lydian international, ltd. · 2015. 6. 15. · 1. i am an independent mineral process engineering...

KDE FORM No. A263a-7/12/99

LYDIAN INTERNATIONAL, LTD.

Development of Amulsar Heap Leach Facility

Preliminary Economic Assessment

PREPARED FOR:

LYDIAN INTERNATIONAL, LTD. 1st Floor, Capstan House

La Route es Nouaux St. Helier, Jersey JE2-4ZJ

Channel Islands

PREPARED BY:

Mr. Joseph M. Keane, P.E. Mr. Richard Kiel, P.E.

Mr. Galen White, BSc (Hons), FGS, MAusIMM Mr. Kent Bannister, MAusIMM CP

Mr. John Eyre, FRICS MIMMM MIQ CEnv

K D Engineering 7701 N. Business Park Drive

Tucson, Arizona 85743

Document No. Q439-03-028-01 Project No. 439-01

12 August 2011

KD Engineering

7701 N. Business Park Drive Telephone: (520) 579-8315Tucson, AZ 85743 Facsimile: (520) 579-3686

E-Mail: jkeanekdengco.com

CERTIFICATE of AUTHOR

I, Joseph M. Keane, P.E. do hereby certify that:

1. I am an Independent Mineral Process Engineering Consultant and contributedto a Report entitled “Lydian International, Ltd., Development of Amulsar HeapLeach Facility - Preliminary Economic Assessment”, dated 12 August 2011 asan associate of the following organization:

K D Engineering7701 N. Business Park DriveTucson, Arizona 85743Telephone: 520-579-8315Fax: 520-579-3686E-Mail: [email protected]

2. This certificate applies to the Report titled “Lydian International, Ltd.,Development of Amulsar Heap Leach Facility - Preliminary EconomicAssessment”, dated 12 August 2011 (the “Prefeasibility Study”).

3. I graduated with a degree of Bachelor of Science in Metallurgical Engineeringfrom the Montana School of Mines in 1962. I obtained a Master of Science inMineral Processing Engineering in 1966 from the Montana College of MineralScience and Technology. In 1989 I received a Distinguished Alumni Awardfrom that institution.

4. I am a member of the Society for Mining, Metallurgy, and Exploration, Inc.(SME #1682600) and the Instituto de Ingenieros de Minas de Chile. I am aregistered professional metallurgical engineer in Arizona (#1 2979) and Nevada#5462).

5. I have worked as a metallurgical engineer for a total of 49 years since mygraduation from university.

6. I have read the definition of “qualified person” set out in National Instrument43-101 (“NI 43-101”) and certify that by reason of my education, affiliation witha professional association (as defined in NI 43-101) and past relevant workexperience, I fulfill the requirements to be a “qualified person” for the purposesof NI 43-101.

KD Engineering Co., Inc.

7701 N. Business Park DriveTucson, AZ 85743

Telephone: (520) 579-8315Facsimile: (520) 579-3686

E-Mail: [email protected]

7. I am responsible for Sections 1 through 13, 17, 18 19, the process portion of21, 22 through 24, 27 and 28 of the above referenced report and the overallassembly of the report. I visited the property during the period 21 through 28May2011.

8. I have not had prior involvement with the property that is the subject of theTechnical Report.

9. I am independent of the issuer applying all of the tests in Section 1.5 ofNational Instrument 43-101.

10. As of the date of this certificate, to the best of my knowledge, information, andbelief, the prefeasibility study contains all scientific and technical informationthat is required to be disclosed to make the prefeasibility study not misleading.

11. I have read National Instrument 43-101 and Form 43-IOIFI, and thePrefeasibility Study has been prepared in compliance with that instrument andform.

12. I consent to the filing of the Preliminary Economic Assessment with any stockexchange and other regulatory authority and any publication by them, includingelectronic publication in the public company files on their websites assessableby the public.

Dated this 12th D. of August 2011

Joserh M. KeanePrint Name of Qualified Person

Signature of Qualified Person

KD Engineering Co., Inc.

