush sizes were 38, 25, 18 and 12 millimeters. The columns

were irrigated at a rate of 10 l/m 2 /h and the leach period was 68 days.

The column leach test results are given in Table 13.10 and the column leach

recovery curves in Figure 13.4.

Table 13.10

Column Leach Test Results Summary

Sample

Crush Size, NaCN Concentration, Gold Recovery,

mm

%

%

A 25 0.05 91.9

A 19 0.05 93.5

A 12 0.05 94.8

B 38 0.05 88.6

B 25 0.05 88.6

B 25 0.075 89.1

B 19 0.025 89.2

B 19 0.05 93.1

B 19 0.075 92.3

B 12 0.025 89.3

B 12 0.05 90.7

B 12 0.075 94.9

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100.00%

Column Leach Recovery Curves

90.00%

80.00%

70.00%

60.00%

50.00%

40.00%

30.00%

20.00%

10.00%

0.00%

0 3 4 5 6 7 10 11 12 13 14 17 18 19 20 21 23 24 25 26 27 30 37 44 51 58 65 68

Leach Time (Days)

25mm 0.5gpl 38mm 0.5gpl 25mm 0.75gpl 19mm 0.25gpl 19mm 0.5gpl

19mm 0.75gpl 12mm 0.25gpl 12mm 0.5gpl 12mm 0.75gpl

Figure 13.4 - Column Leach Curves

(Various Crush Sizes and Cyanide Concentrations)

The test results indicate that for both samples the optimum crush size may be 19

mm and the optimum cyanide concentration 0.05 percent, although, further work will be

required to substantiate these leaching parameters.

Tests using the higher cyanide concentrations also gave higher cyanide

consumptions and the additional gold recovery needs to be related to the additional

cyanide costs. The same is true for the additional costs of crushing to the finer sizes.

13.4 Wardell Armstrong International (2011)

In late 2010 Lydian shipped full core from three diamond drill holes carried out in

each of the deposits, Tigranes, Artavasdes and Erato, to WAI for coarse bottle roll and

column leach test work.

13.4.1 Bulk Density, Specific Gravity and Work Indices

Samples of the drill core from each of the three deposits were dried and crushed

for determination of crushed ore bulk density, by volumetric flasks, and rock specific

gravity, by pycnometry, at Phillips Enterprises, LLC. The results are shown in Table

13.11.

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Deposit/Drill

Hole ID

Erato

DDAM-068

Atravasdes

DDAM-070

Tigranes

DDAM-071

Table 13.11

WAI – Bulk Density and Specific Gravity

Dry Bulk

Interval,

Density,

m

t/m 3

Dry Specific

Gravity

21.4 – 26.2 1.32 2.71

44.9 – 50.0 1.44 2.94

101.8 – 107.0 1.28 2.91

41.5 – 46.0 1.16 3.01

106.0 – 111.8 1.35 3.06

127.0 – 133.8 1.28 2.82

54.1 – 58.9 1.37 2.80

90.2 – 95.5 1.40 2.89

119.0 – 124.4 1.16 3.03

Phillips also conducted crusher work index and abrasion testing on samples of

the drill core. Two of the core intervals for DDAM-070 did not contain core competent

enough for work index and abrasion testing.

Deposit/Drill

Hole ID

Erato

DDAM-068

Atravasdes

DDAM-070

Tigranes

DDAM-071

Table 13.12

WAI – Bulk Density and Specific Gravity

Interval, Crusher Work Index, Abrasion

m

kWh/mt

Index

21.4 – 26.2 18.78 0.2716

44.9 – 50.0 11.61 0.2519

101.8 – 107.0 9.06 0.1348

41.5 – 46.0 N/A N/A

106.0 – 111.8 N/A N/A

127.0 – 133.8 12.53 0.1885

54.1 – 58.9 9.16 0.4908

90.2 – 95.5 8.27 0.3786

119.0 – 124.4 3.39 0.0080

The Atravasdes and Tigranes crusher work indices are considered to be

moderate, and are harder for Erato.

13.4.2 Coarse Bottle Roll Leach Tests

To determine variability in gold recovery arising from rock type, and crush size,

coarse bottle roll leach tests were conducted on individual drill hole intervals at crush

sizes of 100% minus 12 mm and 19 mm.

Results of bottle roll leach tests are shown in Table 13.13 for gold and in Table

13.14 for silver.

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Deposit DDH# Rock Type Description

Table 13.13

Coarse Ore Bottle Roll Leach Test Results (Gold)

Sample

(m)

Head Assay

Au

(g/t)

% Gold Recovery

% Var

Avg. Deposit

% Au Rec

Au Rec

-12 mm -19 mm -12 mm -19 mm

Erato DDAM 068

Artavasdes DDAM 070

FeOx-banded volcanics, slight leaching 62.7-64.2 0.02 - -

Strongly FeOx-banded volcanic 71.3-73.0 1.19 95.4 95.9 -0.5

Pervasive FeOx throughout silicified breccia 85.5-86.4 2.89 98.6 97.0 1.6

Pervasive FeOx throughout silicified breccia 97.9-98.7 0.02 - -

Heavily leached + FeOx gossan 85.1-86.4 0.02 - -

Fault gouge zone 89.5-90.6 2.80 92.3 86.2 6.1

Heavily leached + FeOx gossan 93.9-94.8 14.09 79.4 73.3 6.1

Fault gouge zone + SM bx clasts 99.0-100.0 0.31 80.9 81.9 -1.0

Fault gouge/gossanous zone 104.5-105.9 8.32 93.4 89.6 3.8

Fault zone with clay gouge, host SM volcanics 41.0-42.0 0.36 92.7 93.4 -0.7

Fault zone with clay gouge, host SM volcanics 48.4-49.3 1.09 85.9 92.2 -6.3

97.0 96.5

86.5 82.8

SM volcanics, numerous small-scale (


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Table 13.14

Coarse Ore Bottle Roll Leach Test Results (Silver)

Head

Deposit DDH# Rock Type Description

(g/t)

Sample Assay

(m)

Ag

FeOx-banded volcanics, slight leaching 62.7-64.2 - -

% Silver

Recovery % Var

Avg. Deposit

% Ag Rec

-12 mm -19 mm

Ag Rec

-12 mm -19 mm

Erato DDAM 068

Artavasdes DDAM 070

Strongly FeOx-banded volcanic 71.3-73.0 175 28.6 43.1 -14.5

Pervasive FeOx throughout silicified breccia 85.5-86.4 1.10 56.8 48.6 8.2

Pervasive FeOx throughout silicified breccia 97.9-98.7 - -

Heavily leached + FeOx gossan 85.1-86.4 - -

Fault gouge zone 89.5-90.6 1.85 54.6 45.14 9.5

Heavily leached + FeOx gossan 93.9-94.8 3.25 37.2 42.77 -5.6

Fault gouge zone + SM bx clasts 99.0-100.0 1.40 57.1 42.86 14.2

Fault gouge/gossanous zone 104.5-105.9 6.25 27.0 14.48 12.5

Fault zone with clay gouge, host SM volcanics 41.0-42.0 1.50 62.3 60.3 2

Fault zone with clay gouge, host SM volcanics 48.4-49.3 1.35 48.1 52.2 -4.1

42.7 45.9

44.0 36.3

SM volcanics, numerous small-scale (


%Au Leach Recovery

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The bottle roll leach tests show the previous trend of increasing gold leach

recovery with finer crush size. On average a two percent recovery gain was realized by

crushing from 100 percent minus 19 mm to minus 12 mm.

The gold recoveries for the minus 12 mm tests are plotted against sample

number in Figure 13.5.

100

Coarse Bottle Roll Leach Tests (-12mm)

Amulsar Deposit Types

97.5

95

92.5

90

87.5

85

82.5

80

77.5

75

0 2 4 6 8 10 12 14

Sample

Erato Artavasdes Tigranes

Figure 13.5 - Coarse Bottle Roll Leach Recoveries (-12 mm)

The tests indicate that there is minimal metallurgical variability with respect to

gold recovery as a function of rock and deposit type.

Figure 13.6 plots the gold leach recoveries as a function of head grade. It is

evident that there is no influence of head grade upon gold leach recovery. The two low

leach recovery points, less than 82.5 percent, were for samples from the Artavasdes

deposit. One of the samples had a high gold grade of 14 g/t and thus potentially could

have had insufficient leach residence time, whilst the second sample had a low head

grade of 0.30 g/t.

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%Au Leach Recovery

Lydian International - Amulsar Resource Update and Heap Leach Feasibility Study Page 98

100

Coarse Bottle Roll Leach Tests

Effect of Head Grade on Leach Recovery

97.5

95

92.5

90

87.5

85

82.5

80

77.5

75

0.00 2.00 4.00 6.00 8.00 10.00 12.00 14.00 16.00

Head Grade (g/tAu)

Erato Artavasdes Tigranes

Figure 13.6 - Effect of Head Grade on Gold Leach Recovery

The preceding tabulation indicates silver extraction averages 44 percent from a

2.21 g/t silver head. This work has not been optimized as maximum silver extraction

usually requires higher cyanide concentrations, or finer liberation size than customarily

used for gold leaching.

13.4.3 Column Leach Tests

Column leach test were conducted on samples composited according to rock

type as well as on samples composited according to deposit. Following are the rock

type composites:





Medium Iron Oxide, Siliceous (MPF)

Siliceous Breccia (SB)

Fault Gouge Zone (FG)

Leached and Iron Oxide Rich Gossan, Silicified (GSN)

The results of the column leach tests conducted on the different rock types are

shown in Table 13.15.

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Rock Type

Solution to Ore

Ratio,

kl/t

Table 13.15

WAI – Rock Type Column Tests

Calculated Head

Assay,

g/t

Extraction,

%

Reagent Consumption,

kg/t

Au Ag** Au Ag*** NaCN

*

Ca(OH) 2

MPF 3.59 0.43 1.58 97 56 0.67 2.5

SB 2.96 0.37 2.12 86 67 0.41 2.5

FG 6.41 2.06 1.58 92 37 0.68 2.5

GSN 11.34 3.13 3.25 84 36 2.00 2.5

* Hydrated lime addition to test charge during column loading.

** Silver head assay.

*** Silver extraction based on head and tail assays.

The results indicate that there is minimal metallurgical variability with respect to

rock type, or gold grade, as all four samples attained gold extractions at, or exceeding

84 percent.

Following are the deposit type composites with the results from the deposit type

columns presented in Table 13.16.

Deposit

▪ Erato - MC 068

▪ Artavasdes - MC 070

▪ Tigranes - MC 071

Solution to Ore

Ratio,

kl/t

Table 13.16

WAI - Rock Type Column Tests

Calculated Head

Assay,

g/t

Extraction,

%

Reagent Consumption,

kg/t

Au Ag** Au Ag*** NaCN

*

Ca(OH) 2

Erato 2.66 0.85 1.81 98 64 0.48 2.5

Artavasdes 2.98 0.76 2.24 95 77 0.46 2.5

Tigranes 3.26 1.49 2.39 89 57 0.46 2.5

* Hydrated lime addition to test charge during column loading.

** Silver head assay.

*** Silver extraction based on head and tail assays.

Gold extraction was high, on average 94 percent, from all three deposit type

composite samples.

13.5 Kappes, Cassiday & Associates (KCA) (2011/2012)

In October 2011 bulk samples from surface outcrops, 1/2 split core, and whole

core was delivered to KCA for metallurgical testing. The testwork program entailed

sample preparation, physical test work, head analyses, coarse bottle roll tests and

column tests. The test results are detailed in a KCA report titled "Amulsar Project Report

of Metallurgical Test Work", Report I.D. KCA0120006, March, 2012, and are

summarized below.

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13.5.1 Sample Preparation

Fifteen composite samples were generated at KCA by combining sample bags

and drill core intervals according to instructions from Amulsar. Each composite sample

was stage crushed to the following approximate size distribution:




P 100 12.5 mm

P 80 9.5 mm

P 60 6.3 mm

The composite samples for leach and physical testing are identified in Tables

13.17 and 13.18 respectively.

Table 13.17

KCA - Composite Samples For Head Analyses and Leach Testing

Description Material Type KCA Sample No.

Tigranes – TM 1-18 Bulk 61723

Artavasdes – ATM 1-26 Bulk 61724

Tigranes – DDA-018 ½ Split Core 61730

Artavasdes – DDA-022 and DDA-055 ½ Split Core 61731

Artavasdes – DDA-033 ½ Split Core 61732

Artavasdes – DDA-035 and DDA-055 ½ Split Core 61733

Tigranes – DDA-046 and DDA-076 ½ Split Core 61734

Tigranes – DDA-055 ½ Split Core 61735

Tigranes – DDA-076 ½ Split Core 61736

Tigranes – DDAM-130 Whole Core 61768

Tigranes – DDAM-137 Whole Core 61769

Artavasdes – DDAM-140 Whole Core 61770

Artavasdes – DDAM-148 Whole Core 61771

Artavasdes – DDAM-169 Whole Core 61772

Tigranes – DDAM-174 Whole Core 61773

Table 13.18

KCA - Composite Samples for Physical Test Work

Description

Log

KCA Sample

No.

Tigranes/Artavasdes – DDAM-130, 137, 140, 148, 174 Breccia 61726

Tigranes/Artavasdes – DDAM-130, 137, 140, 148, 169, 174

Hematite Stained

Vuggy Silica

61727

Tigranes/Artavasdes – DDAM-130, 140, 148, 174 Porphyry 61728

Tigranes/Artavasdes – DDAM-130, 137, 140, 148, 169 Vuggy Silica 61729

13.5.2 Physical Test Work

Four composite samples of cut whole core were submitted to Phillips Enterprises,

LLC for determination of abrasion, crush work and ball mill work indices.

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KCA Sample

No.

13.5.3 Head Assay Analyses

Table 13.19

KCA - Abrasion and Work indices

Crush Work

Abrasion

Index,

Index

kWh/mt

Ball Mill Work

Index,

kWh/mt

61726 0.4376 9.16 13.71

61727 0.4595 9.72 14.02

61728 0.1349 6.75 10.72

61729 0.4673 11.25 14.91

Pulverized subsamples of each of the composites generated for leach testing

were subjected to head assay analyses. The results of total and cyanide soluble gold

and silver assays are shown in Table 13.20. Soluble gold and silver averaged 95 and 57

percent respectively.

Table 13.20

KCA – Gold and Silver Head Assays

KCA

Gold, g/t

Silver, g/t

Sample No. Total NaCN Soluble Total NaCN Soluble

61723 4.64 4.56 12.94 4.39

61724 0.66 0.59 7.41 3.47

61730 0.61 0.58 1.30 0.64

61731 0.45 0.38 1.41 1.14

61732 0.96 0.95 2.19 0.39

61733 1.03 1.03 11.50 0.23

61734 1.63 1.48 3.91 0.76

61735 0.93 0.94 2.61 0.85

61736 2.85 2.76 0.99 0.31

61768 1.34 1.13 0.70 0.59

61769 1.60 1.48 0.79 0.63

61770 1.61 1.49 5.01 2.89

61771 0.78 0.83 4.30 1.14

61772 0.48 0.48 4.70 10.23

61773 0.83 0.72 4.49 3.49

Total and cyanide soluble copper, mercury, carbon and total and sulfide sulfur

assays are shown in Table 13.21. While copper solubility is 20 percent or less, the level

of soluble copper in the leach solution will likely attain or exceed the levels of gold and

silver. No provisions have been incorporated into the process design to manage

cyanide soluble copper though the operational practices of maintaining high cyanide

concentration before carbon adsorption and cold stripping of copper from loaded carbon

should be considered if required.

Mercury assays on the head samples were at, or below the detection limit though

in all locked cycle column leach tests mercury was detected on the loaded carbon.

Leached and adsorbed mercury will be managed in the gold room using a retort furnace

to volatilize, condense and capture the metal in elemental form.

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Low levels of carbon and sulfide sulfur predict minimal metallurgical concerns of

“preg-robbing” and elevated cyanide consumption of which neither were observed in the

column leach tests.

Table 13.21

KCA - Copper, Mercury, Carbon and Sulfur Assays

Copper, g/t

Sulfur, %

KCA

Mercury, Carbon,

Sample No.

NaCN

Total

g/t

%

Total Sulfide

Soluble

61723 473 9.8


Lydian International - Amulsar Resource Update and Heap Leach Feasibility Study Page 103




Sodium cyanide was added to the slurry to a target amount of 1.0 grams per

liter sodium cyanide. The bottle was then placed onto a set of laboratory rolls.

Rolling throughout the duration of the test mixed the slurry.

For the bulk composites and 1/2 split core composites, the slurry was

checked at 2, 4, 8, 24 and 48 hours for pH, dissolved oxygen (DO), NaCN,

Au, Ag and Cu. For the whole core composites, the slurry was checked at 2,

4, 8, 24, 48, 72 and 96 hours for pH, dissolved oxygen (DO), NaCN, Au, Ag

and Cu.

After completion of the leach period, the slurry was filtered, washed and dried.

From the dry tailings, duplicate portions were split out and assayed for

residual gold and silver content. The reject material was stored.

The gold and silver extraction results of the bottle roll test are summarized in

Table 13.22.

Table 13.22

KCA - Bottle Roll Tests

Calculated Head Assay, Leach Extraction, Reagent Consumption,

KCA

g/t

%

kg/t

Sample No.

*

Au Ag Au Ag NaCN Ca(OH) 2

61723 4.32 11.97 92 23 0.63 2.0

61724 0.64 4.37 90 73 0.44 1.0

61730 0.59 0.69 98 69 0.17 1.1

61731 0.43 1.28 93 76 0.07 1.1

61732 0.95 0.85 98 28 0.28 2.0

61733 1.18 12.67 96 50 0.17 1.5

61734 1.40 1.78 98 77 0.30 2.5

61735 1.08 1.68 98 75 0.26 1.5

61736 2.54 0.48 97 58 0.31 1.5

61768 1.35 1.21 96 57 0.13 2.0

61769 1.52 1.30 97 53 0.24 1.5

61770 1.49 5.63 95 47 0.40 1.5

61771 0.79 4.33 95 30 0.28 1.5

61772 0.41 10.09 93 96 0.35 1.5

61773 0.78 5.33 87 61 0.56 2.5

* Hydrated lime addition to test.

13.5.5 Column Leach Tests

Column leach tests were conducted utilizing material crushed to 100 percent

passing 12.5 millimeters. The columns were conducted without agglomeration.

During testing, the bulk composites were leached for a period of 75 days with a

sodium cyanide solution. The 1/2 split core composites were leached for a period of 70

days with a sodium cyanide solution. The whole core composites were leached for a

period of 60 days with a sodium cyanide solution.

For the two bulk samples, testing was conducted at initial ore heights of 3.7 and

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3.5 meters. For the drill core samples, initial ore heights varied between 1.8 and 2.1

meters.

All column tests were conducted in locked cycle with carbon adsorption for

precious metal recovery and accounting. The irrigation rate fluctuated between 10 and

12 l/h/m 2 . The initial leach solution contained 1 g/l sodium cyanide and cyanide was

added to the recirculating leach solution to maintain a target level of 0.5 g/l. Hydrated

lime was added to maintain protective alkalinity and initially added to the column test

charge at dosages between 1 and 3 kg/t.

The column test metal extraction and reagent consumption results are shown in

Table 13.23.

KCA

Sample

No.

Solution to

Ore Ratio,

kl/t

Calculated Head

Assay,

g/t

Table 13.23

KCA - Column Tests

Leach Extraction,

%

Reagent Consumption,

kg/t

Au Ag Au Ag NaCN

*

Ca(OH) 2

61723 2.36 4.50 8.72 91 5 0.34 1.5

61724 2.48 0.67 4.11 89 48 0.12 1.0

61730 5.24 0.56 0.52 96 34 0.18 2.0

61731 5.18 0.50 1.21 92 43 0.17 2.0

61732 4.56 0.95 0.76 93 22 0.15 2.5

61733 5.01 1.13 13.21 91 37 0.22 2.5

61734 5.41 1.64 1.30 97 73 0.32 3.1

61735 4.57 1.18 1.44 96 48 0.18 2.0

61736 4.57 2.44 0.47 97 30 0.23 1.5

61768 4.01 1.27 1.35 92 20 0.14 2.0

61769 4.42 1.60 1.16 92 9


Cumulative Gold Extraction

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100%

90%

80%

70%

60%

50%

40%

30%

20%

10%

0%

TM-1 through TM-18 (61738) ATM-1 thru ATM-26 (61741)

DDA-018 (61744) DDA-022 and DDA-055 (61747)

DDA-033 (61750) DDA-035 and DDA-055 (61753)

DDA-046 and DDA-076 (61756) DDA-055 (61759)

DDA-076 (61762) DDAM-130 (61775)

DDAM-137 (61778) DDAM-140 (61781)

DDAM-148 (61784) DDAM-169 (61787)

DDAM-174 (61790)

0 10 20 30 40 50 60 70 80 90 100

Leach Time, days

Figure 13.7 - KCA Columns - Gold Extraction versus Time

In order to scale the column test leach kinetic results to conditions proposed for

industrial processing, refer to Section 13.7, cumulative gold extraction versus

cumulative applied leach solution per tonne of ore is plotted. These rate curves are

shown in Figure 13.8.

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Cumulative Gold Extraction

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100%

90%

80%

70%

60%

50%

40%

30%

20%

10%

0%

TM-1 through TM-18 (61738) ATM-1 thru ATM-26 (61741)

DDA-018 (61744) DDA-022 and DDA-055 (61747)

DDA-033 (61750) DDA-035 and DDA-055 (61753)

DDA-046 and DDA-076 (61756) DDA-055 (61759)

DDA-076 (61762) DDAM-130 (61775)

DDAM-137 (61778) DDAM-140 (61781)

DDAM-148 (61784) DDAM-169 (61787)

DDAM-174 (61790)

0 1 2 3 4 5 6 7

Leach Time, days

Figure 13.8 – KCA Columns – Gold Extraction versus kl/t

Figure 13.8 indicates that an industrial heap leach designed to operate with a

combined primary and secondary leach cycle attaining 2.0 kl/t will extract most of the

gold leaving a small amount for residual leaching in a buried lift (refer to Section 13.7).

The high levels of gold extraction from the KCA tests are consistent across the

two pits and all rock types and gold head grades. This is depicted graphically in Figure

13.9 which plots column test final gold extraction in units of extracted grams of gold per

tonne against the column test calculated gold head assay also in grams per tonne.

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Final Gold Extraction, g/t

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5

Gold Extraction versus Head Assay

4

3

2

1

y = 0.9185x

R² = 0.9958

0

0 1 2 3 4 5

Calculated Gold Head Assay, g/t

KCA Column Tests Linear (KCA Column Tests)

Figure 13.9 – KCA Columns – Gold Extraction Correlation

The strong linear correlation in Figure 13.9 can be used as a predictor for

ultimate gold extraction. Regardless of the source of the ore, gold extraction of

approximately 92 percent will be obtained in a column leach test. The method of

predicting gold extraction for the industrial heap leach is discussed in Section 13.7.

13.5.6 Gold and Silver Accountability

Four different gold and silver head assays for each of the leach composites are

presented in Table 13.24 and 13.25. These head assays are as follows:





Assay of overall split from head sample

Weighted average assay of head screened fractions

Back calculated assay from bottle roll test

Back calculated assay from column test

The excellent agreement between these four assays for gold indicates sound

sample preparation procedures, mass accounting and assays, and diminishes the

likelihood of sampling/assaying bias due to the presence coarse gold.

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KCA

Sample No.

Head

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Table 13.24

KCA – Gold Head Assays

Gold Head Assay, g/t

Weighted

Screened

Fractions

Bottle Roll

Back

Calculated

Column

Test Back

Calculated

Mean,

g/t

Statistics

Relative

Standard

Deviation, %

61723 4.64 4.47 4.32 4.50 4.48 2.9

61724 0.66 0.65 0.64 0.67 0.66 2.0

61730 0.61 0.53 0.59 0.56 0.57 6.1

61731 0.45 0.50 0.43 0.50 0.47 7.6

61732 0.96 1.00 0.95 0.95 0.97 2.5

61733 1.03 1.06 1.18 1.13 1.10 6.2

61734 1.63 1.50 1.40 1.64 1.54 7.4

61735 0.93 1.04 1.08 1.18 1.06 9.8

61736 2.85 2.41 2.54 2.44 2.56 7.9

61768 1.34 1.31 1.35 1.27 1.32 2.7

61769 1.60 1.56 1.52 1.60 1.57 2.4

61770 1.61 1.40 1.49 1.38 1.47 7.1

61771 0.78 0.73 0.79 0.76 0.77 3.5

61772 0.48 0.46 0.41 0.45 0.45 6.5

61773 0.83 0.76 0.78 0.76 0.78 4.2

Poorer accountability is observed for silver as fire assaying for silver is more

difficult and less precise than for gold.

KCA

Sample No.

Head

Table 13.25

KCA - Silver Head Assays

Silver Head Assay, g/t

Weighted

Screened

Fractions

Bottle Roll

Back

Calculated

Column

Test Back

Calculated

Mean,

g/t

Statistics

Relative

standard

Deviation, %

61723 12.94 10.48 11.97 8.72 11.03 16.7

61724 7.41 6.91 4.37 4.11 5.70 29.9

61730 1.30 1.04 0.69 0.52 0.89 39.4

61731 1.41 1.28 1.28 1.21 1.30 6.4

61732 2.19 2.12 0.85 0.76 1.48 52.8

61733 11.50 10.49 12.67 13.21 11.97 10.2

61734 3.91 3.35 1.78 1.30 2.59 48.1

61735 2.61 2.28 1.68 1.44 2.00 26.8

61736 0.99 1.01 0.48 0.47 0.74 41.1

61768 0.70 0.93 1.21 1.35 1.05 27.7

61769 0.79 1.56 1.30 1.16 1.20 26.7

61770 5.01 6.05 5.63 4.77 5.37 10.9

61771 4.30 3.87 4.33 3.91 4.10 6.0

61772 4.70 6.23 10.09 8.90 7.48 32.9

61773 4.49 6.03 5.33 1.75 4.40 42.6

13.5.7 Column Detoxification

Following leaching, detoxification test work utilizing a hydrogen peroxide (copper

catalyzed peroxide) detoxification method was conducted on two selected column leach

tests.


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A portion of the final barren solution was submitted for total cyanide analysis,

weak acid dissociable (WAD) cyanide analysis and copper analysis. Based upon the

total cyanide and copper analysis, hydrogen peroxide and copper sulfate

(CuSO 4·5H 2 O), if needed, were added to detoxify the solution. The columns were then

restarted with the detoxified solution. Hydrogen peroxide and copper were then added

daily to the leach solution based on the total cyanide analysis. Every seven days

thereafter, the solution was sampled and analyzed for WAD cyanide and copper, and

daily hydrogen peroxide and copper additions were adjusted accordingly. The

detoxification procedure was continued until a WAD cyanide value less than 0.2 mg/l

was obtained for three consecutive days. Complete details of the detoxification test

work are appended.

A summary of the results is presented in Tables 13.26 and 13.27.

Table 13.26

KCA – Column Detoxification – Cyanide Species

Detoxification

KCA

Total Cyanide, ppm

WAD Cyanide, ppm

Period,

Sample No.

d

Initial Final Initial Final

61770 26 126 0.48 114 0.10

61773 26 107 1.15 103 0.08

KCA

Sample No.

Sample Dry

Weight,

kg

Table 13.27

KCA – Column Detoxification – Reagents

35% H 2 O 2 10 g/l CuSO 4·5H 2 0

Added, Added,

ml

ml

100% H 2 O 2

Added,

g/t

Cu Added,

g/t

61770 51.7 66.5 850 450 41.8

61773 56.1 71.5 850 446 38.6

13.6 Metallurgical Sample Drill Hole Locations

Through the five phases of column testing, split drill core, whole drill core and

composite samples have been used for testing. The drill hole identifications for each

phase of testing have been identified in the text of this report introducing each phase.

Figure 13.10 presents a plan of the location of all of the samples subjected to

metallurgical investigation. The samples are shown to be representative of ore within

the proposed starter pit and within the final pit shell.

Samples are also representative of depth within the starter and final pit shells.

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Figure 13.10 - Metallurgical Drill Hole Locations

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13.7 Test Work Interpretation and Process Design

The metallurgical results from all five phases of testing confirm that the Amulsar

gold mineralization is amenable to recovery using heap leach technology. The gold

extraction versus crush size column test investigations completed by SGS and WAI

indicate improvements in gold extraction with finer crush size. At particle size

distribution of approximately 12 mm column test gold extraction approached or

exceeded 90 percent as was confirmed in the KCA test work. A nominal 12 mm crush

size is a reasonable lower limit for processing by heap leaching. At the gold grade of the

Amulsar resource, the incremental increase in gold recovery by grinding the ore (SGS

test work) and processing by carbon-in-leach (CIL) technology does not offset higher

capital and operating costs to improve project economics over heap leaching. A trade

off study was completed during the preparation of this feasibility study resulting in a net

present value for the heap leach option at US$ 300 million above that of a CIL option.

As such, conventional three-stage crushing, followed by stacking, leaching and gold

recovery is proposed.

The high rate and ultimate levels of gold extraction obtained in the KCA column

test work were obtained without percolation problems and without agglomeration. The

proposed plant design does not include an agglomeration drum. Crushed ore will be

conveyed to the leach pad and stacked using a radial stacker to an 8 m height, standard

maximum height for conventional stacking equipment. Irrigation will be with sprinklers to

promote evaporation due to positive net water balance at rates between 8 and 10

l/h/m 2 , also standard to the industry.

The leach pad will be divided into cells from which the pregnant leach solution

can be diverted to either pregnant or intermediate solution ponds. The stacking of the

intermediate leach solution onto the leach pad allows operations to upgrade the

pregnant leach solution gold and silver tenor and for the design of a more efficient

carbon adsorption circuit.

13.7.1 Column Test Scale Up - Gold Extraction

The column test work was conducted at column heights shorter than the 8 m lift

height proposed for the industrial operation. KDE has used a common practice of

scaling column test results using the ratio of applied solution to ore under leach. The

design parameters proposed for the industrial leach stage are summarized in Table

13.28.

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Table 13.28

Heap Leach Design Parameters

Parameter Units Design

Lift Height m 8

Solution Application Rate l/m 2 /h 8-10

Primary Leach Cycle days 30

Primary Cycle Solution:Ore Application kl/t 0.5

Secondary Leach Cycle days 80

Secondary Cycle Solution:Ore Application kl/t 1.5

Total Leach Cycle before Burying Lift days 110

Total Solution:Ore Application before Burying Lift kl/t 2.0

The KCA column test work has been scaled up to these industrial conditions by

evaluating the rate of gold extraction versus applied leach solution to ore ratio (S:O) and

equating this rate of extraction to the S:O ratio attained under the industrial conditions.

The levels of gold extraction obtained from the column tests at S:O ratios

equivalent to those calculated for the industrial operation after the primary and

secondary leach cycles are shown in Table 3.29. The ultimate level of gold extraction

under industrial conditions which includes additional leaching of gold from buried lifts of

ore (residual leaching) is set as the gold extraction obtained at the termination of the

column leach tests.

KCA Sample

No.

Primary Leach Cycle

Table 13.29

KCA - Column Test Leach Kinetics

Primary plus Secondary

Leach Cycle

Primary plus Secondary

plus Residual Leach Cycle

(Final Column Test Results)

S:O,

kl/t

Au Ext.,

%

S:O,

kl/t

Au Ext.,

%

S:O,

kl/t

61723 0.5 80 2.0 89 2.36 91

61724 0.5 84 2.0 88 2.48 89

61730 0.5 86 2.0 92 5.24 96

61731 0.5 84 2.0 90 5.18 92

61732 0.5 82 2.0 91 4.56 93

61733 0.5 61 2.0 89 5.01 91

61734 0.5 89 2.0 96 5.41 97

61735 0.5 90 2.0 96 4.57 96

61736 0.5 85 2.0 95 4.57 97

61768 0.5 73 2.0 88 4.01 92

61769 0.5 81 2.0 89 4.42 92

61770 0.5 72 2.0 84 5.87 85

61771 0.5 83 2.0 87 4.16 89

61772 0.5 80 2.0 89 4.21 92

61773 0.5 64 2.0 70 5.43 75

Average 0.5 80 2.0 89 4.50 91

Au Ext.,

%

On average, 80 percent gold extraction will occur during the first thirty days of

primary cycle leaching, an additional 9 percent during the 80 days of secondary cycle

leaching, and an additional 2 percent will be extracted as solution passes through

buried lifts.

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The levels of laboratory column leach test gold and silver extraction are

downgraded by 3 and 5 percent respectively to reflect the attainable full scale heap

leach extraction due to losses on the side of the heap and due to channeling effects.

Additionally, both gold and silver extraction is downgraded 1 percent to account for

losses in the metal recovery processes.

13.7.2 Prediction of Industrial Gold Production

Gold and silver reporting to doré (metal recovery) have been estimated for each

of the Tigranes, Artavasdes and Erato deposits. The averaged KCA column test results

were used to estimate recoveries for the Tigranes and Artravasdes deposits. The WAI

test work was used to estimate the recoveries for the Erato deposit. These recoveries

and the downgrades discussed in 13.7.1are summarized in Table 13.30.

Deposit

Table 13.30

Heap Leach Metal Recovery by Deposit

Tigranes Artavasdes Erato

Au, Ag, Au, Ag, Au,

% % % % %

Source KCA Test work KCA Test work WAI Test work

Column Test Average Extraction 92.00 35.63 90.14 37.14 97.67 64.10

Discount for Industrial Practice 3.00 5.00 3.00 5.00 3.00 5.00

Heap Leach Average Extraction 89.00 30.63 87.17 32.14 94.67 59.10

CIC/Goldroom Recovery 99 99 99 99 99 99

Recovery to Doré 88.11 30.32 86.27 31.82 93.72 58.51

13.7.3 Reagent Consumption

Through the first four metallurgical test phases reagent consumption has been

low. Sodium cyanide consumption ranged between 0.1 and 0.3 kg/t of ore and lime

consumption between 0.6 and 1.4 kg/t. In the fifth phase of testing, the KCA column test

program final average sodium cyanide consumption was 0.2 kg/t and the average lime

consumption was 2.0 kg/t (Table 13.23).

While reagent consumption can be lower in industrial practice than observed in

column testing, the average levels of reagent consumption from the KCA test work are

proposed for the heap leach without discount.

Ag,

%

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14.0 MINERAL RESOURCE ESTIMATES

The Amulsar deposits of Tigranes, Artavasdes and Erato are comprised of

fracture controlled, discontinuous, higher grade mineralization surrounded by halos of

lower grade mineralization. The nature of this style of mineralization requires tight

control of the higher grade drill intercepts in order to not give them more influence on

the overall grade estimate than they represent.

The mineral resource is based on drilling completed up to and through the 2011

drill season and includes a total of 665 core and reserve circulation (RC) holes. The

mineral resource was originally reported in the IMC report “2012 Mineral Resource

Estimate, Amulsar Gold Project, NI 43-101 Technical Report”, dated January 22, 2012.

The reported mineral resource at that time was sub-divided into indicated and inferred

resource classifications. The classification of the resource has now been revised to

include measured resource.

14.1 Summary

Gold and silver grades in the Amulsar model were estimated using inverse

distance to the eighth power (ID8), spherical search ellipsoids of 50 meters in Tigranes

and Artavasdes and 100 meters in Erato and no internal domain boundaries. The basis

for selection of these criteria is discussed in the following sections.

The mineral resource is summarized in Table 14.1 at several cut-off grades, with

the resource at a 0.40 g/t gold cut-off being the tonnage and grade that was selected to

be reported. Higher and lower cut-off grades are presented to show the distribution of

tonnage and grades. At the lower cut-off grade of 0.20 g/t, the 0.62 g/t gold head grade

for the indicated mineral resource would have a value of about US$ 22.00 per tonne (at

US$ 1300/oz gold price) which is well above the estimates of operating costs (about

US$ 3.00/t processing and US$ 5.00/t mining).

The mineral resource is within a floating cone geometry based on US$ 1300/oz

gold price (no credit for silver) and preliminary estimates of gold recovery and operating

costs provided in January 2012. Additional definition of these estimates were

subsequently done and used for the definition of the mineral reserve discussed in

Section 15. The reported mineral resource represents about 98 percent of the model

contained mineralization.

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Au

Cut-off

g/t

Mt

Measured

Au

g/t

Ag

g/t

Table 14.1

Amulsar Mineral Resource

Sum of

Indicated

Measured &

Indicated

Au Ag

Au Ag

Mt

Mt

g/t g/t

g/t g/t

Inferred

0.75 15.4 1.64 4.97 13.0 1.57 4.78 28.4 1.61 4.88 15.0 1.45 4.56

0.50 27.7 1.18 4.18 23.9 1.13 4.14 51.6 1.16 4.16 26.8 1.08 4.39

0.40 36.5 1.00 3.82 32.2 0.95 3.84 68.8 0.98 3.83 35.5 0.92 4.01

0.30 50.2 0.83 3.45 46.4 0.77 3.48 96.6 0.80 3.47 48.1 0.77 3.65

0.20 70.7 0.66 3.11 71.1 0.59 3.13 141.9 0.62 3.12 73.9 0.59 3.12

Mt

Au

g/t

Ag

g/t

IMC has not completed an exhaustive independent review of all of the data

supplied to it and has assumed that all these data, including the assay, survey and

density data on which the model resource estimates are based, are correct to within

normally-accepted limits of error. Much of this data has been reviewed and reported on

by CSA in its May 19, 2011 report and IMC has relied on this review work plus the

ongoing checks by IMC.

The mineral resources are not subdivided by mineral type (oxide, sulfide etc.)

because it is reported by Lydian that the Amulsar deposit is entirely oxide.

14.2 Drilling and Assaying

The drill hole information of collar coordinates, down hole survey and assay

information were provided to IMC electronically by Lydian. Table 14.2 summarizes basic

drilling and assaying statistics for the assay data base supplied to IMC in December

2011.

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Table 14.2

Drilling and Assaying Statistics

Diamond core Reverse circulation Total

Number of holes 207 458 665

Total drilling, m 26,826 62,957 89,783

Average hole depth, m 129.6 137.5 135.0

Total assay intervals 25,719 56,455 82,174

Average assay interval, m 1.04 1.15 1.09

Assayed for gold 25,718 56,153 81,871

Assayed for silver 24,437 49,156 73,593

Assayed for sulfur 24,437 49,156 73,593

Coded for alteration 25,636 56,212 81,848

Coded for lithology 18,337 3,439 21,776

At the time of resource estimation assaying was essentially complete for gold

and about 90 percent complete for silver. There are 25,278 silver assay intervals (34

percent of total intervals assayed for silver) with silver assays less than or equal to 0.25

g/t of which 16,060 intervals (22 percent) with the assay equal to 0.25 g/t, most likely

representing a default or minimum assay assignment. Almost all assay intervals are

coded for alteration but only 26 percent are coded for lithology. The average assay

interval is 1.09 meters. The data base also included information on percent recovery in

core holes (average 96.4 percent), RQD data (not reviewed), and the results of 757

specific gravity measurements run on core samples.

Figure 14.1 shows drill hole locations. The three clusters of holes define, from

north to south, the Erato, Tigranes and Artavasdes deposits. Holes are mostly inclined

WNW-ESE or SSW-NNE at a nominal angle of 60 degrees. Hole spacing in in the more

densely drilled areas of Tigranes and Artavasdes is approximately 25 meters and in the

range of 30 to 50 meters in central Erato.

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Figure 14.1 Drill Hole Location Map

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14.3 Assay Compositing, Grade Statistics and Grade Caps

The assay data was composited into 10 meter bench composites to match the 10

meter bench height in the model. Composite statistics are compared with assay

statistics for gold and silver in Table 14.3.

Table 14.3

Assay and 10 meter Bench Composite Statistics

Gold g/t

Silver g/t

Number Mean St. Dev. Max. Number Mean St. Dev. Max

Assays 81,871 0.357 1.494 22.40 73,593 2.47 13.89 2240.00

10m Comps 7,827 0.350 0.811 19.45 6,535 2.48 5.82 204.30

Figure 14.2 shows log-transformed probability plots of 10 meter composite gold

and silver grades. Gold distribution is effectively lognormal with no indication of

separate high-grade populations and thus for this mineral resource no grade cap was

applied. Silver shows evidence for a high-grade population, but represents only about 1

percent of the data above 15 g/t. Silver is not a large economic benefit to the economics

at Amulsar, thus no grade cap was applied at this time.

Figure 14.2 - Gold and Silver Probability Plots, 10 m Composites

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14.4 Variograms

Suites of covariance variograms were run on 10 meter gold composites to assist

in determining search parameters for grade estimation. The ranges measured from

these variograms, fitted with a single spherical structure, are listed in Table 14.4. The

directions with question marks did not produce reliable variograms.

Table 14.4

Gold Variogram Ranges, meters

Erato Tigranes Artavasdes All Data

All horizontal directions 95 55 40 45

Vertical 100? 80 ? 70

Omnidirectional 100 45 55 50

North 120 50 50? 45

North 22 East 90 50 50? 45

North 45 East 90 45 ? 40

North 67 East ? 40 40 40

East ? 80 40 50

South 67 East ? ? 40 50

South 45 East 90? ? 50 60

South 22 East 120 ? 50 55

Gold variogram ranges show no clear indication of horizontal or vertical

anisotropy in any of the deposits. Ranges in Tigranes and Artavasdes are similar but

ranges in Erato are approximately twice as long, indicating more continuity of gold

grade in this deposit. This may be in part due to the more widely spaced drilling. Figures

14.3(A), (B), (C), and (D) shows the omnidirectional gold variograms for Erato,

Tigranes, Artavasdes and “All Data”.

Silver covariance variograms show plus or minus 50 meter ranges in the

horizontal direction but no vertical range, suggesting that silver grades have horizontal

continuity but little or no vertical continuity.

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Erato

Figure 14.3(A) - Omnidirectional Covariance Variogram, Gold

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Tigranes

Figure 14.3(B) - Omnidirectional Covariance Variogram, Gold

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Artavasdes

Figure 14.3(C) - Omnidirectional Covariance Variograms, Gold

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All Data

Figure 14.3(D) - Omnidirectional Covariance Variograms, Gold

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14.5 Distribution of Mineralization

Review of 10 meter bench composite plans and sections show that gold

mineralization at Amulsar is erratic with general mineralization trends. Figure 14.4 is an

illustrative plan map showing the locations of all 10m composites that exceed 1.0 g/t

gold in the Tigranes and Artavasdes drill holes.

Figure 14.4 - 10m Drill Hole Composites Exceeding 1 g/t gold, Tigranes & Artavasdes

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Figure 14.5 shows the gold grade-thickness product in gram-meters above a

0.10 g/t cut-off in Tigranes and Artavasdes drill holes. A localized WNW-ESE trend is

visible in the western part of Artavasdes but there are no clear trends in other areas.

Figure 14.5 - Gold Grade-Thickness Product, 0.1 g/t Cut-off, Tigranes & Artavasdes

Blue < 20 gm-m, Green 20-50, Orange 50-100, Red 100-200, Pink > 200

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The distribution of mineralization relative to lithology could not be reviewed

because the:



Lithology codes in the assay data base were incomplete,

Lithology codes in the model supplied to IMC were generalized into only three

lithologic types.

Table 14.5 summarizes 10 meter composite gold grade statistics relative to

composite and model lithology codes provided to IMC in intervals where both codes are

present (model codes were assigned to composite by “dipping” them into the model).

The breccia is the highest-grade unit in both cases but the composite codes show over

three times as many breccia intervals as the model codes. The lithology in the block

model was not used to restrict or influence the grade estimation at this time. The

differences between the model lithology and the drill hole lithology codes need further

refinement prior to the next resource estimate.

Lithology

Table 14.5

10 m Composite Gold Statistics by Lithology Code

10m composite data base codes

Model codes

Number Mean St.Dev. Max. Number Mean St.Dev. Max.

Breccia 1,068 0.44 1.01 14.4 298 0.49 1.03 10.8

Andesite 276 0.10 0.26 3.4 378 0.24 0.99 12.8

Volcanics 401 0.31 0.55 4.1 1,159 0.37 0.88 14.4

The distribution of mineralization relative to alteration was also reviewed. It was

found that alteration codes in the model had been generalized into four different types

and that the two that accounted for 80 percent of the blocks (silica-alunite, 34 percent

and silica, 46 percent) both had a mean gold grade of 0.407 g/t.

14.6 Grade Estimation

The Amulsar model extends from 4396950N to 4400290N, from 559810E to

562610E and from 2440 to 2990 meters elevation. With a model block size of

10x10x10m it contains 280 columns, 334 rows and 55 tiers for a total of 5,143,600

blocks.

Grades in the model were estimated using spherical searches of 50 meters in

Tigranes and Artavasdes and 100 meters in Erato, no internal domain boundaries, a

minimum/maximum of 1/12 - 10 meter bench composites and inverse distance to the

eighth power (ID8). The search distances respect the variogram ranges, and the

spherical searches are appropriate in deposits with isotropic variography, discontinuous

high grade mineralization and limited mineralized trends. ID8 was used because IMC's

experience in properties where mined-model comparisons are available shows that a

higher-power inverse distance operator is usually needed to match the model grade

distribution to the blast hole grade distribution, and that the blast hole grade distribution

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is usually best matched when the ID operator “splits the difference” between ordinary

kriging (OK), which overstates tonnes and understates grade, and nearest-neighbor

polygons (NNP), which understate tonnes and overstate grade.

Table 14.6 compares IMC model contained resources for indicated blocks above

a 0.20 g/t gold cut-off using OK, NNP and ID8. OK gives 39 percent more tonnes at 28

percent lower grade than NNP but the same contained ounces. ID8 effectively splits the

difference between OK and NNP for tonnes and grade and gives the same contained

ounces. However, there is no guarantee that ID8 will match the grade distribution that

will be obtained during mining at Amulsar.

Table 14.6

IMC Model Contained Resources at a 0.20 g/t Gold Cut-off for Indicated Class

Model Million tonnes Gold, g/t Million oz.

OK 164.5 0.538 2.85

NNP 118.6 0.749 2.86

OK/NNP, % +39 -28 0

Average OK & NNP 141.6 0.626 2.85

ID8 142.2 0.618 2.83

The Figure 14.6 probability plot shows grade distributions in indicated blocks in

the OK, NNP and ID8 models. ID8 splits the difference between OK and NNP overall

grade ranges.

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Figure 14.6 Gold Probability Plots, Indicated Blocks Only

Black = OK, Red = NNP, Blue = ID8

Silver grades were estimated in the model using the same search parameters as

were used for gold. However, because silver assaying was only 90 percent complete at


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the time of resource estimation (Table 14.2) approximately 5 percent of the estimated

blocks are assigned gold grades but not silver grades.

Model results are summarized in Figures 14.7 through 14.10, which compare 10

meter composite gold grades and model block grades on the 2850 bench at Tigranes

and Artavasdes and on the 2800 bench at Erato.

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Figure 14.7 10 m Composite Gold Grades, 2850 Bench, Tigranes & Artavasdes

Blue < 0.1 g/t, Green 0.1-0.2, Orange 0.2-0.5, Red 0.5-1.0, Pink > 1.0

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Figure 14.8 IMC Model Gold Grades, 2850 Bench, Tigranes & Artavasdes, Indicated

Blue < 0.1 g/t, Green 0.1-0.2, Orange 0.2-0.5, Red 0.5-1.0, Pink > 1.0

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Figure 14.9 10 m Composite Gold Grades, 2800 Bench, Erato

Blue < 0.1 g/t, Green 0.1-0.2, Orange 0.2-0.5, Red 0.5-1.0, Pink > 1.0

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Figure 14.10 IMC Model Gold Grades, 2800 Bench, Erato, Indicated

Blue < 0.1 g/t, Green 0.1-0.2, Orange 0.2-0.5, Red 0.5-1.0, Pink > 1.0

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14.7 Tonnage Estimates

Resource tonnages were calculated by applying the mean density values for the

three major lithologies given in the May 19, 2011 CSA Global 43-101 report to the

corresponding “lith” codes in the block model supplied to IMC. The mean densities are:

▪ Breccia 2.33 g/cc

▪ Volcanics 2.41 g/cc

▪ Porphyritic Andesite 2.24 g/cc

14.8 Resource Classifications

Mineral resources in Tigranes and Artavasdes are classified as; measured with

seven or more holes within the search ellipse, indicated where there are three or more

holes within the search ellipse and as inferred where there are fewer than three. This

criterion gives visually reasonable results, with coherent groups of indicated blocks

located internal to the drilling and inferred blocks on the periphery (Figure 14.12).

The three-hole criterion as the break between indicated and inferred is also

supported by Kriging variances, which are a measure of the errors to which individual

block grade estimates are subject. Figure 14.11, which plots Kriging variance against

the number of drill holes in the search, shows Kriging variance increasing gradually as

the number of holes within the search decreases from seven to three but increasing

more rapidly as the number falls below three. This inflection defines the three-hole

minimum as an appropriate confidence threshold for segregating inferred and indicated

material. The inflection point at seven holes was selected for the split between

measured and indicated classification.

Figure 14.11 Kriging Variance Versus Number of Holes in Search Ellipsoid

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Resources in Erato, where grades were estimated using a 100m search, are

classified using the same criterion with two additional constraints imposed:



No measured or indicated blocks further than 50m from the closest hole

A minimum of two composites required to define inferred material

These constraints were imposed to limit the distance to which resources are

projected outside the limits of drilling in the Erato deposit.

Figure 14.12 Mineral Resource Classification, 2850 Bench, Tigranes & Artavasdes

Measured (Red), Indicated (orange), Inferred (green)

Figures 14.13 and 14.14 show the grade-thickness product of measured -

indicated and inferred blocks in the mineral resource model.

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Figure 14.13 Grade-thickness (gram-meters) of Measured-Indicated Blocks,

All Deposits

Blue < 20, Green 20-50, Orange 50-100, Red 100-200, Pink > 200

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Figure 14.14 Grade-thickness (gram-meters) of Inferred Blocks, All Deposits

Blue < 20, Green 20-50, Orange 50-100, Red 100-200, Pink > 200

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14.9 Mineral Resource Tabulations

Table 14.7 lists IMC reported mineral resources for the combined deposits and

Table 14.8 shows the mineral resources split between Tigranes, Artavasdes and Erato.