KD Engineering

7701 N. Business Park Drive Telephone: (520) 579-8315Tucson, AZ 85743 Facsimile: (520) 579-3686

August 12, 2011

To: Lydian International, Ltd.British Columbia Securities CommissionAlberta Securities CommissionAutorité des marches financiersTSX Venture Exchange

Re: Consent to Filing of Technical Report

I, Joseph M. Keane, P.E., am a “qualified person” as defined in National Instrument 43-101 (NI 43-101) and prepared or supervised the preparation of all or part of thetechnical report entitled “Lydian International, Ltd., Development of Amulsar HeapLeach Facility - Preliminary Economic Assessment” dated 12 August 2011 (the“Technical Report”).

I hereby consent to the public filing of the Technical Report. I also consent to anyextracts from, or a summary of, the Technical Report in the press release dated July25, 2011 (the “Press Release”) of Lydian International.

I certify that I have read the Press Release and that it fairly and accurately representsthe information in the sections of the Technical Report for which I am responsible.

Dated this 12th day of August 2011.

Joseph M. Keane, P.E.

11381597FS LTR CertOfAuthorAmulsarPEA 12AUG11.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

August 12, 2011 113-81597FS.110 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 (Preliminary Economic Assessment) 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 [email protected].

2. I hold the following academic qualifications:

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, Wyoming, and

Kansas.

5. I have worked as a civil and geological engineer in the minerals industry for 20 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 18 years as a consulting engineer on precious and base metals, and 2 years as a geologist and engineer on an operating uranium mine. I am qualified to prepare and review the engineering for the heap leach facility and for geotechnical engineering aspects of the Amulsar project.

7. I visited the property in June 2011.

8. This is the first 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.

10. I am responsible for the preparation of the Technical Report for the Heap Leach Pad and Ponds as

discussed in Sections 1, 17, and 18 and for the geotechnical portions of Section 21 of the PEA. Sincerely, GOLDER ASSOCIATES INC. Richard E. Kiel Senior Geological Engineer cc: Brad Schwab, KDE

mailto:[email protected]�

I:\11\81597FS\0100\0110 LT\AUG11\11381597FS LTR ConsentToFileTechRpt 12AUG11.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

August 12, 2011 113-81597FS.110

Ms. Kim Batorson KD Engineering & Metcon Research 7701 North Business Park Drive Tucson, Arizona 85743-9622

RE: CONSENT TO FILING OF TECHNICAL REPORT

Dear Ms. Batorson:

I, Richard E. Kiel, am a “qualified person” as defined in National Instrument 43-101 (NI 43-101) and prepared or supervised the preparation of all or part of the technical report entitled “Lydian International, Ltd., Development of Amulsar Heap Leach Facility – Preliminary Economic Assessment" dated August 12, 2011 (the “Technical Report”).

I hereby consent to the public filing of the Technical Report. I also consent to any extracts from, or a summary of, the Technical Report in the press release dated July 25, 2011 (the “Press Release”) of Lydian International.

I certify that I have read the Press Release and that it fairly and accurately represents the information in the sections of the Technical Report for which I am responsible.

Dated this 12th day of August 2011.

Sincerely,

GOLDER ASSOCIATES INC.

Richard E. Kiel Senior Geological Engineer

REK/rjg

CERTIFICATE of AUTHOR

As author and QP responsible for a recent Technical Report titled “Mineral Resource Estimate, Lydian International Limited, Amulsar Gold Project. 43-101 Technical Report, Armenia” dated 19 May 2011, and filed on SEDAR, extracts of which are presented by Lydian International Ltd, Report – Development of Amulsar Heap Leach Facility. Preliminary Economic Assessment, prepared by K.D Engineering, 7701 N. Business Park Drive, Tuscon, Arizona 85743, Document No Q439-03-028-01, Project No. 439-01 July 2011 I, Galen White do hereby certify that:

1. I am an Principal Geologist of CSA Global (UK) Ltd, and carried out this assignment for, CSA Global (UK) Ltd, 2 Peel House, Barttelot Road, Horsham, West Sussex, RH12 1DE, UK Telephone +44 1403 255 969, e-mail www.csaglobal.com.

2. This certificate applies to the Report titled "Lydian International, Ltd., Development of Amulsar Heap Leach Facility - Preliminary Economic Assessment", dated 12 August 2011 (the " Prefeasibility Technical Report") prepared by K.D Engineering, 7701 N. Business Park Drive, Tuscon, Arizona 85743, Document No Q439-03-028-01, Project No. 439-01 July 2011.