Table 14.7

Amulsar Mineral Resource for All Areas

Sum of

Au

Measured

Indicated

Measured &

Inferred

Cut-off

Indicated

g/t

Au Ag

Au Ag

Au Ag

Au Ag

Mt

Mt

Mt

Mt

g/t g/t

g/t g/t

g/t g/t

g/t g/t

0.75 15.4 1.64 4.97 13.0 1.57 4.78 28.4 1.61 4.88 15.0 1.45 4.56

0.50 27.7 1.18 4.18 23.9 1.13 4.14 51.6 1.16 4.16 26.8 1.08 4.39

0.40 36.5 1.00 3.82 32.2 0.95 3.84 68.8 0.98 3.83 35.5 0.92 4.01

0.30 50.2 0.83 3.45 46.4 0.77 3.48 96.6 0.80 3.47 48.1 0.77 3.65

0.20 70.7 0.66 3.11 71.1 0.59 3.13 141.9 0.62 3.12 73.9 0.59 3.12

The resource at a 0.40 g/t gold cut-off is the tonnage and grade that was

selected to be reported. Higher and lower cut-off grades are presented to show the

distribution of tonnage and grades. At the lower cut-off grade of 0.20 g/t, the 0.62 g/t

gold head grade for the measured-indicated mineral resource would have a value of

about US$ 22.00 per tonne (at US$ 1300/oz gold price) which is well above the

estimates of operating costs (about US$ 3.00/t processing and US$ 6.00/t mining).

These are within an open pit type geometry based on US$ 1300/oz gold price,

preliminary cost estimates and a gold recovery of 88 percent for all deposits (silver was

not included to define the resource pit geometry). Reporting within a conceptual pit

geometry assures a reasonable expectation of extraction as defined by the CIM

resource reporting standards. At the reported 0.40 g/t gold cut-off grade, the reported

resource is about 98 percent of the model contained resource.

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Classification /

Deposit

Table 14.8

Amulsar Mineral Resource by Classification and Deposit

>= 0.75 g/t Au >= 0.50 g/t Au >= 0.40 g/t Au >= 0.30 g/t Au >= 0.20 g/t Au

ktonnes Au, g/t Ag, g/t ktonnes Au, g/t Ag, g/t ktonnes Au, g/t Ag, g/t ktonnes Au, g/t Ag, g/t ktonnes Au, g/t Ag, g/t

Measured

Artavasdes 6,813 1.67 7.00 11,731 1.22 5.95 15,148 1.05 5.43 19,738 0.89 4.96 26,361 0.73 4.56

Tigranes 4,281 1.69 2.67 7,522 1.22 2.49 9,991 1.03 2.39 13,498 0.85 2.27 18,470 0.69 2.12

Erato 4,294 1.56 4.04 8,414 1.10 3.23 11,391 0.93 2.94 16,982 0.73 2.64 25,914 0.57 2.33

Total 15,388 1.64 4.97 27,667 1.18 4.18 36,530 1.00 3.82 50,218 0.83 3.45 70,745 0.66 3.11

Contained Oz. 812,223 2,458,574 1,052,918 3,720,106 1,179,710 4,489,010 1,332,129 5,574,170 1,496,023 7,064,983

Indicated

Artavasdes 6,287 1.52 5.50 12,454 1.07 4.81 17,159 0.90 4.50 24,227 0.74 4.18 35,763 0.58 3.82

Tigranes 4,028 1.59 4.83 8,043 1.09 3.69 11,039 0.92 3.28 16,015 0.74 2.91 24,709 0.57 2.68

Erato 2,665 1.66 3.01 3,393 1.44 2.73 4,033 1.28 2.55 6,140 0.96 2.20 10,653 0.66 1.85

Total 12,980 1.57 4.78 23,890 1.13 4.14 32,231 0.95 3.84 46,382 0.77 3.48 71,125 0.59 3.13

Contained Oz. 654,856 1,995,158 866,241 3,177,999 985,835 3,977,352 1,143,041 5,188,583 1,339,270 7,155,028

Sum M&I

Artavasdes 13,100 1.59 6.28 24,185 1.14 5.36 32,307 0.97 4.94 43,965 0.80 4.53 62,124 0.64 4.13

Tigranes 8,309 1.64 3.72 15,565 1.15 3.11 21,030 0.97 2.86 29,513 0.79 2.62 43,179 0.62 2.44

Erato 6,959 1.60 3.65 11,807 1.19 3.09 15,424 1.02 2.84 23,122 0.79 2.52 36,567 0.59 2.19

Total 28,368 1.61 4.88 51,557 1.16 4.16 68,761 0.98 3.83 96,600 0.80 3.47 141,870 0.62 3.12

Contained Oz. 1,467,079 4,453,732 1,919,159 6,898,105 2,165,545 8,466,362 2,475,170 10,762,753 2,835,293 14,220,011

Inferred

Artavasdes 3,665 1.27 7.22 9,155 0.87 4.96 12,954 0.75 4.48 18,558 0.63 4.06 26,910 0.51 3.76

Tigranes 4,540 1.34 4.01 7,654 1.04 4.73 10,188 0.89 4.21 14,213 0.74 3.73 21,880 0.57 3.29

Erato 6,769 1.62 3.49 9,977 1.29 3.60 12,360 1.13 3.36 15,334 0.98 3.08 25,137 0.69 2.28

Total 14,974 1.45 4.56 26,786 1.08 4.39 35,502 0.92 4.01 48,105 0.77 3.65 73,927 0.59 3.12

Contained Oz. 696,108 2,195,625 926,715 3,778,717 1,051,949 4,580,105 1,191,962 5,645,394 1,394,625 7,410,191

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Table 14.9 lists the model contained mineralization for all Amulsar deposits by

cut-off and measured-indicated and inferred resource classes. The model contained

mineralization is presented here for comparative purposes and is not the final resources

reported for Amulsar, which are within a conceptual open pit geometry based on US$

1300/oz gold price and preliminary cost and recovery estimates. This was done to

assure a reasonable expectation of extraction as defined by the CIM resource reporting

standards. At the reported 0.40 g/t gold cut-off grade, the reported resource is about

98 percent of the model contained resource. This is due in part to the geometry of the

deposits with higher grade structures surrounded by lower grade mineralization. (It

should be noted that approximately 10 percent of inferred blocks are assigned gold but

not silver grades. It is assumed in Tables 14.9 through 14.12 that the mean silver grade

in these unassigned blocks is the same as the mean silver grade in the assigned

blocks.)

Table 14.9

Model Contained Mineralization, All Deposits

Cut-off g/t Au Million tonnes Au g/t Ag g/t Million oz. Au Million oz. Ag

Measured - Indicated

0.20 142.2 0.618 3.12 2.83 14.3

0.25 117.4 0.701 3.28 2.65 12.4

0.30 96.2 0.795 3.47 2.46 10.7

0.35 80.5 0.887 3.65 2.30 9.4

0.40 68.3 0.979 3.84 2.15 8.4

Inferred

0.20 93.7 0.526 3.12 1.58 9.4

0.25 67.1 0.647 3.57 1.40 7.7

0.30 53.2 0.745 3.93 1.27 6.7

0.35 44.3 0.829 4.20 1.18 6.0

0.40 37.7 0.909 4.42 1.10 5.3

Tables 14.10, 14.11 and 14.12 summarize resources in Erato, Tigranes and

Artavasdes respectively. Coordinate limits for tabulation are: Erato, north of 4399000N;

Tigranes 4398000N to 4399000N; Artavasdes south of 4398000N.

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Table 14.10

Model Contained Mineralization in Erato

Cut-off g/t Au Million tonnes Au g/t Ag g/t Million oz. Au Million oz. Ag

Measured - Indicated

0.20 36.6 0.581 2.31 0.68 2.7

0.25 29.3 0.671 2.44 0.63 2.3

0.30 22.7 0.787 2.57 0.57 1.9

0.35 18.2 0.898 2.71 0.53 1.6

0.40 15.0 1.011 2.87 0.49 1.4

Inferred

0.20 39.5 0.544 2.20 0.69 2.8

0.25 23.7 0.761 2.80 0.58 2.1

0.30 18.5 0.900 3.06 0.54 1.8

0.35 15.7 1.001 3.28 0.51 1.7

0.40 13.6 1.096 3.45 0.48 1.5

Table 14.11

Model Contained Mineralization in Tigranes

Cut-off g/t Au Million tonnes Au g/t Ag g/t Million oz. Au Million oz. Ag

Measured - Indicated

0.20 43.6 0.615 2.34 0.86 3.3

0.25 36.0 0.698 2.46 0.81 2.8

0.30 29.7 0.790 2.58 0.75 2.5

0.35 24.9 0.878 2.70 0.70 2.2

0.40 21.1 0.970 2.83 0.66 1.9

Inferred

0.20 25.2 0.536 3.23 0.43 2.6

0.25 19.9 0.619 3.51 0.40 2.2

0.30 15.4 0.720 3.91 0.36 1.9

0.35 12.7 0.806 4.18 0.33 1.7

0.40 10.8 0.880 4.37 0.31 1.5

Table 14.12

Model Contained Mineralization in Artavasdes

Cut-off g/t Au Million tonnes Au g/t Ag g/t Million oz. Au Million oz. Ag

Measured - Indicated

0.20 61.9 0.640 4.14 1.27 8.2

0.25 52.1 0.719 4.34 1.20 7.3

0.30 43.8 0.803 4.54 1.13 6.4

0.35 37.4 0.887 4.75 1.07 5.7

0.40 32.2 0.969 4.95 1.00 5.1

Inferred

0.20 29.0 0.492 4.30 0.46 4.0

0.25 23.5 0.554 4.54 0.42 3.4

0.30 19.3 0.616 4.80 0.38 3.0

0.35 16.0 0.678 5.05 0.35 2.6

0.40 13.2 0.740 5.32 0.31 2.3

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14.10 Resource Estimation Uncertainty

Table 14.13 summarizes the sensitivity of the resource estimates to modeling

method by comparing the gold resources estimated by OK, NNP and ID8 at 0.20, 0.30

and 0.40 g/t gold cut-offs (measured - indicated blocks only). Relative to the ID8 method

used to estimate current resources OK and NNP show constant differences of 10-20

percent in tonnage and grade, but contained gold ounces compare within 1 percent at

the 0.20 g/t cut-off, 5 percent at the 0.30 g/t cut-off and 9 percent at the 0.40 g/t cut-off.

With the mining cut-off expected to be in the 0.30 g/t range, and with project economics

more dependent on contained gold than on tonnage and grade distribution, modeling

method is not considered to be a major source of uncertainty.

Table 14.13

Sensitivity of Resource to Estimation Approach

0.2 g/t Gold Cut-off 0.3 g/t Gold Cut-off 0.4 g/t Gold Cut-off

Mt Gold g/t M oz. Mt Gold g/t M oz. Mt Gold g/t M oz.

OK 164.5 0.538 2.85 111.1 0.677 2.42 75.4 0.834 2.02

NNP 118.6 0.749 2.86 83.9 0.956 2.58 61.8 1.176 2.34

ID8 142.2 0.618 2.83 96.2 0.795 2.46 68.3 0.979 2.15

OK/ID8, % +16 -13 +1 +15 -15 -2 +10 -15 -6

NNP/ID8% -17 +21 +1 -14 +20 +5 -10 +20 +9

Table 14.14 summarizes the sensitivity of the resource estimates to the vertical

search distance, which is often difficult to quantify in deposits where mineralization is

not cut off at depth by drilling. The results are obtained from NNP runs that use a

constant 50 meter horizontal search but which vary the vertical search between 10

meter (most pessimistic) and 100 meter (most optimistic). Tonnages and contained gold

ounces are analogued respectively by the number of blocks above cut-off (“Blocks”) and

the number of blocks times the gold grade (“Product”).

Table 14.14

Sensitivity of Resource to Vertical Search Distance

Vertical Search

0.2 g/t Gold Cut-off 0.4 g/t Gold Cut-off

Blocks Gold g/t Product Blocks Gold g/t Product

50m 76,218 0.695 52,972 38,522 1.097 42,259

100m 91,075 0.701 63,844 45,961 1.111 51,063

% Relative to 50m +20 +1 +21 +19 +1 +21

10m 66,601 0.684 45,555 33,694 1.074 36,187

% Relative to 50m -13 -2 -14 -13 -2 -14

Varying the vertical search between 10 m and 100 m has only a minor impact on

grade but a 13 to 21 percent impact on tonnes and contained ounces relative to the 50

meter “base case”. Increasing the vertical search from 10 to 100 meter raises contained

gold by 40 percent. Vertical search distance is clearly a significant source of uncertainty

while the deposits remain open to depth.

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15.0 MINERAL RESERVE ESTIMATES

The mineral reserve is the total of all proven and probable category ore that is

planned for production. Section 16 details the mine plan and schedule that have been

determined to be the most economic method of extracting this reserve. The mineral

reserve was established by tabulating the undiluted tonnes and grades of proven and

probable material within the designed final pit geometry that is scheduled as ore to the

crusher over the mine life. A floating cone algorithm (independently verified by Whittle

optimizations) was used to determine the final pit design and internal phase designs.

15.1 Floating Cones

The floating cone optimization algorithm is a commonly used and accepted

industry tool for providing guidance to mine design. The algorithm applies an estimate of

costs and recoveries along with overall pit slope angles to establish theoretical

economic breakeven pit wall locations.

Economic input applied to the cone algorithm is based on the Preliminary

Economic Assessment (PEA) and subsequent estimates as it was one of the first steps

in the development of the mine plan. However, the cone geometries should be

considered as a guide as they do not necessarily account for minimum safe mining

widths or access to sequential phases. The important result of the cones is the

quantification of the relative changes in geometry between the cones as a function of

increasing metal prices and or costs. Lower metal prices result in smaller pits which

provide guidance to the design of the initial and internal phase designs as these are

usually indicative of high value areas of the deposit. The change in cone geometry as

metal prices are increased indicates the best directions for the succeeding phase

expansions to the ultimate pit.

A suite of floating cones was generated using gold prices between US$ 1200/oz.

and US$ 400/oz. with two goals in mind: Firstly, to determine the extents of the ultimate

pit as well as the contained gold within; and secondly to provide guidance for the

optimum interim cutbacks for the initial years of mining.

The parameters in Table 15.1 were used as inputs when generating the floating

cones. The metal recoveries and costs used for the pit definition are preliminary and

different than the metal recoveries and costs generated by the Feasibility Study

because pit definition is one of the initial steps of developing a mine plan. The mining

costs resulting from the Feasibility Study given in Section 21.2 and final recoveries

presented in Table 13.30 are the inputs that were applied to the financial model.

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Gold Price

Silver Price

Mining Cost

Processing Cost

Recovery

Refining Charges

Inc. Haulage Cost

Discounting

Overall Slopes

Table 15.1

Floating Cone Inputs

$400 to $1,200/oz (in $100/oz increments)

1/60*Gold Price ($20/oz when gold is $1,200/oz)

Waste

Ore

Artavasdes/Tigranes

Erato

$2.98/t

$1.98/t

$3.18/t

$3.38/t

Deposit Gold Silver

Artavasdes 84.86 39.88

Tigranes 89.35 23.27

$0.15/gm gold

Erato 93.72 58.51

$0.02/t/bench (below 2800 m elevation)

0.5%/bench

Volcanics

Andesite

Colluvium

Grades less than 0.15 gm Au/t have an applied recovery of 0%

Processing costs were provided by KDE and are based on a 10 Mt/year

throughput rate. Costs for processing ore from Erato were increased by an additional

US$ 0.20/t to account for the longer distance for ore to be hauled to the crusher from

the pit.

A mining cost of US$ 1.98/t for waste, was derived from the PEA when the waste

rock dump was sited on the eastern edge of the Amulsar ridge in relatively close

proximity to the pits. As the waste dump has been relocated due to geotechnical

constraints to a location approximately 4.5 km north of Tigranes/Atavasdes an

additional US$ 1.00/t was added to the waste mining cost.

The inter-ramp slope angles are by lithology:

Volcanics = 45 o for slopes with dip azimuths ranging from 90-360°

42 o for slopes with dip azimuths ranging from 0-90°,

Andesite = 30 o and

Colluvium = 29 o

For the floating cone runs, the interramp slopes were reduced by approximately

3 o to account for haul roads in the pit walls. Slope angles used were recommended by

Golder in their June 2012 Pit Slope Design Report (Golder, 2012c).

42˚

27˚

29˚

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Figure 15.1 illustrates the US$ 900/oz cone that was used as guidance for the

ultimate pit boundary. Figure 15.2 depicts the cones between US$ 400/oz and

US$ 1200/oz sliced at an elevation of 2830. This figure can be compared with a slice of

the phases in the following section at the same elevation. Figure 15.3 shows cross

sections of the US$ 400, 600 and 900/oz cones whose section lines are given in

Figure 15.2.

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Figure 15.1

$900/oz Floating Cone used for Ultimate Pit Design

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Figure 15.2

US$ 400-US$ 1200/oz Floating Cones Sliced at 2830m Elevation

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A A’

B B’

C C’

15.2 Final Pit Design

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Figure 15.3

Cross Sections of US$ 400, US$ 600, and US$ 900/.oz Au Cones

The final pit design is based on the shell generated by the US$ 900/oz cone as a

result of the evaluation of the discounted net value at US$ 1200/oz gold and US$ 20/oz


Lydian International - Amulsar Resource Update and Heap Leach Feasibility Study Page 148

silver prices for all of the cone geometries. Cones were evaluated at discount rates of

0%, 5% and 10%, using the US$ 1200/oz gold and US$ 20/oz silver metal prices and

the same cost estimates that were used in generating the floating cones. Table 15.2

shows the results of the cone evaluations.

The cones above US$ 900/oz. showed no increase in contained value for the

additional material mined. This is also a function of the estimation being data limited as

the cone at US$ 900/oz captures ore up to where drilling is limited and insufficient drill

data exists to classify material as either measured or indicated. Essentially at $900/oz

gold price the ore body is robust enough that all material in the current block mode is

extracted meaning that extensions to the ore body at depth have a high likelihood of

being economic in the future.

Table 15.3 shows the resulting tonnages contained within the cone shapes.

Material above a cut-off of recoverable 0.25 g/t gold is reported as economical material.

The results of the cone evaluation are presented graphically in Figure15.4.

Gold Price

for Cone Run

$/oz

Table 15.2

NPV of Floating Cone Geometries Evaluated at $1,200/oz Au and $20/oz Ag

NPV @ $1,200/oz

no discount

$1,000's

NPV @ $1,200/oz

5% discount

$1,000's

NPV @ $1,200/oz

10% discount

$1,000's

400 906,500 851,500 800,500

500 1,220,000 1,139,000 1,063,000

600 1,376,000 1,280,000 1,192,000

700 1,626,000 1,500,000 1,385,000

800 1,749,000 1,601,000 1,468,000

900 1,761,000 1,611,000 1,477,000

1000 1,760,000 1,610,000 1,475,000

1100 1,759,000 1,608,000 1,472,000

1200 1,755,000 1,604,000 1,468,000

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Gold

Price for

Cone Run

Economic

Material Rec.

Au>0.25 g/t

Table 15.3

Material Contained within Floating Cone Geometries

Contained Metal

Recoverable

Metal

Recoverable Ounces Waste Total

$/oz kt Au g/t Ag g/t Au g/t Ag g/t Au Ag kt kt

400 33,775 0.959 4.23 0.832 1.54 903,461 1,672,272 37,312 71,087

500 50,997 0.884 4.02 0.771 1.55 1,264,124 2,541,365 66,682 117,679

600 62,691 0.831 3.84 0.725 1.45 1,461,282 2,922,563 90,268 152,959

700 82,141 0.790 3.52 0.694 1.37 1,832,779 3,618,023 151,247 233,388

800 90,653 0.796 3.44 0.704 1.37 2,051,850 3,992,946 193,760 284,413

900 92,983 0.790 3.42 0.698 1.36 2,086,651 4,065,680 204,229 297,212

1,000 94,652 0.785 3.40 0.694 1.35 2,111,932 4,108,226 215,274 309,926

1,100 96,618 0.780 3.40 0.690 1.35 2,143,374 4,193,557 227,968 324,586

1,200 97,708 0.778 3.40 0.687 1.35 2,158,130 4,240,867 235,966 333,674

Figure 15.4

Results of Floating Cone Evaluations

A drawing of the final pit is presented in Figure 15.5 at the same scale for

comparison against the $900/oz floating cone in Figure 15.1. This pit is the end result of

mining 7 internal phases that are described in more detail in Section 16.

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Figure 15.5

Ultimate Pit

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15.3 Mineral Reserve Estimate

The mineral reserve for the project is the proven and probable material that is

sent to the crusher over the life of the mine. Due to the location of the Amulsar deposit

on the top of a ridge the construction of a sizeable low grade stockpile near the crusher

is difficult. A stockpile of approximately 655,000 tonnes is generated in the second year

of mining due to the grade of material being economic but not sufficient to displace

higher grade ore which averages above the 0.35g/t recovered cut-off grade. Since no

other low grade stockpile is generated during the mine life, the reserves for the project

are a total of the undiluted ore sent to crusher during mining and the stockpile

generated in Year 2.

The proven and probable mineral reserves for the project are presented in

Table 15.4.

Category

Table 15.4

Mineral Reserves Represent the Undiluted Ore Scheduled to the Crusher

Ore

kt

Contained Recoverable Contained Recoverable

Gold

g/t

Silver

g/t

Gold

g/t

Silver

g/t

Gold

oz

Silver

oz

Gold

oz

Silver

oz

Proven 51,143 0.801 3.37 0.713 1.31 1,317,000 5,541,000 1,172,000 2,154,000

Probable 37,106 0.789 3.43 0.694 1.17 941,000 4,092,000 828,000 1,393,000

Proven plus

Probable

88,249 0.796 3.40 0.705 1.25 2,258,000 9,633,000 2,000,000 3,547,000

*Material in Year 2 above 0.30 g/t recoverable gold stockpiled

1. The gold and silver recoveries vary by deposit area based on the metallurgical testwork given in section 13.7.2. A recoverable

grade for gold and silver is assigned in the block model and used for tabulations

The mineral reserve tonnes and contained ounces stated in Table 15.4 are less

than the tonnes and contained ounces shown in the mine schedule in Table 16.4. The

mine production schedule shown in Section 16 includes dilution estimates which are not

included in the mineral reserves statement. The difference is 6,645 kt of material with a

contained gold grade of 0.15 g/t and contained silver grade of 1.5 g/t. These dilution

grades are supported by the average grade of metal in the model blocks enveloping the

scheduled ore. More discussion on dilution is given in Section 16.

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16.0 MINING METHODS

Mining of the Amulsar deposit is planned to be accomplished by conventional

open pit, truck and shovel mining methods. As part of the mine plan, consecutive mine

phases were designed in accordance with the outputs from the sequential floating

cones. A schedule for the mining of the phases has been developed that moves higher

gold production forward in the mine life to reduce payback periods whilst maintaining

material movements that effectively utilize the selected equipment.

The schedule delivers ore to the crusher at a rate of 5 million tonnes per annum

in the first three years of mine life, increasing to 10 million tonnes per annum following a

crusher capacity increase in the later part of Year 3. After crushing, the ore will be

delivered via conveyor to the heap leach pad for cyanide leaching.

The steps for the development of the mine plan were as follows:

1) Floating cone guidance to phase design

2) Phase designs

3) Mine production schedule (strategy to maximize project return on

investment).

4) Waste material allocation.

5) External haul road design

6) Time sequence mine plan drawings

7) Equipment and Manpower requirements

16.1 Pit and Phase Design

Phases were designed using the floating cones outlined in the previous section

as guidance. The two initial phases were designed based on the US$ 400/oz cones with

the ultimate pit extents guided by the US$ 900/oz cone. No significant value was gained

by increasing the size of the ultimate pit beyond the US$ 900/oz cone, although it is

expected that the ultimate pit will increase as a function of ongoing drilling onsite, which

will upgrade potential resource into indicated and inferred categories.

The following criteria were followed when designing the mining phases:

Table 16.1

Phase Design Criteria

Bench Height:

10 m

Interramp Slopes recommended by Golder

Volcanics: 42-45

Andesite: 30

o double benched

single bench

Colluvium: 29

o single bench

Road Width: 25 m

Road Gradient: 8 % (maximum of 10%)

The majority of haul roads were designed to a lower than industry standard

gradient of 8 percent, to make hauling conditions safer in icy winter conditions.

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In some instances, short segments of the haul roads were increased to 10

percent gradients to achieve desired pit geometries. Sequential phases were designed

with at least 100 meters of bench width between push backs to allow sufficient

operating room for mining equipment.

A total of seven phases are scheduled to be mined to arrive at the current design

ultimate pit limit. The Artavasdes and Tigranes areas are mined out with five phases

and the Erato area is mined in two phases. When sequencing the phases, preference

was given to phases having the lowest cost per ounce of gold produced so as to

maximize cashflow in the early years of the project. Except for the last year of mining,

more than one phase is active at any given period of time to provide adequate ore

exposure while stripping areas for future ore release.

Table 16.2 is a comparison of the designed ultimate pit tonnage with the tonnage

contained in the US$ 900 floating cone; ore tonnages are undiluted. The material

difference between the floating cone and the final pit design is less than 5 percent which

is in line with industry standard and is a function of ramp design and operational

constraints which are difficult to quantify via floating cone or Whittle.

Figure 16.1 shows the pit phases sliced at 2830 m elevation for comparison with

the cones sliced at the same elevation in Figure 15.2. Figure 16.2 show the phases in

cross section to be compared with the cross sections of the floating cones in Section 15;

the section lines are given in Figure 16.1.

Table 16.2

Comparison of Designed Phase Tonnes Against $900 Cone Tonnes at a 0.25 g/t Recovered Gold Cut-off

Volume

Boundary

Rec. Au

cut-off g/t

Mat. > cut-off

ktonnes

Rec. Au

g/t

Rec. Ag

g/t

Rec.

Au oz.

Rec.

Ag oz.

Waste

ktonnes

Total Mat.

ktonnes

Strip. Rat.

w/o

$900 Cone 0.25 92,983 0.698 1.36 2,087 4,066 204,229 297,212 2.20

Ultimate Pit 0.25 89,710 0.697 1.36 2,011 3,912 216,896 306,606 2.42

% Difference -3.65 -0.10 -0.27 -3.75 -3.92 5.84 3.06 9.15

The individual phase tonnages are shown in Table 16.3 at a recovered gold cutoff

grade of 0.25 g/t on an undiluted basis.

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Phase

Table 16.3

Phase Tonnages and Grades at a 0.25 g/t Recoverable Gold Cut-off

Mat > cut

ktonnes

Rec Au

g/t

Rec Ag

g/t

Recoverable Oz.

Waste

kt

Total

kt

Ph1a 4,697 0.849 0.48 128,211 72,487 5,337 10,034 1.14

Ph1 7,083 0.953 0.81 217,024 184,459 10,169 17,252 1.44

Ph2 21,733 0.713 1.97 498,204 1,376,523 44,242 65,975 2.04

Ph3 13,869 0.696 1.56 310,350 695,613 28,382 42,251 2.05

Ph4 18,534 0.592 0.84 352,768 500,549 46,194 64,728 2.49

Art/Tig_Tot 65,916 0.711 1.34 1,506,557 2,829,631 134,324 200,240 2.04

Erato ph1 9,186 0.564 1.63 166,572 481,406 16,713 25,899 1.82

Erato 14,608 0.72 1.28 338,159 601,172 65,859 80,467 4.51

Erato Tot 23,794 0.660 1.42 504,731 1,082,578 82,572 106,366 3.47

Total 89,710 0.697 1.36 2,011,288 3,912,209 216,896 306,606 2.42

Au

Ag

SR

W/O

As the three separate deposits of Artavasdes, Tigranes and Erato have different

gold recoveries, a recovered gold variable was inserted in the resource block model

(based on the deposit wireframes) on a block by block basis to facilitate more accurate

mining and economic modeling. For mine scheduling and reporting purposes, the

recovered gold grade has been used instead of the contained gold grade as this allows

consideration for recovery in planning and prevents lower value ore from having priority

over higher value ore.

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Figure 16.1

Phases Sliced at 2830 m elevation

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A A’

Phase 6

Phase 7

B B’

Phase 1

Phase 4

C C’

Phase 3

Phase 2

Figure 16.2

Cross Sections of Designed Phases Showing Gold Grade in Block Model

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16.2 Mine Schedule

The seven phase designs were scheduled to deliver 5 million tonnes of ore per

annum in the first three years of mine life and 10 million tonnes per annum for the

remainder of the mine life starting in Year 4 upon completion of the expanded crusher

facilities. Waste stripping was scheduled far enough in advance that at any given time,

there is sufficient ore exposed to provide a continuous feed to the crusher from the pit

without stockpiling. At the projected crushed ore rates, the operation has a 12 year mine

life including mining from the pit and the re-handling of the stockpile accumulated in the

second year of mining.

Several criteria were considered when generating the mine schedule:





Maximizing the project NPV by varying the cut-off grades by period to move

the highest value ounces forward in the mine life

Targeting consistent production of over 200,000 ounces of recoverable gold a

year from the heap for the first 3 years following the upgrade to the crushing

facility in Year 3.

Matching and keeping consistent the material movement rates to correspond

with realistic loading units outputs to ensure maximum usage of mine capital

Minimizing ore stockpiling because of the lack of accessible locations for

stockpiles in the vicinity of the crusher and to reduce costs associated with

rehandling.

At US$ 1200/oz. the marginal cut-off grade per tonne of ore is approximately

0.09 grams per tonne.

Marginal Cut-off Grade = (ProcessCost+G&A Cost) = $3.38/t

Gold Price ($/g)

$38.58/g

= 0.087g/t

As the modeling of the ore body shows large continuous volumes of economic

grades, no ore loss has been applied to the schedule as mining is scheduled at a much

higher cut-off than the true internal cut-off grade. Consequently scheduled ore is rarely

bounded by truly uneconomic material and as such ore loss due to strict dilution control

measures is unlikely.

Dilution of the higher grade ore will occur and has been modeled in the schedule

by including an additional 7 percent at a grade of 0.15 g/t. This material is included to

account for some mixing of higher grade material with lower grade but still economic

material at the interfaces of the ore boundaries. The 0.15 g/t dilution grade is below the

lowest cut-off grade for any of the given years which ranges between 0.20 g/t and 0.35

g/t recovered gold.

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To account for this dilution, 93 percent of the desired ore tonnage was scheduled

for any given time period and the additional 7 percent was assumed to be dilution

incurred in the mining process at the grade of 0.15 g/t. For example, in Year 4, 9,300

ktonnes at a head grade of 0.73 g/t contained gold are scheduled to the crusher. It was

modeled that an additional 700 ktonnes of dilution will also be sent to the crusher in

Year 4 with an assumed grade of 0.150 g/t contained gold. This is a realistic estimate

for dilution as 5,659 ktonnes of material greater than 0.15 g/t but less than the cut-off

grade of 0.280 g/t are sent to the waste dump in this period. The average head grade of

this economic but “sub-grade” material is 0.246 g/t recoverable gold. Life of mine, the

average block grade of blocks bounding scheduled ore blocks is 0.22 g/t recoverable

gold.

The resulting mining and crusher feed schedule with material movements is

provided in Table 16.4. Cut-off grade for material sent to the crusher is always

maximized well above the true breakeven cut-off grade in an effort to increase gold

produced for a given throughput in time. In Year 2 of mining, a small low grade stockpile

is generated when the crusher cut-off grade is 0.35 g/t recoverable gold. This is to keep

a consistent mining rate as well as maximize the grade of material fed to the crusher in

early years.

The cut-off grade for the low grade stockpile in this period is 0.30 g/t recoverable

gold. This material is planned for re-handle to the crusher in Years 10 and 12. In total,

385,000 tonnes of ore is stockpiled in pre-production and re-handled to the crusher in

the first quarter of crusher operation. A graphical representation of the schedule is given

in Figure 16.3 showing ore tonnes sent to crusher, waste tonnes mined, and

recoverable gold grade sent to the heap leach pad.

The drop in ounce production in Years 7 - 10 is a result of the commencement of

mining in the early stages at Erato. As the drill density at Erato is less, a larger

proportion of the material inside the ultimate pit shell is in the inferred category and

hence cannot be included in this study. It is expected that as exploration activities

continue in 2012 and 2013, more material will be upgraded from inferred and this drop

in ounces produced can be reduced. Increased understanding of the Erato orebody will

also lead to more optimized stage designs which will also improve the production

schedule in the later years.

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Table16.4

Material Movements Total Annual Summary

Parameter yr1 yr2 yr3 yr4 yr5 yr6 yr7 yr8 yr9 yr10 yr11 yr12 Total

Cut-off Grade rec. Au 0.28 0.35 0.30 0.28 0.26 0.26 0.26 0.24 0.22 0.20 0.26 0.25

Total Mined (bcm) 4,317,300 6,452,700 6,411,300 14,113,500 14,269,100 14,371,100 14,348,400 14,379,800 14,325,400 14,299,800 12,824,800 549,500 130,662,700

Total Mined (t) 10,064,000 15,127,000 15,193,000 33,466,000 33,500,000 33,500,000 33,500,000 33,500,000 33,500,000 33,500,000 30,433,000 1,323,000 306,606,000

Waste Mined (bcm) 2,735,997 4,065,867 4,311,687 9,936,188 10,081,896 10,164,218 10,132,056 10,152,273 10,097,873 10,304,795 8,664,585 156,005 90,803,440

Waste Mined (t) 6,314,000 9,471,880 10,193,000 23,466,000 23,500,000 23,500,000 23,500,000 23,500,000 23,500,000 23,959,000 20,433,000 375,000 211,711,880

Ore Mined (bcm) 1,581,303 2,386,833 2,099,613 4,177,312 4,187,204 4,206,882 4,216,344 4,227,527 4,227,527 3,995,005 4,160,215 393,495 39,859,260

Ore Mined (t) 3,750,000 5,655,120 5,000,000 10,000,000 10,000,000 10,000,000 10,000,000 10,000,000 10,000,000 9,541,000 10,000,000 948,000 94,894,120

Stkpl Rehandle (bcm) 162,066 - - - - - - - - 193,141 - 82,559 437,766

Stkpl Rehandle (t) 385,000 - - - - - - - - 459,000 - 196,120 1,040,120

Au

Ore Grade (Au) Insitu (g/t) 0.915 0.966 0.957 0.730 0.812 0.870 0.639 0.658 0.549 0.562 0.815 1.508 0.750

Ounces (Au) Insitu (Oz) 110,375 175,557 153,784 234,711 261,160 279,822 205,529 211,505 176,540 172,454 262,012 45,973 2,289,423

Ore Grade (Au) Rec. (g/t) 0.806 0.849 0.837 0.630 0.701 0.753 0.561 0.578 0.493 0.521 0.764 1.414 0.665

Ounces (Au) Rec (Oz) 97,217 154,365 134,631 202,544 225,501 241,967 180,275 185,957 158,656 159,688 245,558 43,086 2,029,444

Ag

Ore Grade (Ag) Insitu (g/t) 2.12 3.16 5.02 4.85 3.76 3.94 3.43 3.00 2.27 2.14 2.35 3.20 3.27

Ounces (Ag) Insitu (Oz) 255,464 574,921 806,334 1,558,899 1,209,553 1,265,710 1,104,352 963,816 728,417 656,738 754,352 97,612 9,976,166

Ore Grade (Ag) Rec. (g/t) 0.64 0.97 1.56 1.54 1.19 1.25 1.19 1.02 1.00 1.19 1.37 1.87 1.21

Ounces (Ag) Rec (Oz) 77,656 176,065 251,430 495,976 384,109 400,444 383,821 327,271 320,700 364,514 441,372 57,113 3,680,473

Strip Ratio (W:O) 1.68 1.67 2.04 2.35 2.35 2.35 2.35 2.35 2.35 2.51 2.04 0.40 2.23

Material Processed Total

Tonnes Crushed (t) 3,750,000 5,000,000 5,000,000 10,000,000 10,000,000 10,000,000 10,000,000 10,000,000 10,000,000 10,000,000 10,000,000 1,144,120 94,894,120

Au

Feed Grade Insitu (g/t) 0.915 1.046 0.957 0.730 0.812 0.870 0.639 0.658 0.549 0.553 0.815 1.310 0.750

Feed Ounces Insitu (Oz) 110,375 168,122 153,784 234,711 261,160 279,822 205,529 211,505 176,540 177,667 262,012 48,195 2,289,423

Feed Grade Rec. (g/t) 0.806 0.920 0.837 0.630 0.701 0.753 0.561 0.578 0.493 0.511 0.764 1.224 0.665

Feed Ounces Rec. (Oz) 97,217 147,858 134,631 202,544 225,501 241,967 180,275 185,957 158,656 164,249 245,558 45,032 2,029,444

Ag

Feed Grade Insitu (g/t) 2.12 3.27 5.02 4.85 3.76 3.94 3.43 3.00 2.27 2.15 2.35 3.05 3.27

Feed Ounces Insitu (Oz) 255,464 525,125 806,334 1,558,899 1,209,553 1,265,710 1,104,352 963,816 728,417 691,955 754,352 112,192 9,976,166

Feed Grade Rec. (g/t) 0.64 1.00 1.56 1.54 1.19 1.25 1.19 1.02 1.00 1.17 1.37 1.68 1.21

Feed Ounces Rec (Oz) 77,656 160,632 251,430 495,976 384,109 400,444 383,821 327,271 320,700 375,437 441,372 61,622 3,680,473

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16.3 Waste Movement

Figure 16.3

Graphical Presentation of Mine Schedule

At Amulsar, the distance from the pit to the waste dump is approximately 4.5 km.

The waste dump, designed by Golder, has a design capacity of 158 million tonnes with

room for expansion subject to environmental approvals. Over the mine life, 150.5 million

tonnes of waste are sent to the dump. At the end of the mine life, an opportunity to back

fill the pits becomes available. This has three benefits, namely:




Decreasing the haulage requirements which reduces operating costs;

Increasing total waste storage capacity with minimal environmental impact

and;

The prevention of the formation of pit lakes after mining is complete.

In total, 61.1 million tonnes primarily from the latter stages of Erato are scheduled

to be backfilled into the Artavasdes and Tigranes pits. Figure 16.4 shows the final

planned backfill of the mining phases. Table 16.4 outlines the waste movements based

on the schedule.

In Years 4, 5 and 6, waste material is sent to a waste stockpile east of the

Artavasdes pit. This material below the cut-off grade in Years 4 through 6 is stockpiled

rather than sent to the waste dump to prevent an unnecessary spike of trucks during

these years. It is re-handled to the pit backfill in Year 11 when more truck shifts are

available. Runoff from these temporary stockpiles will be collection in ponds adjacent to

the crushing plant.

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Figure 16.4

Pit Backfill at End of Mine Life

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Period

Table 16.5

Waste Movement Required for Mine Schedule

To Dump

ktonnes

To bkfill

ktonnes

To Stkpl

ktonnes

Total Waste

Moved

ktonnes

Yr1 6,314 6,314

Yr2 9,472 9,472

Yr3 10,193 10,193

yr4 19,666 3,800 23,466

yr5 19,700 3,800 23,500

yr6 23,500 23,500

yr7 23,500 23,500

yr8 23,500 23,500

yr9 8,862 14,638 23,500

yr10 5,875 18,084 23,959

yr11 20,433 20,433

yr12 7,975 7,975

Total: 150,582 61,130 7,600 219,312

16.4 Low Grade Stockpiles

As discussed previously, a small low grade stockpile was designed near the

crusher to hold 655,000 tonnes of low grade material between 0.30g/t and 0.35 g/t

mined in the second year of mining. This provides a buffer of approximately 3 weeks of

feed capacity to the crusher should mining in the pit be interrupted for an extended

period. If not rehandled earlier, this lower grade tonnage is scheduled to be sent to the

crusher at the end of mine life.

Figure 16.6 shows the proposed stockpiles at the end of Year 10 in the mine life.

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Figure 16.6

Proposed Stockpiles at the end of Year 10

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16.5 External Haul Roads and Time Sequence Drawings

As all three deposits daylight at the top of hills, initial access to the phases will be

with external haul roads. One main waste dump access road has been designed to

handle all material from the Tigranes / Artavasdes pit as well as the Erato pit later in the

mine life. Temporary haul roads will be used to terrace down the hillside until such time

as the pit no longer daylights, at approximately 2850RL. At this point permanent internal

haul roads and ramps will be constructed to ensure access to the deeper benches

extracted later in the mine life.

The design of haul roads and the planned pit progression can be seen in the

annual drawing Figures 16.7 through 16.19.

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Figure 16.7

End of Pre-Production

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Figure 16.8

End of Year 1

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Figure 16.9

End of Year 2

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Figure 16.10

End of Year 3

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Figure 16.11

End of Year 4

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Figure 16.12

End of Year 5

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Figure 16.13

End of Year 6

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Figure 16.14

End of Year 7

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Figure 16.15

End of Year 8

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Figure 16.16

End of Year 9

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Figure 16.17

End of Year 10

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Figure 16.18

End of Year 11

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Figure 16.19

End of Year 12

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16.6 Mining Equipment Fleet

Mine mobile equipment was selected to meet the production requirements as

outlined in Table 16.3. Although operations are targeted to run 365 days per annum, it

has been assumed that inclement weather will render this impossible for 35 days per

annum. Consequently, mining activities are scheduled for 330 days per annum. Shift

lengths for production personnel will be 11 hours with two shifts in a day. Maintenance

personnel both for fixed plant and mobile equipment will work 12 hour shifts, day shift

and night shift. Refueling and minor maintenance work / servicing will be performed in

the 2 hours between shifts. The bulk of the lost mining days are planned to occur in the

winter months during the first and fourth quarters. A three crew rotation is planned for

the Armenian workforce consisting of 14 days on and 7 days off. Expatriates will be

working a 9 weeks on, 3 weeks off roster.

16.6.1 Drill and Blast

Blast hole drilling will be carried out with a fleet of Sandvik DP1500i drills. These

machines are versatile rigs capable of drilling vertical through to horizontal holes of

diameter 80mm – 140mm. Due to the undulating terrain in the early years of mining

these smaller track mounted rigs were given preference to larger blasthole rigs to

provide flexibility in blast design and allow easier access across the deposit. Although

contractors exist in country to provide drilling services, for the purposes of this study it

has been assumed that Lydian will purchase and operate these machines.

It is planned to drill 127mm holes at 3.68 meter spacing. These drill holes will be

sampled and assayed for ore control. Dry blast holes will be loaded with ANFO and wet

holes with emulsion. A number of suppliers exist in country for the provision of

explosives and Lydian has already completed preliminary discussions regards the

supply and management of ANFO, emulsion and packaged explosives in country.

Capital costs have been allocated to account for the construction of magazines and the

purchase of specialized explosives trucks. However, this is primarily due to the early

stages of negotiations with in country suppliers; it is likely that Lydian will ultimately

outsource the explosives supply and management.

16.6.2 Load and Haul

In Years 1-3 the primary loading units will be one 180t Cat 6018 and one 290t

Cat 6030 hydraulic backhoe excavator. These will be supplemented by an additional,

Cat 6018 and Cat 6030 hydraulic backhoe excavators in Year 4.

The smaller machines were chosen in the earlier years to better facilitate the

mining of the hillside. Backhoe configuration was chosen in preference to face shovel as

it allows for greater selectivity during ore mining as well as providing increased flexibility

during the construction phase and early years of mining. A number of these machines

are already in use elsewhere in Armenia and it is expected that access to trained

operators and mechanics will be simpler due to their prevalence.

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For the expansion to 10Mtpa in Year 4, Cat 6030’s were chosen as they allow for

greater production due to their increased size but offer sufficient similarities in terms of

controls etc that upgrading the skills of existing operators should not be overly arduous.

An additional Caterpillar 992K loader has been included in the loading units for

stockpile re-handling to the crusher and other ancillary work. This machine may be used

as a substitute for the Cat 6018 excavators if required.

Hauling of material from the pit to the crusher, dump and stockpiles will be

accomplished with 90t Caterpillar 777G Haul Trucks. The 777G truck is considered an

optimum match for the Cat 6018 excavator and its versatility and size will offer

considerable advantages in the early years of mining as well as construction. Current

capital and operating cost estimations assume that the Cat 777G is used for the

duration of the mine life. However, subject to exploration drilling and increases in

ultimate pit size and mine life, consideration may be given to utilizing 135t Cat 785

trucks as a match for the Cat 6030 excavators in Year 4.

Haul truck productivity was based on a detailed haul time simulation over

measured haul profiles. Truck performance characteristics were based on Caterpillar

published truck specifications for the Cat 777G model. Haul profiles were measured for

each material type, from each pushback to each destination on a quarterly/yearly basis.

These profiles account for the gradient and design of the haul road so as to accurately

model truck speed and cycle times for each period in the mine schedule.

Equipment productivity for excavators and ancillary equipment was calculated on

a shift basis based on Amulsar rock and operating conditions. Productivity for each

machine was calculated based on shift length, planned and unplanned stoppages,

machine utilization, and operator effectiveness. Calculated productivities were then

benchmarked against comparable machines in similar environments.

The productivity per shift and the tonnage requirements set the number of

operating shifts needed per year to move the material. Availability and utilization were

applied to determine the required number of operating units and overall fleet size which

can be viewed in Table 16.5.

16.6.3 Ancillary Equipment

Caterpillar D10 tractor dozers have been selected as the primary materials

handling option for the waste dump, the stockpiles, road construction and for in pit

operations. To supplement the track dozers, Caterpillar 824 wheel dozers will be used

in loading areas and on haul roads to keep floors clean and free of debris that may

damage tyres. A Caterpillar 16M grader will also be purchased to manage haul roads

and dumps to ensure optimum performance from trucks and reduce maintenance costs.

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Dust suppression will be provided by a Caterpillar 777G water truck that will

source water from catchment areas adjacent to the haul road, waste dump and crushing

facility. Light construction work, trenching and general housekeeping will be handled

using a Caterpillar 336 excavator.

Table 16.6 summarizes the mine mobile equipment fleet for the mine life.

Equipment Type

Table 16.6

Equipment Requirements by Time Period

Time Period

Yr-1 Yr1 Yr2 Yr3 Yr4 Yr5 Yr6 Yr7 Yr8 Yr9 Yr10 Yr11 Yr12

Sandvik DP1500 Drill 0 3 2 3 7 7 7 7 7 7 7 6 1

10.0 cu m Exc 0 1 1 1 2 2 2 2 2 2 2 2 2

17.0 cu m Exc 0 1 1 1 2 2 2 2 2 2 2 2 2

Cat 777 Haul Truck 0 19 19 20 37 37 37 37 34 25 24 24 18

Cat D10 Track Dozers 1 3 3 3 3 3 3 3 3 3 3 2 2

Cat 824G Wheel Dozer 1 1 1 1 1 1 1 1 1 1 1 1 1

Cat 16M Motor Grader 1 1 1 1 1 1 1 1 1 1 1 1 1

Cat 777 Water Truck 1 1 1 1 1 1 1 1 1 1 1 1 1

Cat 992G Wheel Loader 1 1 1 1 1 1 1 1 1 1 1 1 1

Cat 336DL Excavator 1 1 1 1 1 1 1 1 1 1 1 1 1

TOTAL 6 32 31 33 56 56 56 56 53 44 43 41 30

16.7 Personnel

Salaried staff requirements are expected to be 44 persons per year; 16

expatriates and 28 nationals (Table 16.7). Labor requirements for operations and

maintenance increase to approximately 208 persons in the last quarter of Year 1. Labor

requirements remain in the lower 200’s of persons until Year 4 when the labor required

increases to 345 persons. The persons required remains in the mid 300’s until Year 8

when the personnel requirements begin to decrease. An allowance for vacation,

sickness and absenteeism (VSA) is included in the overall labor requirement. Table

16.8 is a summary of the mine operations and maintenance personnel requirements.

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JOB TITLE

Table 16.7

Salaried Staff Labor Requirements

Personnel by Time Period

Yr -1 Yr1 Yr2 Yr3 Yr4 Yr5 Yr6 Yr7 Yr8 Yr9 Yr10 Yr11 Yr12

Mine Manager 1 1 1 1 1 1 1 1 1 1 1 1

MINE OPERATIONS:

Mine Superintendant 1 1 1 1 1 1 1 1 1 1 1 1 1

Mine Leading Hands 2 6 6 6 6 6 6 6 6 6 6 6 6

Blasting Foreman 2 2 2 2 2 2 2 2 2 2 2

Supervisors/trainers 1 3 3 3 3 3 3 3 3 3 3 3 3

Mine Operations Total 4 12 12 12 12 12 12 12 12 12 12 12 10

MINE MAINTENANCE:

Maint. Manager 1 1 1 1 1 1 1 1 1 1 1 1 1

Maint. Superintendent 1 1 1 1 1 1 1 1 1 1 1 1 1

Maint. Lead Hand 1 3 3 3 3 3 3 3 3 3 3 3 3

Maintenance Planner 1 2 2 2 2 2 2 2 2 2 2 2 2

Supervisors/trainers 1 3 3 3 3 3 3 3 3 3 3 3 3

Mine Maintenance Total 5 10 10 10 10 10 10 10 10 10 10 10 10

MINE ENGINEERING:

Technical Services Super. 1 1 1 1 1 1 1 1 1 1 1 1 1

Senior Mine Engineer 1 1 1 1 1 1 1 1 1 1 1 1 1

Mining Engineer 1 2 2 2 2 2 2 2 2 2 2 2 2

Surveyor 1 2 2 2 2 2 2 2 2 2 2 2 2

Surveyor Helper 1 3 3 3 3 3 3 3 3 3 3 3 3

Mine Engineering Total 5 9 9 9 9 9 9 9 9 9 9 9 9

MINE GEOLOGY:

Senior Mine Geologist 1 1 1 1 1 1 1 1 1 1 1

Grade Control Geologist 3 3 3 3 3 3 3 3 3 3 3

Sr Geotechnical Engineer 1 1 1 1 1 1 1 1 1 1

Geotechnical Engineer 1 1 1 1 1 1 1 1 1 1

Sampler 6 6 6 6 6 6 6 6 6 6 6

Mine Geology Total 0 10 12 12 12 12 12 12 12 12 12 12 0

TOTAL PERSONNEL 14 42 44 44 44 44 44 44 44 44 44 44 30

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JOB TITLE

MINE OPERATIONS:

Table 16.8

Mine Hourly Labor Requirements

Personnel by Time Period

Yr-1 Yr1 Yr2 Yr3 Yr4 Yr5 Yr6 Yr7 Yr8 Yr9 Yr10 Yr11 Yr12

Drill Operator 0 9 9 9 18 18 18 18 18 18 18 17 2

Shovel Operator 0 3 3 3 6 6 6 6 6 6 6 5 6

Loader Operator 0 3 3 3 6 6 6 6 6 6 7 6 6

Haul Truck Driver 0 54 56 59 109 110 111 111 101 73 71 71 54

Track Dozer Operator 1 6 6 6 6 6 6 6 6 6 6 5 4

Wheel Dozer Operator 0 2 2 2 3 3 3 3 3 3 3 2 1

Grader Operator 1 3 3 3 3 3 3 3 3 3 3 3 3

Service Crew 4 12 12 12 12 12 12 12 12 12 12 12 9

Blasting Crew 6 6 6 6 6 6 6 6 6 6 6

Floating Operator 3 3 3 3 3 3 3 3 3 3 3 3

Laborer 6 6 6 6 6 6 6 6 6 6 6

Operations Total 6 107 109 112 178 179 180 180 170 142 141 136 88

MINE MAINTENANCE:

Mechanic 2 41 42 43 74 74 74 74 69 56 55 52 29

Welder 1 20 21 21 36 36 36 36 33 27 27 26 14

Electronics Tech. 1 6 7 7 11 11 11 11 10 9 8 8 5

Fuel & Lube Man 2 6 6 6 6 6 6 6 6 6 6 6 6

Tire Man 2 6 6 6 6 6 6 6 6 6 6 6 6

Laborer 1 3 3 3 3 3 3 3 3 3 3 3 3

Maintenance Total 9 82 85 86 136 136 136 136 127 107 105 101 63

VS&A at 10% 2 19 19 20 31 32 32 32 30 25 25 24 15

TOTAL LABOR 17 208 213 218 345 347 348 348 327 274 271 261 166

Maint./Operations Ratio 1.50 0.77 0.78 0.77 0.76 0.76 0.76 0.76 0.75 0.75 0.74 0.74 0.72

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17.0 RECOVERY METHODS

Development of the Amulsar Project will be conducted in two phases:



Phase I is the construction of a facility to process ore at a rate of 5 Mtpa.