3. I graduated with a degree of Bachelor of Science, with honours in Geology

from the University of Portsmouth, UK in 1996. 4. I am a registered Member of the Australian Institute of Mining & Metallurgy

(Membership # 226041) and a fellow of the Geological Society of London (# 1003505).

5. I have worked as a geologist for a total of 15 years since my graduation from

university. 6. I have read the definition of "qualified person" set out in National Instrument

43-101 ("NI 43-101") and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purposes of NI 43-101.

http://www.csaglobal.com/

7. I am responsible for a Technical Report titled “Mineral Resource Estimate, Lydian International Limited, Amulsar Gold Project. 43-101 Technical Report, Armenia” dated 19 May 2011, and filed on SEDAR, extracts of which are presented by Lydian International Ltd in this Technical Report.

8. I have had prior involvement with the property that is the subject of the

Technical Report, having co-authored a Technical Report on the project in 2009 and 2010, both filed on SEDAR.

9. I am independent of the issuer applying all of the tests in Section 1.5 of

National Instrument 43-101. 10. As of the date of this certificate, to the best of my knowledge, information, and

belief, the content of this Technical Report extracted from the earlier Technical Report contains all scientific and technical information that is required to be disclosed to make the Report not misleading.

11. I have read National Instrument 43-101 and Form 43-101F1, and the

Prefeasibility Study has been prepared in compliance with that instrument and form.

12. I consent to the filing of the Preliminary Economic Assessment with any stock

exchange and other regulatory authority and any publication by them, including electronic publication in the public company files on their websites assessable by the public.

Dated this 12th Day of August 2011 Galen White BSc(Hons) MAusIMM FGS

August 12, 2011 To: Lydian International, Ltd.

British Columbia Securities Commission Alberta Securities Commission Autorité des marchés financiers TSX Venture Exchange


I, Galen White., am a “qualified person” as defined in National Instrument 43-101 (NI 43-101) and I am responsible for a recent Technical Report titled “Mineral Resource Estimate, Lydian International Limited, Amulsar Gold Project. 43-101 Technical Report, Armenia” dated 19 May 2011, and filed on SEDAR, extracts of which are presented by Lydian International Ltd, Report – Development of Amulsar Heap Leach Facility. Preliminary Economic Assessment, prepared by K.D Engineering, 7701 N. Business Park Drive, Tuscon, Arizona 85743, Document No Q439-03-028-01, Project No. 439-01 July 2011 (the “Technical Report”). I hereby consent to the public filing of the Technical Report. I also consent to any extracts from, or a summary of, the Technical Report in the press release dated July 25, 2011 (the “Press Release”) of Lydian International. I certify that I have read the Press Release and that it fairly and accurately represents the information in the sections of the Technical Report for which I am responsible. Dated this 12th day of August, 2011. ____________________________ Galen White BSc(Hons), MAusIMM, FGS.

Certificate of Author_KRB LR.docx Page 1 of 1

CERTIFICATE OF AUTHOR

KENT R BANNISTER

MAusIMM (CP)

As the co-author of this Technical Report on the Amulsar Gold Project of Lydian International Ltd., in Armenia, I, Kent

R. Bannister do hereby certify that:

1) I am a Director of CSA Global Pty Ltd, and carried out this assignment for, CSA Global Pty Ltd, Level 2, 3

Ord Street, West Perth, WA 6872 Telephone: +61893551677, www.csaglobal.com.

2) I hold the following academic qualifications: Associate Diploma Mining Engineering, from the Royal

Melbourne Institute of Technology, Melbourne, Australia.

3) I am a registered Member and Certified Practicing (CP) Mining Engineer of the Australasian Institute of

Mining & Metallurgy (Membership # 102511).

4) I have worked as a Mining Engineer in the minerals industry for 37 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. After graduating in 1974 and mining for

two years in Kalgoorlie and Mt Isa, I have gained over 37 years of experience in base metals, gold,

uranium, manganese, mineral sands, nickel and iron ore. During this time, I have held positions as

Mining Engineer, Senior Mining Engineer, Assistant Superintendent, Principal Engineer, Mine Manager,

Registered Manager, General Manager, Consultant, Technical Advisor and Company Director.