In the third year of operation Phase II will be constructed to increase

throughput to 10 Mtpa in year four. The Phase II expansion will essentially

entail installation of a duplicate Phase I facility, though some of the unit

operations and ore handling equipment will be initially installed to support

the 10 Mtpa processing rate.

Attached in Appendix 1 are Design Criteria for Phase I and Phase II.

Attached in Appendix 2 are Equipment Lists for Phase I and Phase II.

Attached in Appendix 4 are the following Flowsheets for Phase I and Phase II as

well as General Arrangement, Civil, Electrical and P&ID drawings.

- 10-F-01 Flowsheet and Mass Balance - Primary Crushing

- 13-F-01 Flowsheet and Mass Balance - Secondary Crushing

- 15-F-01 Flowsheet and Mass Balance - Tertiary Crushing

- 17-F-01 Flowsheet and Mass Balance - Lime Addition

- 19-F-01 Flowsheet and Mass Balance - Ore Stacking

- 20-F-01 Flowsheet and Mass Balance - Heap Leach

- 23-F-01 Flowsheet and Mass Balance - Carbon Adsorption

- 25-F-01 Flowsheet and Mass Balance - Stripping and Refining

- 27-F-01 Flowsheet and Mass Balance - Carbon Reactivation

- 30-F-01 Flowsheet and Mass Balance - Utilities and Reagents

The Overall Flowsheet is shown below in Figure 17.1.

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Figure 17.1 Amulsar Overall Flowsheet

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17.1 Crushing Facility

Ore is processed through three stages of crushing to a target crush size of 100

percent minus 12 millimeters.

17.1.1 Primary Crushing

Run-of-mine ore is delivered to the primary crusher feed hopper, or adjacent

stockpile, by rear-dump haul trucks. A static grizzly screen above the hopper limits the

top size of rock fed to the crusher to 700 mm. Below the hopper, an apron feed

transfers ore at a controlled rate to the vibrating grizzly screen. Grizzly screen

oversize, plus 100 mm material, feeds the primary jaw crusher. Grizzly screen

undersize joins the crusher product on the primary crusher discharge conveyor which

feeds the primary crusher transfer conveyor taking the ore to the second stage of

crushing. The primary crushing circuit reduces the size of run-of-mine from a

maximum of 700 mm to approximately 80 percent passing 165 mm. The rock breaker

is installed to serve the static grizzly and the monorail crane and air compressor

support jaw crusher operation. Dust is controlled at the feed pocket by water sprays

and at the screens and transfer points by dust collection/filtration in the bag house.

Tramp iron is removed from the crushed product by way of the magnet mounted above

the discharge of the discharge conveyor.

In the Phase II expansion, the entire primary crushing circuit is duplicated

except both phases share a common run-of-mine stockpile, dust bag house, air

compressor and transfer conveyor.

17.1.2 Secondary Crushing

Primary crushed product is fed into the coarse ore storage bin. Two apron

feeders transfer the ore to the coarse ore transfer conveyor which feed ore at a

controlled rate to the secondary vibrating screen deck. The screen deck oversize, plus

100 mm and plus 28 mm, is fed to the secondary cone crusher. Screen deck

undersize joins the secondary crusher product on a transfer conveyor for delivery to

the third stage of crushing. Secondary crushing reduces the primary crushed product

to approximately 80 percent passing 32 mm. The crane and air compressor are

installed to support crushing operations and dust is controlled at the screen deck and

crusher by collection/filtration.

The Phase II expansion shares the coarse ore storage bin, crane, air

compressor and product transfer conveyor with Phase I, but requires installation of two

additional apron feeders, one vibrating screen deck and one secondary cone crusher.

17.1.3 Tertiary Crushing

Secondary crushed product is discharged onto the fine ore screen tripper

conveyor and delivered to the fine ore screen feed bin. The belt feeder delivers ore

from the bin to the double deck vibrating screen. Screen oversize, plus 30 mm and

plus 17 mm, reports to the screen oversize tripper conveyor and discharged into the

tertiary crusher feed bin.

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Two belt feeders deliver the screen oversize material to two tertiary short cone

crushers. The tertiary crushed product is discharged onto the fine ore screen tripper

conveyor and re-circulates back to the vibrating screen. The screen undersize,

approximately 80 percent passing 12 mm, reports to the fine ore collection conveyor

which discharges onto the fine ore transfer conveyor. The fine ore transfer conveyor

delivers ore to the crushed ore tripper conveyor and into the crushed ore surge bin.

Four belt feeders transfer crushed ore from the surge bin to the overland conveyor.

Tertiary crushing is supported by the air compressor and crane hoist. Dust is

controlled at all transfer points, the screen and crushers by collection/filtration.

The Phase II expansion requires installation of two second tertiary vibrating

screens with belt feeders and two additional tertiary cone crushers with belt feeders.

17.1.4 Stacking

The current overland conveyor system consists of three connecting overland

conveyors, followed by a series of twenty four portable conveyors, ending with a

radial stacking conveyor. The first conveyor is approximately 4 kilometers in length

and spans from the crushing plant to the northwest corner of the heap leach pad.

The second conveyor is approximately 1.2 kilometers in length and continues south

along the west side of the heap leach pad. The third conveyor is approximately 1.2

kilometers and includes a tripper conveyor. The tripper provides stacking capability

over the area of the pad without a need to increase the number of portable

conveyors.

The overland conveyor system was designed by a third party vendor,

Paakkola Conveyors OY. The 4 kilometer main conveyor has a proposed straight

line routing down the mountain from the crushing plant to the heap leach pad.

Paakkola proposed this routing due to the complexity of placing bends in the

conveyor, which would require 900 times the belt, resulting in a minimum of

approximately 1 kilometer in length to place a curve in the conveyor. Utilizing

multiple shorter conveyors would require more maintenance and conveyor

components (e.g. drives, drive ends, tail ends, etc.). The trade-off of using a curved

conveyor or multiple conveyors versus the suggested straight routing resulted in no

advantages in cost savings or required earthworks. The succeeding overland

conveyors also have proposed straight-line routing design based on the same

criteria. Ore is discharged from the stacking conveyor onto the heap leach pad in 8

meter high lifts.

Pebble lime is added on the overland conveyor from a storage silo via screw

feeders with the rate of lime addition varying with tonnage.

17.2 Heap Leach Facility

Golder completed and submitted to Lydian a separate document detailing a

feasibility-level design and cost estimate for the heap leach facility including the leach

pad and collection ponds (Golder, 2012c). Prior to selecting the final location (known

as Site 6), a thorough review of Heap Leach Facility Site Alternatives Analysis (Golder,

2012j) was undertaken jointly by Golder, WAI and Geoteam.

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Heap leaching consists of stacking the crushed ore on the leach pad in 8m lifts

and leaching each individual lift to extract the gold and silver. Barren leach solution

(BLS) containing approximately 0.5 g/l sodium cyanide (approximately 250 ppm free

cyanide) is applied to the ore heap surface using drippers at an application rate of 10

l/h/m 2 . The overall leaching cycle for the ore is at least 140 days total with 30 days of

primary leaching, 80 days of secondary leaching and 30 days of leaching as a buried

lift. This is equivalent to a solution-to-ore application ratio of 3 cubic meters per tonne

of ore. Leaching commences as the BLS piping is installed on the surface of the first

heap lift with a sufficient area to accommodate the applied solution flow rate.

The solution percolates through the ore to the impermeable pad liner where it

collects in a network of perforated solution collection drain pipes installed within a 0.6

meter thick granular cover drain fill layer above the liner. The leaching process is

carried out as a two-stage counter-current leach in order to maximize the gold tenor to

the gold recovery process. Leach solution of intermediate strength is used as recycle

leach solution (ILS) to leach freshly stacked ore. This produces a higher gold grade

pregnant leach solution (PLS) reporting to the pregnant pond.

17.2.1 Leach Pad

The lined leach pad will be constructed in three phases to provide an ultimate

ore heap of 95 Mt stacked in three stages. Each pad phase will be divided into two

cells for a total of six cells.

The Phase 1 (Starter) pad will be constructed at the southern end of the gently

sloping plateau at Site 6 with grade fill placed to maintain pad grades between 0.5%

and 3% to accommodate the stability requirements. In addition to stability

considerations, the grading in Phase 1 accommodates the solution drainage

requirements and provides a sufficient surface to stack the first ore lift on the Phase 1

pad to accommodate the active leaching area requirement. The toe fill will extend

within the central valley of the site northward from the southern pad toe limits until it

daylights into the existing ground.

Ore will be stacked on the Phase 1 pad in a maximum of seven 8 m thick,

horizontal lifts to develop a Stage 1 ore heap with a capacity of 18 Mt during the initial

3.3 years of operations. The Phase 1 leach pad will have an area of 479,690 m 2 and

the Stage 1 heap will have a top surface elevation of 2,229 m. The Phase 1 leach pad

may be constructed in sub-phases to further minimize initial capital costs.

The Phase 2 leach pad will consist of a 465,000 m 2 expansion of the pad to the

north, providing for the stacking of the Stage 2 ore heap above the Stage 1 ore heap

and Phase 2 leach pad. The Stage 2 ore heap will consist of five additional horizontal

lifts above the Stage 1 ore heap level and will have a nominal top surface elevation of

2269 m. The Stage 2 ore heap will add capacity for an additional 27 Mt, which is

projected to occur through the end of Year 6 of operations.

The Phase 3 leach pad will consist of a final 461,120 m 2 expansion of the leach

pad to the north, providing for the stacking of the Stage 3 ore heap above Stages 1

and 2 and the Phase 3 leach pad in horizontal lifts for a nominal maximum heap height

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of 72 m above the ultimate leach pad, with the heap top lifts stepped to match the

sloping pad grade. Stacking of the Stage 3 ore heap is projected to continue through

Year 11 of operations to provide an approximate total ore heap capacity on the

Ultimate pad of 95 Mt.

If additional leachable ore is identified beyond the Stage 3 ore heap capacity, a

fourth pad expansion to the north may be constructed. Up to 120 Mt of ore heap may

be stacked on the pad including the Phase 4 expansion.

The pad will have a basal composite liner system consisting of a 2-mm (80-mil)

linear low-density polyethylene (LLDPE) geomembrane underlain by a 0.3-m minimum

thick compacted low-permeability soil liner. The geo-membrane will be smooth in

most areas and will include a double-side textured strip along the downgradient toe of

the pad to enhance heap stability.

The drainpipe network above the leach pad liner will be embedded within the

0.6 m thick liner cover drain fill composed of free-draining, hard, and durable granular

material. Solution and storm runoff flows collected by the drainpipe network in each

pad cell will be routed via transfer pipes through the leach pad cell spillways to the

process ponds, and will be directed by valve control to either the pregnant or

intermediate ponds.

A limited and targeted leak collection and recovery system (LCRS) will be

constructed beneath the leach pad composite liner that will consist of a series of

transmissive drains connected to down-gradient sumps. The LCRS drains will be

underlain by a secondary LLDPE geo-membrane liner. Should a leak ever occur

through the pad liner and be intercepted by the LCRS drain, it would flow through the

drain to the LCRS sump located at the low point of each pad cell, where it would be

removed via a pump. The LCRS will be constructed beneath the pad areas where the

highest potential for elevated hydraulic head and/or concentrated flows occur, e.g., at

the down-gradient cell divider berm locations and beneath the primary solution

collection pipes.

A stock-proof mesh fence with locking gates will be constructed around the

perimeter of the leach pad to prevent wildlife from reaching the pad and ore heap. An

additional purpose of the fence is for public safety and to deter unauthorized access

into the pad area.

17.2.2 Collection Ponds

The collection ponds consist of process (PLS and ILS) ponds and a storm event

(storm) pond sized in accordance with the project design criteria. Additionally, an

overflow pond will be constructed down-gradient of the storm pond. The collection

ponds and overflow pond will be constructed during the Phase 1 leach pad

construction. The collection pond crest elevation will be approximately 15 m lower

than the pad’s lowest point for cut and fill quantity optimization.

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Solution and storm water flows from the pad cells will be routed to the process

ponds. A common divider berm will be constructed between the pregnant and

intermediate ponds for solution and storm water overflow conveyance between these

ponds. A spillway will be constructed between the intermediate pond and the storm

pond for storm water overflow conveyance to the storm pond.

The combined process pond capacity is approximately 94,600 m 3 to bottom of

freeboard depth. The storm pond capacity is approximately 218,040 m 3 to bottom of

freeboard depth.

The process ponds are sized to contain 8 hours of normal operational solution

flow and 24 hours of solution drain-down flow from the ore heap for the Ultimate pad in

case of operational shutdown due to pump failure or power loss. Considering a

maximum solution flow rate of 2,848 m 3 /hr the 8 hours of normal operational storage

and 24 hours of emergency drain-down storage require 22,784 m 3 and 68,352 m 3 ,

respectively, for a combined volume of 91,136 m 3 . The approximate process pond

capacity of 94,600 m 3 exceeds this combined volume, and therefore the ponds provide

for full passive containment below their freeboard for these flows.

The storm pond was sized to accommodate project design criteria of 150% of

the 100-year, 24-hour design storm event runoff from the Ultimate pad and collection

pond areas. Considering a 100-year, 24-hour design storm depth of 95 mm and an

ultimate area of approximately 1,405,800 m 2 across which precipitation would be

collected, 150% of the design storm over this area would generate a maximum runoff

of 200,330 m 3 (assuming no uptake into the heap). The storm pond capacity of

approximately 218,040 m 3 exceeds this design runoff volume, and therefore the storm

pond provides for full containment below its freeboard of the design contingency

containment criteria.

The storage capacity of the process ponds and storm pond were also evaluated

against the expected inflows that would occur during the wettest month on record. A

maximum monthly precipitation of 213.8 mm was observed in the 41 years of

precipitation data. This precipitation would generate a maximum containment volume

of 300,560 m 3 from the ultimate facility. The heap is expected to retain a portion of

this volume through uptake of the ore from the delivered water content of 3% to its

field capacity water content of 10%. Considering an ore stacking rate of 10 Mtpa

(833,300 Mt per month), the ore will uptake approximately 58,300 m 3 of water or

solution during this month. Considering 8 hours of normal operational flow storage

(22,800 m 3 ), 300,600 m 3 of water from precipitation, and 58,300 m 3 of water or

solution uptake into the heap, the net volume in the ponds at the end of the wettest

month on record would be 265,000 m 3 . The approximate combined process ponds

and storm pond capacity of 312,640 m 3 exceeds this net volume, and therefore the

ponds would provide for full containment below their freeboard during this month.

An additional overflow pond will be constructed downgradient of the storm pond

to contain potential overflow discharge from the storm pond, should a low probability

event or series of events ever occur that exceed the project design containment

criteria.

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The process ponds will have a composite double geo-membrane liner system

comprised of top (primary) and bottom (secondary) geo-membranes, with an

intermediate LCRS layer. The bottom geo-membrane will be underlain by a 0.3 m

thick compacted low-permeability soil liner. The bottom geo-membrane will be a 2-mm

(80-mil) thick smooth LLDPE and the top geo-membrane will be a 2-mm (80-mil) thick

single-side textured high-density polyethylene (HDPE) with texturing at top for traction

(as a safety consideration). The LCRS between the two geo-membranes will be a

transmissive geo-composite that is connected to a LCRS sump. Should a leak ever

occur through the top geo-membrane, it would flow through the geo-composite to the

LCRS sump, where it would be removed via a pump. The design intent of the LCRS is

to ensure that no hydraulic head occurs on the bottom geo-membrane, thereby

removing any driving force required for seepage to occur through that geo-membrane.

The storm pond will have a composite liner system consisting of 2-mm (80-mil)

single-side textured HDPE geo-membrane with texturing at top for traction, underlain

by 0.3-m minimum thickness compacted low-permeability soil liner. A 0.3 m thick layer

of cover fill will be placed at the bottom of the storm pond to protect the exposed geomembrane

liner from wind, weather, and ultra-violet damage considering that this pond

will be empty under normal operating conditions.

The overflow pond will be lined with a 0.3 m thick compacted low-permeability

soil liner.

A stock-proof mesh fence with locking gates will be constructed around the

perimeter of the collection ponds to prevent wildlife from reaching the fluids in the

ponds. An additional purpose of the fence is for public safety and to deter

unauthorized access into the collection ponds area.

Top netting will be provided above the process ponds fence to prevent birds

from accessing the fluids in the ponds. If occasional bird access still occurs, additional

deterrent will be employed by using floating plastic balls.

17.3 Process Plant

The process plant consists of an ADR Plant, electrowinning cells, a gold room

and reagent handling equipment.

For the Phase II expansion, essentially a duplicate carbon adsorption train of

five stages and an additional electrowinning cell will be installed; the gold room and

reagent handling facilities will be initially sized to accommodate the increase in metal

production.

The entire process plant designed by Summit Valley Technologies treats seven

tonne batches of pregnant 6 x 16 mesh carbon. The plant processing steps include

carbon adsorption, carbon acid wash, carbon stripping, carbon regeneration, carbon

handling, sodium cyanide and sodium hydroxide mix/storage, electrowinning, and

smelting.

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The sourcing, transportation, handling, use and disposal of any hazardous

substances will be regulated in accordance with relevant management plans prepared

in accordance with international best practice. Framework plans are presented to

support the ESIA submission. More detailed management plans will be developed at

appropriate times before commissioning.

Brief descriptions of each processing step are presented below.

17.3.1 Carbon Adsorption

Pregnant leach solution is pumped into the ADR plant, passes over a trash

screen, and enters the bottom of the first carbon adsorption column. The solution

flows up through the bed of carbon, over the column top and down into the bottom of

the second carbon adsorption column. This is repeated for a total of five carbon

adsorption stages and the design is such that solution flows by gravity through the

columns. Upon exiting the fifth stage of adsorption the solution, now barren, flows

through a carbon safety screen and into the barren solution surge tank. Barren

solution is pumped back to irrigate the heap leach pad.

The carbon flows through the five stages of adsorption counter-current to the

solution. Periodically, once or twice per day, carbon is pumped from the first carbon

adsorption column to the acid wash vessel, or alternatively, the strip vessel. Carbon

from the second carbon adsorption column is pumped into the first column, the third

into the second, and so on. Fresh or regenerated column is added to the fifth carbon

adsorption column. Wire samplers are installed on the pregnant and barren leach

solution lines. The adsorption plant contains a safety shower and a sump with pump

to return solution to the fifth carbon adsorption column.

For the Phase II expansion, a duplicate set of carbon columns is installed with

associated screens, pumps, and samplers though shares the Phase I sump, shower

and barren solution surge tank.

17.3.2 Carbon Acid Wash

Loaded carbon is preferably pumped to the acid wash vessel prior to stripping.

The acid wash vessel, constructed of fiberglass reinforced plastic, holds seven tonnes

of carbon. Hydrochloric acid, diluted to approximately 3 to 5 percent, re-circulates

through the carbon bed for a period of one to two hours. The acid removes inorganic

contaminants from the carbon surface, typically calcium. Caustic solution is pumped

into the vessel to neutralize the acid followed by fresh water. The caustic solution and

wash water report to the neutralization tank which is pumped to the barren solution

tank via the carbon safety screen. The washed carbon is pumped to the desorption

circuit. The acid wash circuit is supported by the safety shower, the sump with pump to

return solution to the neutralization tank and the exhaust fan to vent acid fumes to the

atmosphere.

The Phase II expansion does not require modification to the acid wash circuit.

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17.3.3 Carbon Stripping

Metal is desorbed from the carbon in the strip vessel. The strip vessel holds

seven tonnes of carbon and operates under conditions of elevated temperature and

pressure. Barren strip solution (lean electrolyte) flows up through the bed of carbon

and strips gold from the carbon (rich electrolyte). The rich solution flows through a

carbon bucket trap, a plate and frame heat exchanger to exchange heat with the

barren strip solution, another trim heat exchanger to further cool the solution before

reporting to the electrowinning cell feed tank. Following electrowinning the discharge

solution reports to the barren strip solution tank. Caustic and sodium cyanide are

added to the barren strip solution, which is pumped through the plate and frame heat

exchanger, past an electric immersion heater, and back into the bottom of the strip

vessel. Once or twice per day, the stripped carbon is transferred preferably to the

kiln dewatering screen for thermal regeneration, or alternatively, to the carbon sizing

screen to be returned to the adsorption circuit. The stripping circuit is supported by a

safety shower, wire samplers on the barren and electrowinning feed solutions and a

sump with pump to discharge solution to the adsorption circuit trash screen.

In the Phase II expansion, an additional strip vessel with all auxiliary

equipment is installed except for the sump and safety shower.

17.3.4 Carbon Regeneration

Stripped carbon is pumped to the kiln dewatering screen. Transfer solution and

fine carbon flow to the carbon fines tank. Carbon sized above 16 mesh reports to the

kiln feed bin. By way of the screw feeder, the carbon is passed into the rotating carbon

reactivation kiln. Under a steam atmosphere and at temperatures between 550 and

700 degrees Celsius, organic fouling is removed from the carbon. Carbon exits the kiln

and reports to the carbon quench tank. The reactivated carbon is pumped to the

carbon sizing screen. Transfer water and fine carbon report to the carbon fines tank.

The carbon sized above 16 mesh reports to the activated carbon storage tank and, as

required, is pumped back into the fifth carbon adsorption column.

17.3.5 Carbon Handling

The virgin activated carbon is attritioned prior to being introduced into the

adsorption circuit. The carbon is placed into the carbon attrition tank with process

solution and mechanically agitated for 20 to 30 minutes. This process breaks off any

platelets or sharp corners of the particles, which would have easily broken off while

in the adsorption column. Fines generated in this step can amount to 3 to 5 percent

of the initial carbon weight. The attritioned carbon is pumped to the carbon sizing

screen. Properly sized carbon falls into to the activated carbon storage tank. Fine

carbon and transfer solution report to carbon fines tank. The carbon slurry in the fine

carbon storage tank is pumped to the filter press. The filtrate flows to the barren

solution surge tank. The filter cake is packaged in 50-gallon drums for off-site

shipment and treatment.

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17.3.6 Electrowinning and Smelting

The electrowinning feed solution is pumped from the feed tank into the

electrowinning cell. Cell electrical power is supplied by the rectifier. Metal is

deposited from solution onto stainless steel mesh cathodes. The metal free solution

flows to the electrowinning cell discharge surge tank and from there to the barren

strip solution tank. Periodically, the sludge is washed from the cell cathodes and is

pumped to the plate and frame filter press. The filtrate reports to the barren strip

solution tank. The filter cake is placed into the electric retort. Dry cake is blended

with flux in the flux mixer and then smelted in the induction bullion furnace.

The slag is periodically reprocessed in the furnace though is ultimately

disposed of on the leach pad. The doré is packaged for off-site shipment. The gold

room operations are supported by the exhaust fan over the electrowinning cell, the

dust collector over the furnace, the high pressure water sprayer and the sump with

pump discharging spill/wash solution to the barren solution strip tank.

In the Phase II expansion an additional electrowinning cell and rectifier will be

installed.

17.3.7 Reagent Handling

In addition to the aforementioned lime silo, facilities are provided to handle the

bulk caustic and sodium cyanide. Raw water and sodium hydroxide briquettes, or

flakes, are added to the caustic mix tank to a make-up concentration of 25 percent.

The caustic/mix transfer pump re-circulates the solution and then transfers it to the

sodium cyanide mix tank. Sodium cyanide is added to the mix tank to obtain a 20

percent concentration. This concentrated solution is transferred to the sodium

cyanide storage tank and distributed to the barren solution surge tank and the barren

strip solution tank. This reagent handling station is supported by a safety shower and

the sump pump discharge reporting to the barren solution surge tank.

Metering pumps and lines deliver anti-scalant directly from 50-gallon drums to

the barren solution surge tank.

A pump and solution line delivers concentrated hydrochloric acid from

standard drums or carboys to the dilute acid tank.

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18.0 INFRASTRUCTURE

18.1 Existing Infrastructure and Services

18.1.1 Location

The Amulsar Gold Project covers an area of 130 km 2 , located in south central

Armenia.

18.1.2 Site Access and Roads

The Amulsar area is located 170 km by sealed road from the capital city of

Yerevan, and 15 km by gravel track from the town of Jermuk or 4 km from the town of

Gorayk. The license area straddles the boundary between Vayots-Dzor and Syunik

provinces and incorporates part of the main highway south from Yerevan into Iran.

For the Amulsar project the road will be upgraded from Gorayk to the plant

and mine site to allow for heavy loads. In addition a new bridge will be constructed

over the Vorotan River. The road from Jermuk to the plant and mine site will also be

upgraded however this road is not intended for heavy loads.

18.1.3 Buildings

An employee camp owned by Geoteam will be established at the site. The

camp will have capacity for 200 people in single and shared person accommodation

units and the facilities include a kitchen, laundry, office, workshop, warehouses,

sewage treatment plant, diesel and fuel tanks / mess building and diesel generator.

The remaining employees will live in the nearby towns of Jermuk or Gorayk.

Geoteam has also established an exploration sample preparation and

core/sample storage facility in the village of Gorayk.

It is assumed the contractor will provide a temporary camp to house

approximately 550 people located either near the process plant or in a nearby town.

Senior management will also have a small camp in a nearby town for management,

vendors and equipment vendors.

Additional non-process buildings are discussed in Section 18.2.7.

18.1.4 Resources & Infrastructure

Infrastructure near the project site is very good. The town of Jermuk is 15 km to

the north and the village of Gorayk some 6 km to the south east of the Amulsar

project.

There is good infrastructure surrounding the Amulsar project. This includes the

main sealed highway between Yerevan and Iran, high tension power lines and

substations, a gas pipeline from Iran, year round water from the Vorotan River and a

fibre optic internet cable. As a consequence of the project location on the top of a

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mountain ridge, a reasonable amount of infrastructure will need to be constructed

during project development. In order to ‘fast track’ the project consideration will be

given to constructing portable or skid mounted equipment.

18.1.5 Communications

The exploration camp is currently serviced by satellite dish based internet and

TV connection. Mobile phones work on most parts of the project area and a telephone

connection is available at the exploration camp.

18.1.6 Personnel

As part of the company’s commitment to adding value to the local

communities and building capacity in Armenia, the bulk of the steady state Amulsar

workforce will be Armenian. Ideally, the majority of the workforce will be sourced

from the four local towns, Gorayk, Gndevaz, Jermuk and Saravan. However, given

the lack of extractive industry in these community’s it is expected that a significant

percentage of the highly skilled workforce, i.e. engineers, geologists, metallurgists,

mechanical and electrical tradesmen with mining and processing experience will

need to be recruited from Yerevan and other regional centers in the country.

Positions that cannot be filled locally will be staffed with suitably qualified expatriates

on fixed term contracts, with the ultimate goal of developing qualified Armenia

individuals for these jobs. Initial expatriate numbers are expected to be less than

10% of the workforce with a reduction targeted to less than 5% as local staff gain the

necessary skills to replace them.

All Armenian operations staff will work a 14 days on 7 days off roster.

Personnel recruited locally will continue to be based in their home town, whilst those

recruited from greater Armenia as well as expatriates will be accommodated in

housing provided by the company in close proximity to the mine. Expatriate

personnel will work a 9 weeks on 3 weeks off roster.

The bulk of the workforce, approximately 85%, will be employed in the mining

and processing departments. As on the job training is possible during the

construction period is it expected that come commissioning, the mine operations

roles such as equipment operator, drill and blast assistant, survey assistant and

service crew will be filled almost entirely by local villagers. Training of mobile

maintenance personnel will be supported by the local Caterpillar dealer, Zeppelin

who have specialized training facilities in Russia as well as extensive experience on

other mine sites in Armenia.

Processing personnel for operation and maintenance of the crushers, ADR

plant and conveyors will, in all likelihood be sourced from other mining and heavy

industrial projects in Armenia. As Amulsar is the first gold heap leach project in the

country, external expertise in the form of expatriates will be required to set operating

procedures and train the local workforce in the early years.

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Literacy rates in Armenia are exceptionally high, over 99% for the adult

population. This rate, coupled with Universities in Yerevan offering degrees in

engineering, mining, geology and finance, amongst other things means that there is

a readily accessible pool of graduates with the appropriate skills to fill the technical

and support functions at the mine. Again, in the early years they will be supported by

experienced expatriates to set up operating procedures but in time this requirement

will reduce and it is expected that the bulk of middle management at the mine will be

Armenian.

Salaries have been benchmarked against comparable operations in Armenia

and it is expected that given the working roster Amulsar should be able to attract and

retain people from outside the local region.

The total workforce during operation is estimated at around 550 employees.

The total workforce during construction is estimated at 600.

Construction activities will be split into two categories, earthworks, which

include the heap leach and waste dump construction and fabrication, which will

encompass the crushers, ADR plant, overland conveyor and associated

infrastructure.

Earthworks associated with the heap leach and waste dump will be completed

by local Armenian contractors with assistance from company equipment on the bulk

excavations. Assessment of local contractors in country has been undertaken and a

number of firms exist that have the capability to complete this work.

Specialized fabrication work associated with the heavy infrastructure will

require significantly more expatriate assistance. It is expected that an international

firm will be awarded an EPCM (Engineer, Procure, Construct, Manage) contract to

facilitate the installation of the crushers, conveyor and ADR plant. This firm will

employ the bulk of the expatriates required for the construction of the project. A part

of their mandate will be to maximize the employment of local personnel and to utilise

local sub-contractors where suitable skills exist.

Housing for construction personnel will be at a dedicated camp onsite sized

according to the construction requirements. The workforce overflow during the peak

construction period would be housed in Jermuk, which would be in the region of 400

people by current estimations. Post construction, the construction camp would be

used during operations to accommodate about 200 Armenian staff. It is anticipated

that expatriate and the non-local Armenian management workforce expected to be in

the region of 150 people will be staying at Jermuk. The balance will be leaving in

nearby towns.

The table below summarizes the personnel required during operations by

department.

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Table 18.1

Summary of Operations Personnel

Department

Number of Personnel

Mining 350

Processing 134

Logistics & Finance 27

Environment, Health and Safety 19

Site Administration & Security 22

Total 552

Throughout the construction and operations phase of Amulsar, the company

intends to develop the following through its recruitment and training practices:

• Improvement of local skills to facilitate initiatives that benefit both Amulsar

and the local community;

The development and dissemination of international best practices to the

company and contractor workforces;

• Investment in local businesses to upgrade their ability and increase the

amount of goods and services sourced from local communities around the

mine.

A summary of the senior management on the mine site during steady state

operations can be viewed below.

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MINE MANAGER

Finance Manager

Mine Production

Manager

Process Manager

Engineering

Manager

Logistics &

Support Manager

OH & S Manager

Security Manager

Environmental

Manager

Chief

Accountant

Maintenance

Superintendent

Crushing &

Conveying

Superintendent

Electrical

Superintendent

Camp

Superintendent

Training

Superintendent

Security

Contractor

Community

Liaison

Tech Services

Superintendent

ADR & Heap

Superintendent

Mechanical

Superintendent

Warehouse

Superintendent

Doctor

Environmental

Officer

Production

Superintendent

Metallurgist

IT

Superintendent

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18.1.7 Power Supply

The country has ample electric power from nuclear, hydro and heat electro-power

plant sources. Power lines and sub-station infrastructure are located in close proximity to

the project area. There is also a hydro-electric power plant on the Vorotan River, which is

in the final stages of construction and will have an installed capacity of 1.8 MW. There are

plans to increase the capacity to 2 MW.

Power is not currently reticulated to the Project site although domestic usage

power is available at neighbouring main towns to the south and east. The supply of power

in Armenia is controlled by the Armenian Electrical Networks company (AEN) that owns

the distribution channels of the country in an arrangement whereby in this region power is

purchased from the AEN distribution grid at 35 kV, stepped down at AEN owned

substations and reticulated as required to consumers.

Based on a study by the local Power Network Design Institute, power for the

project will be fed from two sources - 220/110/35 kV “Yeghegnadzor” and 100/35/10 kV

“Sisakan” substations through a 110 kV overhead line. There is a 35kV line option from

“Gorayk” substation, however, it has low reliability due to frequent power outage and

using this line will result in energy losses.

A two-chain 12 km 110 kV overhead power line will be constructed for power

supply to the mine site. This line will connect to “Sisakan” 110 kV overhead line, which in

its turn is connected on one side to 110 kV rods of the “Sisyan” 110/35/10 kV substation,

and on the other side to 110 kV rods of the “Gndevaz”, “Vorotan 3” and “Vorotan 2” 110

kV substations and to “Yeghegnadzor” 220/110/35 kV substation. In addition to these

lines, in case of this option, a 12 km 35 kV overhead single-chain line will be built and

connected to “Gorayk” 35 kV overhead line.

The 110kV rated utility transmission lines will be the primary source for supplying

power to the mine site. The additional 35kV line option will serve backup power in case

the 110 kV line fails. The 35kV line will be capable of supplying power to only few

processes in the plant to keep critical equipment on-line for facilitating a safe shut-down

or for keeping critical processes in operation till the primary source of power can come

on-line. The utility transmission voltage (110 kV and 35kV) will be stepped down to 6kV,

at the main substation, for reticulation around the site. From there the power would be

distributed to the crushing plant, waste dump area and water treatment plant, conveying

and stacking system, and the ADR plant. All these areas will have their own dedicated

transformers, where the 6 kV power will be stepped down to 400/220 V.

18.1.8 Power Distribution

Upon review of the most recently proposed equipment list, a total electrical load

of approximately 22.6 MW was determined. The electrical load is summarized in

Table 18.2.

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Table 18.2

Mine Power Requirements (by Area)

Electrical Load

Area

(MW)

Crushing Plant 6.7

Overland Conveying + Stacking 4.9

Solution Management Pumps 3.1

ADR Plant + Camp 4.0

Water Treatment Plant + River Pump Station 1.35

Exploration Camp + Offices + Mine Shops 0.8

Reserve capacity (Future additions) 1.75

Total 22.6

The electrical system upon entering the mine site will be configured as a radial

type system. The utility transmission line voltage (110 kV and 35 kV) will be stepped

down to 6 kV at the mine main substation, via two 25 MVA, 3 phase, 50 Hz

transformers. These transformers will provide power to 6 kV switchgear consisting of a

main circuit breaker, a tie breaker, and several feeder breakers for distributing power to

the crushing area, waste dump area & water treatment plant, overland conveying and

stacking, heap leach area, ADR plant and camp site.

The mine main substation is located in the crushing area. The crushing area

power requirements include primary crushing, secondary crushing, tertiary crushing and

screening, lime addition, administration offices, warehouse and mine pit requirements.

The 6 kV overhead power line to the waste dump area will distribute power to the water

treatment plant, exploration camp, fresh water pumps and other facilities in those areas.

There will be three (3) 6 kV overhead transmission lines running along the overland

conveyor route; the first to provide power for overland conveying and stacking, the

second to provide power to the heap leach area and ADR plant, and the third to provide

power for the camp site and other miscellaneous buildings in the area.

Depending on the load, the distribution voltage of 6KV will be utilized directly

(operating voltage for motors greater than 200 kW) or it will be further stepped down to

a 400 VAC, 3-phase, 3-wire system for feeding motors below 200 kW. The 400 VAC will

be further stepped down to feed lighting loads at 400/230VAC and 120 VAC to facilitate

instrumentation requirements and general office equipment (receptacles, computers,

printers, etc.). Power distribution design will follow the federal, state and local

standards.

The mine site will be provided with a grounding grid to which all building steel,

equipment, etc. will be connected for safety. This grounding grid will consist of a #4/0

AWG bare copper conductor buried below ground connecting all items previously

mentioned. All above ground connections except connections to building steel will be

mechanical type connections so that equipment can be removed or replaced easily. All

underground connections including those to building steel will be of the thermoweld

type. A test well will be provided for periodically measuring / testing the resistance of the

ground grid. Grounding design will follow the federal, state and local standards.

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Lighting will be of the high intensity discharge type. High pressure sodium type

light fixtures will be utilized for exterior areas and high bay interior applications. Metal

halide lighting fixtures will be utilized indoors for low bay application and where color

rendition is a factor. Fluorescent lighting fixtures will be used in interior applications

such as office lighting, electrical rooms, etc. All areas will be equipped with emergency

light fixtures utilizing battery packs which will provide a minimum of 90 minutes of

illumination. Lighting levels will be designated by the Illumination Engineering Society

(IES) published guidelines.

A computer based data gathering system, Supervisory Control And Data

Acquisition System (SCADA), will be incorporated in the control and monitoring of all

process operations. The SCADA system will use remote termination devices to channel

appropriate control and monitoring signals from field locations back to the central

processing unit (CPU) computer where an operator can physically operate equipment

from his computer work station. The SCADA system will be based on equipment types

preferred and designated by the Owner. The configuration of the SCADA will be based

on the latest industrial standards. A programmable logic controller (PLC) system will be

installed in respective areas, gathering information from the input and output signals

from instruments and motor control equipment. The SCADA will process and record all

communications with respective PLCs. An uninterruptable power supply (UPS) will

provide power to each PLC.

Standby diesel generators will be provided to handle emergency situations at the

heap leach pad area and ADR plant, respectively. These generators will be connected on

the secondary side of the distribution transformer in respective areas. A 4000 kW

generating station at the heap leach pad will provide power to select solution

management pumps and other equipment that may affect the process production line

should they stop operating. A 1000 kW, 480V rated generator at the ADR plant will

provide power to agitators, sump pumps and other equipment that may affect the process

production line should they stop operating. A 225 kW generator by the water treatment

plant and one more 225 kW unit by the crushing plant for emergency situations.

18.2 Site Development

The Project will require development at the following major locations:






The mining areas

The mine surface facilities, including the mine administration building, truck

shop, mine workshops, refuelling area, mine control areas and explosives yard

The crushing plant area

The ADR plant, leach pad and storage ponds

Waste dump and water treatment plant

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Road and site access

Power line tie into the local utility

Raw water sourcing and distribution

The ESIA team provided regular input during the feasibility study preparation into

the site selection and design decisions for all major infrastructure to ensure that

environmental, health, safety and social considerations inform the mine design process.

The following describes the engineering site preparation requirements at each

location. The proposed overall general arrangement layout drawing is shown in Figure

18.1.

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Figure 18.1 - Proposed Overall Site General Arrangement Layout

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18.2.1 Mine Surface Facilities

The mine service area will be supplied with power by an overhead power line from

the plant. The mine administration and warehouse will be connected to the main PABX at

the administration building. An optical fibre cable will connect mine service area

computers to the main server. Potable water will be supplied from the water treatment

facility at the plant.

The mine operations services area will include the following facilities:










Mine Administration Office

Cleaning

Mine Cafeteria

Heavy vehicle workshop / store and washdown bay

Light vehicle workshop

Heavy vehicle fuelling station

Light vehicle fuel station

Fuel storage

Magazine

18.2.2 Crushing Plant

The ROM pad and crushing plant will be located to the North of the main open pit.

Sub-surface conditions will be further defined with additional geotechnical testing

for building foundations and will be supervised by Golder. This drilling is in addition to the

extensive geotechnical work already undertaken and is scheduled to be completed by the

end of the 2012 drilling season.

The crushing plant site will be cleared and grubbed to remove organic material,

contoured for drainage and then capped with laterite to allow heavy vehicle traffic during

construction. There is extensive laterite available in the area.

18.2.3 Leach Pad and Collection Ponds

The leach pad and ponds are described in the heap leach facility write-up in

Section 17.2.

18.2.4 Waste Dump Facility

The waste dump facility (WDF) at Site 13 consists of the waste dump (WD), and

an influent equalization basin (IEB), wastewater treatment plant (WWTP), and

evaporation pond (EP), located downgradient of the WD and utilized for the collection

and treatment of mine-influenced water draining from the WD. Diversion channels will

be constructed upgradient of the WDF to divert storm and snowmelt runoff from

upstream catchments away from the WD, IEB, WWTP and EP.

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Waste Dump

The WD will be constructed in three phases. The WD phase areas will be

approximately 465,500 m 2 , 506,800 m 2 and 360,100 m 2 for Phases 1, 2 and 3,

respectively, for a total WD area of 1,332,400 m 2 . Waste material will be deposited on

the WD in nominal 8 meter thick lifts at the natural angle-of-repose of approximately

1.4H:1V. Benches with a nominal width of 16.8 m will be constructed between lifts to

provide an overall exterior waste pile slope of 3.5H:1V, to be compatible with closure

requirements. The Phase 1 WD may be constructed in sub-phases to further minimize

initial capital costs.

The Phase 1 waste will be deposited on the Phase 1 WD in the southern portion

of the Site 13 valley in 12 horizontal lifts to a top surface elevation of 2430 meters. The

Phase 1 waste is approximately 20 Mt and will be placed in the WD during the initial two

years of operations. The Phase 2 waste will be deposited north of the Phase 1 WD to fill

the valley to elevation 2430 meters. The Phase 2 waste is approximately 54 Mt that will

be deposited in three years. The Phase 3 waste will consist of depositing nine additional

horizontal lifts above the Phases 1 and 2 WD to a top surface elevation of 2502 meters.

The Phase 3 waste is approximately 84 Mt to be placed in six years, and will bring the

total Phases 1, 2 and 3 WD capacity to approximately 158 Mt. This capacity may be

reduced slightly when considering access ramps within the WD and operational

constraints. The WD may be expanded higher up the hillside to the southwest to

accommodate additional waste material if needed.

The WD will be lined with a 0.45 meter thick compacted low-permeability soil

liner. An underdrain system will be constructed within the WD footprint beneath the soil

liner to drain non-contact groundwater/subsurface seepage to the IEB and prevent the

seepage from entering the waste above the WD base liner. The WD will have 1.5 meter

high perimeter berms to prevent rainfall and snowmelt water within the WD that comes

in contact with the waste (contact water) from overflowing the WD. This water will be

collected by an overdrain system constructed above the WD base liner and routed to

the IEB.

Influent Equalization Basin

The IEB was sized in accordance with the project design criteria to store 24

hours of the WD maximum estimated underdrain flow plus overdrain flow from the 100-

yr/24-hr storm event (snowmelt and precipitation), and to provide flow control to the

WWTP. The IEB storage capacity is approximately 739,400 m 3 to the 0.6-m freeboard

depth. The IEB will be constructed during the Phase 1 WD construction by building an

earthen dam across the narrow valley downgradient of the WD. The dam crest elevation

will be approximately 5 meters lower than the WD downgradient toe elevation.

Groundwater/subsurface seepage flow from the WD underdrains and

rainfall/snowmelt contact water from the WD overdrains will be routed to the IEB. The

collected water will be pumped from the IEB where it will be tested for compliance with

discharge criteria, and if needed, routed to the WWTP for treatment.

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The IEB will have a composite liner system consisting of 2-millimeter (80-mil)

thick single-side textured HDPE geomembrane with texturing at top for traction,

underlain by 0.3-meter thick compacted low-permeability soil liner. An underdrain

system will also be constructed within the IEB footprint beneath the liner to drain

groundwater/subsurface seepage to a collection sump located downgradient of the IEB.

Water collected in the sump will be tested for quality and released if non-impacted, or

pumped to the IEB if the water quality criteria are exceeded.

Wastewater Treatment Plant

The WWTP will receive water from the IEB. Treatment processes have been

developed based on the projected water quality characterization of the combined flows

from underdrains and overdrains. The IEB and WWTP capacities have been designed

to accommodate high flows associated with snowmelt, with operation of the WWTP at a

constant rate for about eight months per year. Final treated effluent water quality

discharge requirements are yet to be finalized. The WWTP effluent is projected to

comply with Armenian maximum allowable concentration (MAC) Category II standards.

Category III standards (more lenient) have been considered, but the conceptual design

and cost estimation for the WWTP is conservatively based on the more stringent

Category II effluent targets.

The design flow rate for the WWTP is 182 m 3 /hr. The WWTP will operate 24

hours per day, seven days per week for eight months per year (roughly April through

November). High flows in spring will accumulate in the IEB and will be gradually treated

through the drier summer months. The final unit operation in the wastewater treatment

process is a spray-enhanced solar evaporation pond. Use of the evaporation pond limits

the operational season for water treatment. The IEB, WWTP and EP are conceptually

sized with capacity to treat twelve months of accumulated flow in the 8-month operating

season.

The contaminants of potential concern (COPCs) are based on comparison of the

projected influent water quality characterization and the Category II discharge

standards. COPCs include metals, sulfate, and suspended solids. Treatment operations

for these COPCs include:




Chemical precipitation (lime treatment) for metals removal;

Microfiltration for suspended solids removal; and

Reverse osmosis for sulfate removal.

All ancillary equipment (chemical reagent feeds) have been included in

conceptual design, as well as secondary waste handling equipment (dewatering

chemical precipitation sludge).

Treated water will be discharged to the Vorotan River. Secondary waste sludge

from chemical precipitation will be suitably disposed of on site in the WD.

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Solids accumulated in the evaporation pond may be removed for disposal or

disposed in-place at the end of the WWTP life.

The WWTP is expected to operate in post-closure mode for some period of time,

currently estimated at 10 years. Further study of post-closure flows from the overdrains

and underdrains is needed to more accurately predict post-closure water treatment

requirements and duration of operations. Golder has completed net acid generating

(NAG), synthetic precipitation leaching process (SPLP), and humidity cell tests on

representative waste rock types. This is summarized in Section 24.

Evaporation Pond

Reverse osmosis brine from the WWTP will drain by gravity to the EP for

evaporation. The EP was sized to meet the brine storage requirements. The EP storage

capacity is approximately 61,420 m 3 to the 0.6-m freeboard depth. The EP will be

constructed during the Phase 1 WD construction by excavating into the sloping terrain

on the uphill side and filling earthen embankments on the downhill side. The pond crest

elevation will be 20 m higher than the IEB crest level.

The EP will have a composite liner system consisting of 2-millimeter (80-mil)

thick single-side textured HDPE geomembrane with texturing at top for traction,

underlain by a 0.3-meter minimum thick compacted low-permeability soil liner. An

ultrasonic system will be provided for the EP to prevent birds from accessing the fluid in

the pond.

A stock-proof mesh fence with locking gates will be constructed around the

perimeter of the IEB, WWTP and EP for public safety and to deter unauthorized access

into the waste water treatment area. The fence will also prevent wildlife from reaching the

fluid in the ponds.

18.2.5 Accommodations

The final strategy for accommodating all construction personnel, employees and

security personnel during the construction period will be defined as part of the detail

engineering effort. The basis of the cost estimate included an allowance to house 200

Lydian employees on-site and the remaining employees in nearby towns. Employees

who live outside the area would be placed in hotels while local employees could live in

existing accommodations. The contractor will provide housing for all construction

personal and this cost was included in the construction labor rate.

Prior to detail engineering of the on-site housing facility potential vacant or run

down housing opportunities will be investigated in Jermuk.

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18.2.6 Roads & Site Access

For supplies, material and equipment can be shipped to the ports of Poti or Batumi,

Georgia then trucked through Georgia and Armenia to the Amulsar project site. Airfreight

through the Zvartnots International Airport in Yerevan is also possible.

There is a sealed road from Yerevan to the Iranian border passing to the south of

the project area and a sealed spur road to the town on Jermuk. The current project

access is gained via a gravel/dirt road from the Jermuk road. A further gravel/dirt road

runs along the Vorotan river valley to the town on Gorayk. The sealed roads to the site

turn-off are adequate for all Project transport requirements. The existing gravel/dirt site

access road to the mine site will need to be widened over its entire length of 20 km as

noted below, and maintained for all weather operation, providing the main means of

access to the mine site and associated infrastructure The gravel/dirt road from Gorayk

can also be used to access the Amulsar site and will also require upgrading.

The roads required for the Project are:







Plant access road

Village access road

Leach pad & ponds access roads

Waste dump and WWTP area roads

Mine haul roads

Borefield access road

18.2.7 Non–Process Buildings

The location of the various site buildings is shown in Appendix 3. The following

non process buildings have been included in the capital cost estimate:












Administration and engineering building

Plant Warehouse / Workshops

Maintenance Shop

Mine Truck shop

Lube Storage

General Storage

Truck Wash Station

Guard Gate

Gas Pump Building

Explosive magazine

Camp including the following;

a. Dining Area

b. Sleeping Quarters

c. General Store

d. Laundry

e. Infirmary

f. Sewage Treatment plant

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Laboratory services for exploration, the mine and process facilities will be

supplied by ALS Minerals (ALS). ALS has expressed interest in constructing a

laboratory either on site or in close proximity to serve not only the mine but other

regional requirements. Based on an analytical requirement volume and determinations

estimate, ALS will supply and staff the laboratory accordingly. Lydian will make

payments monthly to ALS for the analytical support. For the FS, this monthly cost is

estimated as the staffing cost of a Chief Chemist, two additional chemists, four assayers

and four sample preparation technicians plus US$10 per sample and 40,000 samples

per year.

18.3 Water Source

The water source for the project will come from the Vorotan River with potable

water delivered as required. A sitewide water balance was prepared as part of the

Integrated Water Studies by Golder and adequate supply is available for production,

makeup water, and general use in supplementing mine infrastructure needs for additional

requirements such as dust control.

18.3.1 Potable Water Supply

Potable water will be used for drinking water, cleaning, change rooms, laboratory

water and safety showers.

Potable water is not required for the process requirements.

The design potable water demand is 115 m 3 /day based on 300 l/person per day in

the staff quarters and 40 l/day for staff not resident in the quarters. A further 70 m 3 /d will

be required within the plant (ablutions, laboratory, safety showers, etc). Accordingly, a

supply of 370 m 3 /d has been allowed and will be purchased from local community

supplies.

18.3.2 Raw Water Distribution System

Based on the water balance study and hydrological assessment, there is adequate

raw water available from the Vorotan River. An abstraction permit will be required in

accordance with Armenian legislation and at this time it appears there is no reason to

suspect that this will not be granted. KDE has included a raw water storage and

distribution system for the project that pumps to the following areas;







ADR Plant Area

Camp Area

Water Treatment Plant

Truck shop area

Crusher Area

Mine Area

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In order to minimise the number of services it is proposed to provide firewater via

the raw water system. A diesel driven pump will start automatically on loss of raw water

pressure to provide a secure fire service. A minimum volume of water will be held in the

raw water pond at all times.

For exploration drilling purposes Geoteam currently holds a water use permit from

a small pond on the western side of the pit.

Raw water required for the operation of the plant will be sourced from the Vorotan

River. It is assumed that the well for raw water will be located on the shore of the river

somewhere downstream of the hydro-electric power plant. The tentative location for the

extraction well is shown on site plan. A perforated pipe with an installed pump (i.e, sump)

will be buried adjacent to the river and surrounded with drainage gravel to allow water to

flow into the sump and be pumped through the delivery piping system. A local contractor

will be utilized to design and build the extraction well to assure this meets local codes.

The exact location of the extraction well will be determined during the detailed

engineering phase of the project and after getting necessary environment permits.

Geoteam is completing the permitting process required to source water from the river for

this project.

The water is pumped at the required rate to an adjacent tank located out of the

flood plain. From the storage tank, the water will then be pumped to ADR and crushing

plant through 4 and 3 km high pressure buried pipeline with a booster pump station if

required.

The electrical power of 6 kV will be brought to site by a overhead power line

spurred from ADR plant which then will be stepped down to 400V for use.

18.3.3 Process Water Supply

Process water will be prepared at the process plant and will be recycled to the

extent possible. Makeup water will be kept to a minimum. Process water quality will be

monitored and, provided it is acceptable, will be used in the following areas:





Leach ponds

Screen sprays

Carbon transfer

Process plant washdown

18.3.4 Sewage Waste Water Treatment

Sewage waste water treatment will be required at the man camp, ADR process

plant, crushing plant, truck shop and contact water treatment plant area. An allowance

was included with the building costs to include a septic system for each of these

facilities.

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In the event that geotechnical testing indicates that a septic system is not

appropriate due to ground conditions then a pre-engineered sewage waste water

treatment system would be placed at the man camp and sewage from the other

locations would be hauled to this facility for treatment. Regardless of the type of

sewage treatment facility required this facility will be designed to meet local regulatory

requirements.

18.4 Waste Disposal

Two landfill disposal sites will be constructed in accordance with the EU Landfill

Directive for non-hazardous and hazardous waste. The sites are small and would be

adjacent to each other within the Rock Allocation Area.

Engineering requirements for a hazardous waste landfill/cell are a basal and side

wall liner with a permeability and thickness equivalent to 1 x 10 -9 m/sec at 5m thick.

Engineering requirements for a non-hazardous waste landfill/cell are a basal and side

wall liner with a permeability and thickness equivalent to 1 x 10 -9 m/sec at 1m thick.

The Directive states that where the geological barrier does not naturally meet the

above conditions, it can be completed artificially, but must be no less than 0.5m thick

(again on the base and side walls) and be of equivalent standard (i.e. for a hazardous

waste landfill/cell it would have to be equivalent to 1 x 10 -9 m/sec at 5m and for a nonhazardous

landfill/cell equivalent to 1 x 10 -9 m/sec at 1m).

Non-hazardous waste generation is estimated to be less than 5,000 t. Hazardous

waste generation is estimated to be considerably less. The landfills would be

constructed with leachate management and treatment systems.

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19.1 Marketing Studies

19.0 MARKET STUDIES AND CONTRACTS

The product from the Amulsar Project will be doré bars containing a mixture of

gold and silver and other impurities. The precious metal content of the bars is estimated

to be between 90 and 99 percent gold plus silver. Doré bars produced at the mine will

be weighed and assay samples collected. These high grade samples will be analyzed

both on site and at an independent laboratory. The weight of the bar combined with the

assay values allows the calculation of the ounces of gold plus silver contained in each

bar and thus the overall value. Typically gold and silver doré bullion is sold through

commercial banks and metal dealers. Sales prices are obtained on the World Spot or

London fixes and are easily transacted.

The doré bars will be shipped by a secure carrier to a precious metal refiner,

probably located in Europe or Asia. Upon arrival at the refinery, the bars are weighed and

samples are taken to determine the precious metal content. The refiner will schedule

periodic processing of the Amulsar doré in separate crucibles. The products from the

refinery are separate refined gold and silver ingots known as good delivery bars. The

option exists to take physical metal or to employ a trading account to monetize the bullion.

Once the mine has established an operating history at the refinery, payment of

typically 90 percent of the estimated shipment value will be forwarded at the company's

account at the commercial bank that manages the gold and silver sales for the company

as the bullion is transferred from the company to the secure carrier. The remainder of the

money is transferred to the client once final bullion assays have been agreed upon by the

refiner and the company. Usually the company CFO manages the account as a source of

immediate funds or, alternately, gold and silver can be kept in inventory. Typical shipping

and refining costs are approximately US$ 5 per ounce of gold refined.

19.2 Contracts

As of this writing, the company has not entered into contractual agreements with

civil contractors or engineering, procurement and construction management contractors.

However, potential contractors have been interviewed and shortlisted.

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20.0 ENVIRONMENTAL STUDIES, PERMITTING AND

SOCIAL OR COMMUNITY IMPACT

Wardell Armstrong International (WAI) was instructed by Lydian International

Limited (Lydian) to undertake an Environmental and Social Impact Assessment (ESIA)

of the Amulsar Gold Project in the Republic of Armenia (RoA). This reporting process

involved the following key steps:

• Preparation of a Scoping Study by WAI - to set out the main project

parameters, outline national legislative and international best practice

requirements and identify any potential environmental and social impacts;

• Baseline data collection - designed in accordance with the above, with

data collection principally undertaken by in-country specialists with input

from WAI (in relation to the social baseline conditions), Golder

(contributing to seismicity and water resources).

• Impact prediction, assessment and mitigation, with regular interaction with

the Amulsar Feasibility Study (FS) team; culminating in the:

- Preparation of the ESIA document - disclosure draft due for

submission in Q3 2012;

- Preparation of Framework Environmental and Social Management

Plans, Stakeholder and Community Engagement Plan and

Framework Mine Closure Plan; these framework management

plans set out principles and outline management strategies for

Lydian, and form the basis of an Environmental and Social

Management System to be developed and adopted by Lydian for

the Amulsar gold Project.

In order to produce an ESIA to satisfy international requirements, WAI’s remit

has been to review and incorporate data and reports collected and prepared by

Geoteam, Golder and other FS contributors, together with various appointed Armenian

and international specialists. Significant specialist contribution has been provided by Dr.

Clive Hallett (acid rock drainage), Eddie Jewell Associates (noise data modelling), SKM

Colin Buchanan (traffic impact assessment), Environmental Resources Management

Group (archaeology and cultural heritage), Dr. Joanna Treweek (biodiversity), Shape

Consulting Limited (community health) and Radman Associates (Radiation Protection

Advisors). Therefore, in addition to the above listed reports, various supporting

deliverables and activities have also been undertaken throughout the ESIA process.

These include formal and informal stakeholder engagement and the iterative integration

of environmental and social considerations within Project design and development.

Information provided by third parties has been referenced as appropriate and has been

detailed in full in the ESIA.

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Where relevant, key environmental and social aspects have been included and

identified throughout this FS. The focal elements of the ESIA process are summarized

in this Section and the full ESIA is presented as a separate document. Aspects covered

in the ESIA, such as geology, geotechnics, geochemistry and seismicity are fully

detailed in the ESIA and geochemistry and seismicity have been further discussed in

Section 24. For completeness this report should be read in conjunction with the full

version of the ESIA (WAI Report No. EO-52-0088-2); and the Scoping Study (WAI

Report No. EO-52-0088-1, February 2011).

It should be noted that, whilst exercising all reasonable diligence in checking and

confirming it, WAI has relied upon the data presented by others in undertaking the ESIA

and cannot comment on the adequacy of actual field sampling undertaken, laboratory

procedures or any interpretation of data by others. Some of the documents reviewed by

WAI have been translated from Armenian, necessitating WAI to interpret and use the

information with caution.

Erato

The ESIA has been undertaken for the extraction and processing of ore from

Tigranes and Artavasdes up to and including Year 12 of the Project. Whilst resource

drilling has been undertaken at Erato, currently the resource is indicated only and has

not been included within the ESIA.

In terms of the FS, the economic contribution and resource allocations

associated with the commencement of mining of Erato in Year 7 of the Project has been

identified, in that provision has been made for the waste from Erato to be

accommodated in the current WD, with potential for progressive backfilling of Tigranes

and Artavasdes open pit. Similarly, Erato ore will form the late stages of the proposed

development of HLF. However, the environmental and social studies required with

respect to mining operations at Erato will require full assessment in an ESIA addendum

to be completed at a later stage of the Project. Additional studies will necessitate the

consideration of baseline data, including biodiversity field studies, hydrogeological field

data and modeling, soils and land capability assessments, together with further ARD

characterization and extended visual impact analysis. Stakeholder engagement would

also take place to inform and explain the nature of these studies.

20.1 Location, Environmental and Social Setting

The Project is located in central southern Armenia and straddles two administrative

provinces, or Marzer, namely Vayots-Dzor and Syunik. The Project area is largely open in

nature with no areas of woodland and is characterized by a temperate climate of long cold

winters and short relatively cool summers. The typical landscape at the Project is shown

in Photo 20.1 below.

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Photo 20.1 Typical Landscape at the Project

The landscape ranges in altitude from approximately 1500 masl to a ridge of

2988 masl, where the gold deposits are located (Amulsar Mountain). Elevated areas are

rocky rounded mountain ridges with steep sided slopes, leading to large undulating

plateaus and river valleys, some of which are locally incised by gorges of the Rivers

Arpa and Vorotan. The environment is relatively pristine, being unaffected by any

industry in the immediate area, and is characterized by grassed foothill meadows,

prairies and sub-alpine to alpine landscapes as the elevation increases. The River

Vorotan and associated catchment is located within the license area. The Vorotan feeds

the Spandaryan Reservoir, located approximately 5 kilometers to the south of the

proposed heap leach facility (HLF).

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Figure 20.1 Project Environs showing proposed infrastructure and regional setting (rivers, villages topography). Jermuk is situated on the H-43 highway, to the north of Kechut

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There are a number of ephemeral and permanent surface water features in the

foothills of the Project area, supporting a range of natural and semi-natural habitat and

fauna. The land use within the project area includes; seasonal grazing land (moderately

high elevations within the alpine meadow grasslands), hay cropping and winter grazing

(lower elevations, supporting sub alpine meadow grasslands). Local residents and

visitors make use of the Vorotan, Arpa and Darb river systems for recreational fishing. A

small proportion of local residents also hunt for recreation within the Project area.

Three rural communities with a combined population of approximately 2000

people lie within the Project vicinity. These include the communities of Saravan

(consisting of the villages of Saravan, Saralanj and Ughedzor, which is only inhabited

during the summer months), Gorayk and Gndevaz. The communities of Saravan and

Gndevaz are in Vayots Dzor Marz, some 5 - 9 km west and southwest of the deposit.

Gorayk village is located in Syunik Marz, and is situated 15km from the centre point of

the Project; the edge of the village is approximately 1.8km south of the HLF and 5km

from the south east of the open pit. The main livelihood for villagers is subsistence

agriculture.

The closest city is Jermuk (which includes the associated village of Kechut)

which is located approximately 14 km and 7.5 km from the north-west of the proposed

open pit and the waste rock dump respectively. Jermuk and Kechut together have

approximately 6000 residents. Jermuk has several natural hot springs, health resorts

and spas. Jermuk hosts an established mineral water bottling plant and emerging

tourist industries.

20.2 ESIA and Permitting

20.2.1 Scope of the ESIA

The ESIA fully describes the policy, legal and administrative framework under

which the Project will be developed and under which the assessment was carried out,

as well as a description of the Project covering geographical, ecological, social and

temporal aspects. It includes baseline data describing the physical, biological, cultural

and historical conditions and the environmental and social impacts associated with

project implementation. Mitigation measures needed to minimize impacts to an

acceptable level are presented, as well as an analysis of feasible alternatives. Key

framework management plans covering environmental, health and safety, social

management and community development have been formulated and presented,

together with an Environmental and Social Action Plan for the delivery of the Project

from construction to operation and eventual closure.

The integration of the ESIA team with the specialists engaged in the FS has

allowed many potential impacts to be prevented or designed out at early stages of the

study. Similarly, the integration process provides the means for appropriate and

practical mitigation measures to be included in the designs.

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While the submission of the international ESIA is not an Armenian regulatory

requirement (the Armenian EIAs (ShMAGs) fulfill this role), the ESIA will be made

available to the Ministry of Nature Protection and other Government departments. The

ESIA will also be circulated to relevant financial institutions and made available publicly

in both Armenian and English language versions. Public input is welcomed and will be

considered in the decision-making process of the relevant financial institutions. In this

regard the findings of this study will be used in the compilation of the final ShMAG

documentation, based on the detailed design of the project, and submitted for approval

by the government authorities.

The ESIA is prepared in line with the requirements of an international standard

ESIA, specifically the International Finance Corporation Performance Standards (IFC-

PSs) and the European Bank of Reconstruction and Development’s Performance

Requirements (EBRD-PRs). The objective is that the Project will be acceptable to IFC

and EBRD, and other financial institutions that are signatories to the Equator Principles

(EPFIs). In January 2012 the IFC introduced updated Performance Standards, and it is

these 2012 PSs that have been applied in the compilation of the ESIA.

The Equator Principles apply the IFC’s environmental and social screening

criteria, to reflect the magnitude of impacts understood as a result of assessment:

• Category A - Projects with potential significant adverse social or

environmental impacts that are diverse, irreversible or unprecedented and

may affect an area broader than the site facilities subject to physical works;

• Category B - Projects with potential limited adverse social or environmental

impacts that are few in number, generally site-specific, largely reversible and

readily addressed through mitigation measures; and

• Category C - Projects with minimal or no social or environmental impacts.

The IFC and EBRD have been investors in Lydian International since the

company began early exploration activities on the Project. Accordingly the ESIA

examines the potential environmental measures needed to prevent, negate, minimize,

mitigate or compensate for adverse effects, and to improve environmental performance,

whilst seeking to optimize the positive benefits that the Project may accrue. On the

grounds of the Project being classifiable as Category A, the ESIA is required to

integrate environmental and social considerations into Project design and to conduct

consultation and disclosure accordingly.

Although the Amulsar Project has the potential to incur environmental and social

impacts, the ESIA and this Section of the FS demonstrates that these are manageable

to avoid, prevent or to reduce to acceptable levels, in accordance with Armenian and

international standards.

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20.2.2 Republic of Armenia Environmental Impact Assessment

The requirements under the RoA Environmental Impact Assessment procedure

(ShMAG) are slightly different in terms of process, method and presentation, to those

required of an ESIA by international funding agencies. Therefore two processes were

undertaken, following parallel paths and based on common baseline data and project

parameters.

WAI has provided input into the conceptual ShMAG reports for the crushing,

conveying and heap leach facility (HLF). Under Armenian law the mining operation and

the HLF/ADR operations are permitted separately and thus require separate

submissions. The ShMAG reports have been prepared by in-country experts ‘Eco Audit

LLC’, based on baseline data principally provided by Geoteam.

20.2.3 Permits and Licensing

Subsequent to the exploration phase, and prior to development of the mine,

several permits and licenses will be required. These include those outlined in Table 20.1

below.

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Table 20.1

RoA Permits Required for Development of Amulsar Mine

License/Permit Title Application/Provision Status Comment

Mining Licence To permit extraction of ore Granted. Valid until 2034.

Technical Safety Approve that the design follows all Armenian safety regulations. Granted. Valid until the life of the ML, unless there are changes in the design of the Open Pit operations.

Rock Allocation Area

Change in land use from agriculture to industrial required to accommodate all

mining infrastructure and get construction permit

Granted for open pit, waste rock dumps & crushing. Valid until 2034.

Once the Concession Agreement is signed with the Government, the new RAA will be granted which will include all the HLF and parts of general

infrastructure. As part of RAA the land status change will be done automatically after RAA is granted with defined limits.

Water abstraction & discharge licence To permit the use and the discharge of water. No application has been made as yet. The company has a water extraction permit for exploration activities.

Air emission permit To permit the emissions to the Air. Not yet applied for. The company has air emission permit for exploration activities.

Explosives permit (store, transport, use) To permit the use and the storage of the explosives material.

The Company will contract a company that will have both blasting and storage permits. As such, Geoteam will not need to get these permits.

ICMC cyanide supplier compliance Company is committed to become ICMI compliant, thus the transporter and The Company will purchase CN from a producer that is ICMC compliant, or working on becoming compliant, taking professional advice from a ICMI

the producer should be compliant as well.

Lead Auditor to ensure that viable options are in line with the ICMC..

Construction and Architecture permits To get the approval that all the design corresponds to Armenian Standards Not yet applied for.

and Norms.

Gas and power use designs and To permit the gas and power use.

Not yet applied for.

construction expertise and permits

Waste Passports

To give the class of hazard to the different waste types and permit the locating Not yet applied for.

of the waste and its disposal.

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20.3 Significant Project Consumption and Releases

The ESIA details the effects and influences of the Project, which are

significant for the environmental and social impacts (both positive and negative). The

most significant are summarized here and detailed in the ESIA:

Land Take - the Rock Allocation Area (the area within which mining activities

are authorized) will be approximately 4,700 hectares (ha). While Lydian intends to

maintain seasonal nomadic grazing access to the majority of land under its control,

and grazing does not occur on all of the 4,700 hectares due to altitude and

topography restrictions, at least a small proportion of land currently in use for grazing

is likely to be sterilized, either permanently or temporarily.

Permanent sterilization would occur at the:



Waste rock dump; and

HLF.

Temporary sterilization will result from the development of mine infrastructure

which will be in place for the period of mining and includes the following elements:

• Waste rock stockpile;

• Waste water treatment plant & basins;

• Conveyor;

• Utilities;

• Mine camp;

• HLF ponds;

• Truck shop;

• Exploration camp;

• Maintenance shop & offices;

• Explosives magazine; and

• Road to crusher.

Certain haul routes and access roads will be retained as a part of the mine

reclamation plan for longer term maintenance and aftercare management.

The ESIA provides details of the footprint for individual areas within the mine

site. In summary (see Figure 20.2), during operations, it is estimated that the area of

direct disturbance from the development of the mine and associated infrastructure

will be in the order of 510 hectares. However, because elements of the infrastructure

such as haul roads, the conveyor and site access roads will also reduce access for

grazing and other recreational activities, it is estimated that a further area of

approximately 260 hectares will be indirectly affected as consequence of fencing and

bunding. In addition, there will be approximately 270 hectares where access to

grazing land will be restricted by controlled passage of animals via crossing points

on mine haul roads.

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Figure 20.2 Footprint of Mine Development (throughout the operational life)

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Labor and Services - Approximately 600 people will be directly employed

during construction, 550 during various phases of the operation and 20 for closure

care and maintenance. In addition, where services are outsourced, local contractors

will be offered the opportunity to bid for tenders.

Energy and Diesel Consumption - Total electrical load is estimated to be

22.6MW. Total diesel consumption will be 9 million litres /year (based on

consumption estimated in year 2 of the Project).

Air Emissions - Dust is defined as particulates from 1 to 75 micrometers in

size, however, 95 percent of dust will normally be over 30 micrometers in size and

will be subject to aerodynamic and gravitational effects, which determine the

distance they travel before settling out of suspension. Main Project dust sources will

be open pit excavations, crushing plant/overland conveyor, ore stacking and

associated haulage and deposition activities. The latter is chiefly haulage of waste

rock from the open pit to the waste rock dumps and ore from the pit to the run-ofmine

(ROM) ore stockpile and crushing facility.

Greenhouse Gas Emissions - The calculated annual emissions of CO 2 and

other greenhouse gases are predicted to be 91,900 Total CO 2 e/yr (tonnes),

predominantly from grid electricity consumption for crushing and grinding (c.67,800

tonnes per annum) and diesel consumption (c. 24100 tonnes per annum).

Greenhouse gas generation has been based on published data, however it should

be noted that there is no coal power generation in Armenia, which relies on a

combination of nuclear, hydroelectric and gas.

Water Uses and Discharges - Total water requirements for processing

(including HLF, crusher and dust suppression water) will average 1.9 Mm 3 per year

during Phase 1, rising to 3.2 Mm 3 per year during Phase 2. Average requirements

are broken down on Table 20.2. This water will be sourced from the Vorotan River.

Water for domestic and potable needs, requiring a supply of 136,000m 3 /yr

abstracted from a combination of local spring supplies and Vorotan River, which may

be shared with the new Gorayk water supply. Water sourcing and requirements are

further detailed in the Water Balance prepared as part of the Integrated Water

Studies by Golder.

Table 20.2

Average Water Requirements

Water Requirement m 3 /hour m 3 /day Supply Source

Potable water at camp 15.4 370

Local springs (share with new

Gorayk supply) or Vorotan River

2 x hoppers at crusher

Dust suppression pond/Influent

45.8 1100

site

equalization basin (IEB)

General dust

suppression

Process water for

leaching - Phase 1

Process water for

leaching - Phase 2

SUM (assumes summer

season during Phase 2)

19.7

22 (4 months/year)

272(4 months/year)

472(4 months/year)

Dust suppression pond/IEB

147 3,528 Vorotan River

147 3,528 Vorotan River

374.9 8,998

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Water discharge from the site will be from the WWTP, which treats the water

to an acceptable standard, and is designed to release water to the Vorotan

catchment at a maximum rate of 182m 3 /hour.

The HLF will be a closed system, and no discharge will occur from this area.

Foul water from the mine camp will be treated to a high standard, and released to

the Vorotan catchment.

Principal Wastes - The mine will generate diverse waste streams throughout

all development phases, including solid construction wastes, domestic and technical

(processing and assaying wastes) wastes, domestic effluents and runoff waters.

Liquid waste streams (effluents and residual liquids) will not be released to the

environment unless they conform to Armenian regulatory requirements and

internationally recognised concentrations. The management of wastes will be

undertaken in line with the various framework management plans, accompanying the

ESIA.

Waste Rock - The geochemical characterisation study, reported in full in the

ESIA, identifies that the majority of Amulsar waste rock lithologies have some Acid

Forming and Metal Leaching potential. All waste rock will be placed on the waste

rock dump sited to the north of the mine (see Figure 20.2) and contact waters will be

treated in a WWTP. Due to known properties of the waste rock, the WWTP has been

designed to capture, divert and neutralise all associated contact waters to meet the

regulatory standards. Detailed design and operation will be informed by an Acid

Rock Drainage Management Plan.

Water Treatment Sludge - These comprise the residues from the treatment

of surface water run-off and leachates from the waste rock dump. The sediments are

removed as a sludge containing metal hydroxides. This volume of sludge is

insignificant, with annual productions in the order of 69t/year (equivalent to 49m 3

/year). The sludge will be re-incorporated in the Waste Dump.

Ore - Ore will be stacked on the HLF which will be operated as a closed

system with adequate environmental protection (as described in Table 20.5). ARD

test work has been undertaken on ore material showing that it is also acid forming

and metal leaching and plans have been designed to manage this acid generation.

Any ROM stockpile will be of very short residence time, small-scale, lined, and

include runoff collection/diversion to a storage facility in order that contact waters are

treated prior to discharge.

Hazardous Materials & Reagents - These will include sodium cyanide

(NaCN), lime, caustic soda (NaOH) and hydrochloric acid (HCl), together with diesel

oil. Appropriate signage and MSDS will be used. Chemical-specific first aid training

will be provided to staff. The use, transport, storage and handling of cyanide at the

site will be controlled by documented management procedures and formal

management plans and procedures in accordance with the International Cyanide

Management Code (ICMC) and the ESIA framework Cyanide Management Plan.

Measures to avoid, respond to and treat spills and emergency situations are outlined

in the ESIA framework Spill Prevention and Emergency Response Plan.

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20.4 Environmental Context

20.4.1 Geology and Soils

The Amulsar high sulfide gold deposit is hosted in an Upper Eocene to Lower

Oligocene calc-alkaline magmatic-arc system. Detailed regional and site-specific

geology, together with the current block model, has been summarized in preceding

chapters and is detailed within the ESIA.

Soil types, broad characteristics and indicative pH have been identified in the

general Project area and over 2,000 exploration soil samples have been tested for

heavy metal content. Targeted samples have also been tested for extended

environmental suites including potentially toxic heavy metals ions and cations,

radiological parameters, hydrocarbons, cyanide and microbiology. A geotechnical

soil sampling regime within the proposed footprints of major mine infrastructure has

been undertaken. This data is considered in detail in the ESIA with respect to the

assessment of impacts on soil quality and land use.

20.4.2 Radioactivity

It is understood that some residents in the Project vicinity are concerned

about the impact of radioactivity arising from the Project as ‘radioactive dust’ or in the

form of radon. The main source of dust from proposed mining activities will be from

disturbed rock, and to a lesser extent, soil. Uranium (U) and Thorium (Th)

concentrations from over 2,000 samples of soil and 46,000 samples of rock (ore and

waste) have been provided by Geoteam from their extensive exploration program

undertaken across the license area.

The measured U and Th concentrations have been reviewed by Radman

Associates (Radman), a UK-based firm of accredited Radiation Protection Advisors

and the concentrations have been compared with typical reported levels of these

elements in Armenian soils (United Nations Scientific Committee on the Effects of

Atomic Radiation, UNSCEAR) and found to be below these levels.

Household surveys for baseline radon levels were conducted in households in

the villages of Gorayk, Saravan, Saralanj and Gndevaz. The measurements were

taken in 149 locations in December 2010 until March 2011 and indicated pre-existing

elevated radon levels in Gorayk village, primarily due to poor ventilation in housing.

20.4.3 Seismicity

Armenia is situated within the Caucasus region in the vicinity of the Alpine-

Himalaya seismic belt and at the juncture of the African, Arabian and Indian tectonic

plates. It is a global region of moderate to high historic earthquake activity. The

Amulsar license is located within a seismically active region of the Arabia-Eurasia

plate boundary zone.

Detailed studies evaluating the regional seismic profile and seismic hazard

class of the Project area were completed by Golder and were included in their report

titled Earthquake Hazard Assessment and Seismic Parameters for the Amulsar Gold

Project (Golder, 2012b), summarized in Section 24.

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The results of this study have been used to develop appropriate seismic

design criteria for major mine infrastructure in accordance with international and

Armenian guidance and building codes.

20.4.4 Water Resources

Groundwater Characteristics

The bedrock of the mountain and proposed open pit has been shown by

hydraulic testing to have a low permeability. Exploration has shown that the

alteration of bedrock to clay is extensive, and therefore the low permeabilities

measured are likely to be representative of the bulk geology of the open pit. The

mountain-top topography of the proposed open pit further reduces the groundwater

inflow potential, since, unlike a flat or valley setting, there are no adjacent waterbearing

strata which would drain towards the excavation; at least for the majority of

the open pit’s life. A 2D groundwater model has been generated for the open pit,

and indicates a maximum groundwater inflow of approximately 850 m 3 /day (9.8

L/sec).

Hydraulic testing at the WD has indicated that this is a groundwater discharge

area, with marginally higher groundwater inflows, but the aquifer here can still only

be considered of local importance here. The HLF area is situated within poorlyfractured

basalts, and also does not have significant groundwater potential. Neither

the WD, nor HLF construction will involve excavation into the saturated subsurface.

Infiltration of precipitation is limited by the low permeability of the geology,

leading to the development of many low-flow, short-pathway local mountain springs

at the open pit and WD areas. Numerous springs have been identified within the WD

and open pit areas, but it is likely that the number of active springs varies according

to season. Four of the springs appear to support/be associated with an area of

suspended marsh. Another spring drains to the Benik Pond within the Darb

catchment, west of the mountain. There are reportedly more springs at the WD area

and, as discussed below, Gorayk village.

Groundwater quality and flow (pathways and volumes), together with a

calibrated groundwater model, a Conceptual Site Model and site-wide water balance

are presented in the ESIA based on the Integrated Water Studies Report prepared

by Golder (Golder, 2012i).

Surface Water Characteristics

The Project license area is bisected approximately north-south by a

catchment divide, with the Darb catchment (a subcatchment of the Arpa) to the west,

and the Vorotan catchment to the east. Virtually all Project activities will take place

within the Vorotan River catchment. River flows at both of these catchments have

been significantly altered by human intervention, including the following:


The Vorotan River has a tunnel at Spandaryan reservoir which diverts flow

to Kechut Reservoir (as part of measures to augment the flow at Lake

Sevan). It is understood that the flow entering Kechut Reservoir from the

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tunnel is approximately 30 l/s. The water intake at Spandaryan has not

been opened since the tunnel was completed due to geotechnical

difficulties, and the water flowing from the tunnel outlet is assumed to be

groundwater.


Within the Darb catchment, to the north-west of Amulsar Mountain, is a

small pond, referred to as ‘Benik Pond’. The pond is approximately 1ha in

area, and is noteworthy for its naturally low pH and wetland biodiversity. It

is interconnected with local springs and surface water channels.

As identified previously, of the proposed main infrastructure associated with

the Project, only a portion of the open pit will be located within the Darb

catchment. The HLF, WD and most roadways will be situated within the Vorotan

catchment. The surface water quality, Conceptual Site Model, runoff and drainage

characteristics of the Project area are outlined in detail in the ESIA.

Lake Sevan is the largest lake in Armenia, and in the Caucasus Region. Its

basin makes up one sixth of the total territory of Armenia. The lake water is of

unusually high quality for a lake of its size and position. During the Soviet period,

flows were artificially increased from the lake, leading to dramatic falls in lake surface

area, and, among other impacts, a decline in biodiversity and water quality. The lake

remains an important national resource for water supply, electricity, fishing and

recreation. Measures to restore the quality and size of the lake have been ongoing

since the 1980’s, and included flow-augmentation tunnel interconnections with the

adjacent Arpa River basin. The, as yet, uncommissioned flow-augmentation tunnel

from the Vorotan River basin (Spandaryan-Kechut Tunnel) is one of these

interconnections.

Community Water Supplies

Drinking water for the village of Gorayk is supplied by three groundwater

springs originating in pasture land north east of the village. A new groundwater

source is envisaged to be developed by the village administration, as there are

existing water quality issues with the current supply. Drinking water is currently also

used for irrigation. Water for animals is sourced from groundwater filled storage

tanks on the outskirts of the village.

The Saravan village cluster is connected to the regional water mains for

drinking water. This water is shared with livestock. A multitude of groundwater

springs are used for irrigation water.

There is a reported pipeline and canal, fed by the headwaters of the River

Vorotan (upstream of the Project license area), which supplies the village of

Gndevaz with irrigation water. This essentially transfers a portion of the flow from the

Vorotan catchment, to the Arpa catchment.

Drinking water for Gndevaz is supplied through a water intake pipe which was

installed in the downstream section of an adjacent spring. This is further outlined in

the ESIA.

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20.4.5 Biodiversity

Protected Areas and Areas of International Significance

The Project is located below the southern edge of the Caucasus Mixed

Forests (CMF), which is a Global 200 Ecoregion (238 Ecoregions have been

identified by WWF as priority areas for global conservation because of their

important biodiversity). The Caucasus Mixed Forest Ecoregion covers a wide area

of 170,300 sq. km including portions of Georgia, Russia, Azerbaijan as well as

Armenia. It has been assigned a status of ‘critical/endangered' due to rapid land use

changes, including widespread deforestation. The Project site itself does not

currently support forests of the type prioritized within the Ecoregion.

The 2010 IUCN Red List of Threatened Species identifies around 50 species

of globally threatened animals in the Caucasus region as a whole.

The whole of Armenia is within a Birdlife International Endemic Bird Area 1 ,

which covers an area of 170,000 km 2 and includes portions of Azerbaijan, Georgia,

Iran, Russia and Turkey. The EBA is important for several restricted-range species

as well as breeding populations of raptors and reflects the importance of the

Caucasus as a center of bird endemism. Armenia provides important habitat for

many migratory bird species as part of an international flyway between Africa and

Europe, notably migratory raptors.

There are two Important Bird Areas (IBAs) in the vicinity of the Project:

Jermuk and Gorayk IBAs (see Figure 20.3). The IBAs constitute “Key Biodiversity

Areas” according to the definition in IFC Performance Standard 6 and have been

identified at national level using the globally standardized criteria which underpin the

KBA methodology 2 . The Concession Area partially overlaps with Gorayk IBA and the

proposed Heap Leach Pad location (Site 6) is located partially within it. The IBA was

designated primarily for its breeding colony of lesser kestrel Falco naumanni and the

boundary represents the limits of an assumed hunting area around the breeding

colony. The status of lesser kestrel on the IUCN Red List has decreased from

Vulnerable down to Least Concern though it is still listed as Vulnerable on the

Armenian Red List and the only breeding colony in the country is at Gorayk, making

it important in a national context. Gorayk IBA also was identified because of a large

number of other species including Egyptian Vulture (Neophron percnopterus) which

is listed as Endangered by IUCN as well as several other raptor species and a large

number of passerine and wetland birds.

The closest National Park to the site is the Sevan National Park located

approximately 44km to the north- north west of the Project. Three specially protected

Natural Areas (2 are only proposed and one is active) are located in the vicinity of

the Project as illustrated in Figure 1.3, below: Jermuk (proposed) is 2.9km north

1 BirdLife International (2012) Endemic Bird Area factsheet: Caucasus. Downloaded from

http://www.birdlife.org on 04/07/2012

2 To meet the KBA criteria, a site must contain: One or more globally threatened species; One or

more endemic species which are globally restricted to the site or surrounding region; Significant

concentrations of a species (e.g. important migratory stops, nesting sites, nurseries or breeding

areas); and/or Globally significant examples of unique habitat types and species assemblages.

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west of the WD, Herher Open Woodland (proposed) is 5.1km, west north west of the

WD and Jermuk Hydrological (operational) is 6.4km, north of the WD. WWF has put

forward proposals to develop an additional National Park encompassing these State

Reserves and the Jermuk Important Bird Area. These proposals are still under

discussion.

There are several wetland habitats present within the Project area, generally

within the Vorotan River, its valleys and tributaries form an extensive network of

surface drains within the Project area. Habitats include the Benik pond, River

Vorotan and tributaries, suspended marsh and mires wet meadow; some of which

form part of the functioning ecosystem supporting the species identified within the

IBA.

Protected Species

The Project affects four main vegetation types: alpine meadows, sub-alpine

meadows, mountain steppe and steppe grassland. There are also some wetlands.

Mount Amulsar is at an elevation of 2988m asl and has alpine vegetation. The mine

pit and crushing plant are located in areas above 2,100m with rocky outcrops, scree

slopes and alpine vegetation, including one plant which is included in the Armenian

Red Data Book. Other mine components are located on sub-alpine meadows and

steppe grassland. Vegetation has been analyzed using satellite imagery in the ESIA

and further details of the density of vegetation within each of the project components

have been considered.

Desk top information indicates several IUCN listed and Armenian Red Book

(1990) species which have a high potential occurrence in the Project area. The

Armenian Red Book was updated in 2010. Field studies confirmed the presence of 9

plant species which were listed in the 1990 Red Book but which were removed from

the 2010 version. In 2012, a Red Book (2010) plant species (potentilla

porphyrantha) was identified in the area of the open pit, and further baseline work is

being conducted to clarify the importance of the Amulsar population in a national

context.

The Project Area supports seven species of bird that are listed in the

Armenian Red Data Book and one which is listed as an endangered species in the

IUCN Red List. Further details of the ornithological studies are presented in the

ESIA.

Two species of dorcadion beetle (D. Bistriatum and D. Sevangense – the

latter of which is listed in the Armenian Red Book), the IUCN Vulnerable Apollo

butterfly (Parnassius apollo) and rock viper (Montivpera raddei – also included in the

Armenian Red Book) have been identified in the license area. No Red Listed fish

have been observed.

There are some other globally endangered species which are known to have

used the area in the past. These include the Caucasus leopard (an endangered subspecies)

and the Bezoar goat. Brown bear (included in the Armenian Red Data

Book) has been recorded in the vicinity of the Vorotan River near the proposed HLP

location. This is regarded as a keystone species in the region, though its populations

have declined dramatically and it is now likely to be a rare visitor. Together with

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Bezoar Goat It is one of the mammal species targeted for action in Armenia in the

recently issued Conservation Plan for the Caucasus Ecoregion (WWF 2012).

Moufflon and Bezoar goat are known to occur within the Jermuk area and NGO’s

have queried their presence in the project site, however they have not been

identified by field studies. The Caucasus leopard is likely to be locally extinct.

Evidence of the brown bear (Ursus arctos) has been found in the Project area.

Ecosystem Services

In terms of ecosystem services, foothill grasses and other species provide

seasonal grazing for sheep and cows in foothill zones where homeopathic species

are also known to occur. A questionnaire has been undertaken to assess which

areas are foraged by local people for plants and the types used for household/dietary

and medicinal purpose. The results provide information on the relative abundance

and local availability of the species used and have been used in the assessment of

impacts on ecosystem services and informal land uses.

Grasslands support seasonal grazing for sheep and cows; in the foothill zones

medicinal species are also known to be present. A questionnaire has been

undertaken to assess which areas are foraged by local people for wild plants and the

types used for household/dietary and medicinal/homeopathic purpose. The results

provide information on the relative abundance and local availability of the species

used and have been used in the assessment of impacts on ecosystem services and

informal land uses.

Habitat Designation

The license area is largely open in nature with a relatively high density of

surface water features.

Most land within the concession area is “natural” habitat according to the

definition in IFC PS6. The relatively large footprint of the Project means that

achieving No Net Loss of natural habitat is likely to require consideration of offset

activities.

IFC Performance Standard 6 (2012) sets out criteria for identifying areas

which might constitute “critical habitat”. The site as a whole could potentially

constitute critical habitat under criterion:

i) Because it provides habitat of significant importance to a globally

Endangered species (Egyptian vulture);

and possibly under criterion;

ii) Because it provides habitat for globally significant concentrations of

migratory species.

Additional criteria referred to in IFC PS6 Guidance Notes which could

potentially apply are: “Concentrations of Vulnerable (VU) species in cases where

there is uncertainty regarding the listing, and the actual status of the species may be

EN or CR” and “Habitat necessary for the survival of keystone species”.

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Additional baseline studies are being conducted to further refine the impact

assessment on biodiversity values at the Amulsar project. If critical habitat is

confirmed, biodiversity offsets will be developed to ensure no net loss occurs,

focusing on areas which could support the species and habitats which are being

impacted. One option for an appropriate offset may be to support the development

of the Jermuk National Park. Impacts to protected floral species will also be

minimized through project design where possible.

A Biodiversity Management Plan is being developed at present and will define

clear management and mitigation strategies for all biodiversity impacts and outline a

comprehensive monitoring plan. The results of continuing biodiversity studies will be

reported in the ESIA.

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Figure 20.3 - State Sanctuaries and Important Bird Areas in relation to the Project Exploration Licence

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20.4.6 Air Quality

There are no significant urban or industrial emission sources within the area, and

therefore the existing levels of related gasses (SO 2 , NO x , hydrocarbons, smoke

particulates, etc) are low to very low.

Baseline monitoring of particulates (total deposition) and gases (including SO 2

NO x , CO, CH 4 and volatile organic compounds) at the Project and nearest communities

will continue and will be supplemented by construction and operational monitoring in

order to assess and control (if necessary) emissions to air.

20.4.7 Noise and Vibration

There are no major urban centers or industrial activities in the region that would

result in significant levels of noise. The small hydro-electric power plant located close to

the proposed waste rock dump (WD) creates negligible noise impact in the Project area.

The M-2 public highway to the south of the Project experiences relatively

constant traffic conditions over a 24 hour period, with some seasonal variation, however

the baseline local traffic flows on the links in the study area are very low and within the

study area and local village residential receptors, traffic noise is generally considered

inaudible.

The baseline noise environment is typical rural, and experiences very low

background levels throughout the day and night.

Ground vibration and air overpressure results from blasting operations at mining

operations and construction projects. There are no operations of a similar nature in the

locality, and are absent in the current baseline conditions. The potential for effects

associated with blasting fall into two categories; those causing nuisance and those with

the potential for causing damage to structures. The principal source of vibration and

overpressure will result from blasting to remove rock from the open pit.

20.4.8 Visual and Landscape Aspects

The Project site is in a remote location, with small population in villages in the

local area. The landscape is characterized by steep and rolling topography dominated

by the Amulsar Mountain and the river valleys.

Landscape effects associated with a development relate to changes to the fabric,

character and quality of the landscape and how that is perceived by stakeholders who

have opinions on and/or will be potentially affected by the Project. This includes the

surrounding communities, seasonal herders and visitors to the area. Jermuk markets

itself as a spa town and which includes tourism and associated leisure activities such as

skiing. Tourism necessitates that the overall value and importance of the area is

considered from an aesthetic point of view.

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20.5 Environmental and Social Impact Assessment (ESIA)

A comprehensive ESIA has been prepared and for the purpose of summarizing

the potential impacts and approach to mitigation design and management the

environmental and social impact analysis has been considered separately in Tables

20.3 and 20.4. The importance magnitude, in terms of significance, taking account of

mitigation (i.e. the residual impact), has been defined as:

NEGLIGIBLE to MINOR:

MODERATE:

MAJOR:

Not significant.

Not significant subject to suitable management or

action plans, including potential for offsite

enhancement; otherwise

Significant.

20.5.1 Environmental Impact Assessment

20.3.

A summary of the environmental impacts from the ESIA are considered in Table

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Table 20.3

Summary of findings from the Environmental Impact Assessment

Potential Impact

PHASE

Source & Description

Inherent Impact Mitigation Residual Impact

Soil Quality Mountain meadow and pasture soils at elevation above 2000m.

At lower elevations, the soils are generally, brown, black earths or

alluvial in the Vorotan River valley. Soil quality will be affected by

earthworks carried out by the Project.

Moderate

Thin soil profile can be handled and stored using standard

mining operations. Use of relatively long, low storage

mounds


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Table 20.3

Summary of findings from the Environmental Impact Assessment

Potential Impact

PHASE

Source & Description

Inherent Impact Mitigation Residual Impact

Biodiversity

Habitats

The Project area has Alpine and Sub Alpine habitats as well as

extensive areas of montane steppe and steppe grasslands which

are managed through extensive summer grazing. The habitats are

“natural” according to the definition in IFC PS6. The grassland

communities are species rich and include some species of

conservation priority, particularly in the alpine communities.

Moderate

The Project has a large footprint of disturbance on natural

habitat with high species richness which is extensive in

Armenia but also declining rapidly in Armenia and is

threatened globally. There will be a residual impact

requiring measures through a biodiversity action plan to

achieve no net loss.

MODERATE

Flora

Fauna

The vegetation is generally important because of its species

richness, rather than because it supports individual species of

high conservation priority. The Project affects one plant species

Potentilla porphyrantha which is listed as Critically Endangered in

the Armenian Red Book.

The diversity of habitats, combined with the presence of abundant

water and low levels of disturbance and human modification have

contributed to a rich diversity of species. These include species

which are included in the IUCN Red List and the Armenian Red

Data Book. Animal species will be exposed to loss of habitat

through the Project’s physical footprint, increased levels of

disturbance and other indirect impacts associated with mining and

mineral processing.

Moderate/ Major

Moderate/ Major

A species action plan is being developed which is likely to

include measures to protect a proportion of the population

and efforts to translocate individuals located within the

Project footprint.

There is some potential for off-site mitigation to maintain,

and in some cases enhance biodiversity within the locality

taken as a whole.

MODERATE

MINOR/ MODERATE

Birds

The Project area is part of a global flyway and is particularly

important for migratory and breeding raptors, including Egyptian

Vulture which is Endangered on the IUCN Red List. Other species

are listed as Vulnerable and are included in the Armenian Red

List. Parts of the Project affect Gorayk IBA directly, which is

considered critical habitat. The valley of the River Vorotan outside

the IBA boundary provides important supporting habitat for many

of the listed species in the Gorayk IBA.

Major

The location of the HLF has been subject to assessment of

detailed design alternatives and the current proposed

location is considered to be the only feasible design.

Measures will be needed to offset the resulting loss of

hunting area for lesser kestrel (a primary reason for

designation of Gorayk IBA). Some species such as

corncrake will be difficult to provide mitigation for. Species

listed in the IBA which will be affected by the Project will be

included in a Biodiversity Action Plan supported by

monitoring.

MODERATE

Air Quality Dust Emissions of dust result from surface mining activities, including

overburden / rock and ore extraction, haulage and processing. In

addition, fugitive emissions of dust result from areas of bare and

disturbed ground in dry windy conditions. Vulnerable receptors

include: flora and fauna, surface water, and human communities

in terms of nuisance and potential health impacts. A very high

proportion of fugitive dusts will settle within a short distance from

the emission point or source. The mine activities are remote from

human settlements and the land close to the mine is not subject to

intensive agricultural usage.

Negligible to Moderate

Dust management plan to prevent release of particulates

into the area surrounding the site. Long term monitoring will

enable analysis of the effectiveness of the management

plan and provides feedback to manage operational

procedures and seek alternatives to minimize emission of

particulates over the life of the mine

NEGLIGIBLE

Gaseous emissions

Acid gases from vehicle exhaust fumes related to construction,

extraction, loading and haulage operations. Vulnerable receptors

include human receptors. However, these activities do not take

place within a close proximity to settlements.

Negligible

Regular maintenance and scheduling of all vehicles used at

the mine. Site speed restrictions on haul roads to reduce

optimize vehicle use and fuel efficiency.

NEGLIGIBLE

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Table 20.3

Summary of findings from the Environmental Impact Assessment

Potential Impact

PHASE

Source & Description

Inherent Impact Mitigation Residual Impact

Greenhouse gases

Greenhouse gas emissions derive from onsite diesel use from

mobile plan, oil and gas for heat. Offsite source derives from

electrical usage for static plant, conveyors, lighting and general

industrial and business uses.

The local environment is vulnerable to climate change, in relation

to natural habitats, duration of snow cover and change in mean

annual, mean and maximum daily temperatures.

Minor

Of the fuel sources, operational procedures should be

directed towards the efficient use of diesel and heating oil

that result in lower GHG emissions/kW.

Electrical supply in Armenia is principally sourced from

nuclear and gas, with future projects to develop both

geothermal and wind that should reduce the embedded

GHG emissions associated with grid electrical supply.

MINOR

Noise

Emissions associated with operational plant within the mine.

Vulnerable receptors include dwellings within settlements (closest

is 5km distance), herders and summer grazing flocks, and

disturbance to off-site wildlife.

Negligible to moderate

Site noise levels will be maintained within an acceptable

range of between 32 to 37 LA eq dBL. At these levels

operations will be inaudible at nearest settlements and

have a negligible effect on potential disturbance of summer

grazing and herding.

NEGLIGIBLE

Air Overpressure

Blasting operations for extraction will result in instantaneous noise

emission termed air over pressure. Due the remote location of the

mine, there are limited human receptors; however there is the

potential for disturbance of fauna and summer grazing herds.

Minor

Blasting design and practice can be used to mitigate this

impacts and actual emission can be controlled to a level

between 94dBL and 109dBL.

MINOR

Vibration

Ground vibration that result from the blasting operations. Due the

remote location of the mine, there are limited human receptors;

however there is the potential for disturbance of fauna and

summer grazing herds.

Negligible

Blast pattern designed to prevent excessive vibration

and negate any potential.

NEGLIGIBLE

Landscape &

visual

Visual intrusion of activities

Excavation of mining void and ancillary infrastructure – exposed

and elevated ridge with views from surrounding areas with

potential views towards the open pit from Saralanj to south.

Above ground structures and mounds, these are dispersed within

the development area and tend to be visually confined, by

surrounding relief. Structures and waste rock dump will be visually

prominent close to, but not from visually important receptors – due

to topography and distance separation.

Negligible to

Moderate, depending on

viewpoint

The footprint of direct disturbance has been defined and

will be maintained by delineating the outer boundary with a

perimeter soil mound. Shape and seed the outer face of the

mound.

Appropriate maintenance, as identified in ESAP directed to

best practice in terms of the appearance of operational

areas.

NEGLIGIBLE TO MINOR

NEGLIGIBLE

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20.5.2 Summary of Environmental Impacts

By adopting a wide range of impact management and mitigation measures, it is

considered that any potential residual environmental impacts can be reduced to a

moderate or low (or below) level. Mitigation measures include reduction of emissions

by active dust control; biodiversity offset measures, blast limitation, water management

and treatment, monitoring programs with stakeholders consultation and participation,

good waste management practices, and progressive rehabilitation. Full mitigation

measures are defined in the ESIA, which has been subject to an iterative process in

parallel with mine design and operational practice. Therefore, mitigation measures that

formed part of the design have been included in the development costs. In addition,

there are a range of associated management plans which will be incorporated into the

design and operations, through procedures, as well as Lydian’s ESMS. Early adoption

of management plans allows for efficiency of design and therefore the potential for cost

savings and these have been considered during the iterative process of site design.

20.6 Social Context and Baseline

20.6.1 Archaeology and Cultural Heritage

A phased assessment has been undertaken by in-country experts. The initial

desk study phase for the Project area (5000ha) did not indicate the presence of any

sites or features of cultural or archaeological significance within the license. However,

subsequent reconnaissance surveys revealed the presence of several features of

potential archaeological interest, such as tombs and kurghans (graves). Most features

are located away from areas of proposed maximum disturbance and some appear to be

already disturbed and degraded.

Currently none of the features identified are thought to necessitate in-situ

preservation which could affect the progression of development. All features are being

recorded by a State archaeologist; working in concert with international archaeologists.

Appropriate measures will be identified for off- or on-site preservation, as appropriate.

20.6.2 Demographic, Land-Use, Family Structure and Migration Patterns

The study area for socio-economic considerations is comprised of the villages of

Gorayk, Saravan (including Saralanj and Ughedzor), and Gndevaz, all of which lie

within a 9km radius around the Project, as well as the city of Jermuk (and the

associated village of Kechut), located 14km from the Project. Socio-economic baseline

data were obtained through reconnaissance visits, a household survey covering all rural

households and a sample of Jermuk households, a number of focus groups with

community members, as well as semi-structured interviews with a range of community

members, community leaders and administrators.

The rural settlements within the study area have a naturally aesthetic setting,