6) I have not visited the property.

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

8) 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.

9) I am independent as defined by NI 43-101 regulations and have provided consulting services to Lydian

International Ltd.

10) I am responsible for the preparation of the CSA Global Scoping Study Report No: R268.2011 involving the

mining engineering and open cut design and associated economics presented by Lydian International, Ltd

Report, Development of Amulsar Heap Leach Facility, Preliminary Economic Assessment, prepared by K

D Engineering, 7701 N. Business Park Drive, Tucson, Arizona 85743, Document No. Q439-03-028-01,

Project No. 439-01, July 2011.

Dated this 12th day of August, 2011

Kent R Bannister. MAusIMM (CP) Director Mining and Projects

http://www.csaglobal.com/

Consent letter Page 1 of 1

August 12th, 2011

To: Lydian International, Ltd. British Columbia Securities Commission Alberta Securities Commission Autorité des marchés financiers TSX Venture Exchange


6) I, Kent Bannister, am a “qualified person” as defined in National Instrument 43-101 (NI 43-101) and prepared or supervised the preparation of all or part of the technical report entitled “CSA Global Scoping Study Report No: R268.2011 involving the mining engineering and open cut design and associated economics presented by Lydian International, Ltd., Development of Amulsar Heap Leach Facility - Preliminary Economic Assessment" dated 12 August 2011 (the “Technical Report”).

I hereby consent to the public filing of the Technical Report. I also consent to any extracts from, or a

summary of, the Technical Report in the press release dated July 25, 2011 (the “Press Release”) of Lydian

International.

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

sections of the Technical Report for which I am responsible.

Dated this 12th day of August, 2011

____________________________

Kent Bannister MAusIMM CP

Wardell Armstrong InternationalWheal Jane, Baldhu, Truro, Cornwall, TR3 6EH, United KingdomTelephone: +44 (0)1872 560738 Fax: +44 (0)1872 561079 www.wardell-armstrong.com

Wardell Armstrong International is the trading name of Wardell Armstrong International Limited,Registered in England No. 3813172

Registered office: Sir Henry Doulton House, Forge Lane, Etruria, Stoke-on-Trent, ST1 5BD, United Kingdom

UK Offices: Stoke-on-Trent, Cardiff, Edinburgh, Greater Manchester, Liverpool, London,Newcastle upon Tyne, Sheffield, Truro, West Bromwich. International Offices: Almaty, Beijing

ENERGY AND CLIMATE CHANGE

ENVIRONMENT AND SUSTAINABILITY

INFRASTRUCTURE AND UTILITIES

LAND AND PROPERTY

MINING, QUARRYING AND MINERAL ESTATES

WASTE RESOURCE MANAGEMENT

CERTIFICATE OF AUTHORJOHN M. EYRE

As the co-author of this Preliminary Economic Assessment on the Amulsar Mine Project of LydianInternational Limited, in Southern Armenia, I, John M. Eyre do hereby certify that:

1) I am an Associate, and carried out this assignment for, Wardell Armstrong International Limited,Wheal Jane, Baldhu, Truro, Cornwall, United Kingdom TR3 6EH, tel. +44 (0)1872 560738, [email protected]

2) I hold the following academic qualifications:

RICS (Direct Entry Examinations) Minerals Surveying North Staffordshire Polytechnic, UK1975-1978

3) I am a registered Fellow of the Royal Institution of Chartered Surveyors (Minerals and Environment)Membership No. 00058203, a Member of the Institute of Mining, Materials & Metallurgy, aMember of the Institute of Quarrying, an Associate Member of the Institute of EnvironmentalManagement and Assessment and a Chartered Environmentalist;

4) I have worked as a minerals surveyor and resource manager in the minerals industry for 39 years;

5) I am familiar with NI 43-101 and, by reason of education, experience and professional registration; Ifulfill the requirements of a Qualified Person as defined in NI 43-101. My work experience includes14 years in operations working at underground and surface mining operations, 16 years as a SeniorLecturer at the Camborne School of Mines, University of Exeter, 7 years as a consulting mining-environmental director on precious and base metals , energy and industrial minerals. I am qualifiedto review and comment on the environmental and social matters relating to the Amulsar Project;

6) I last visited the property in June, 2011;

7) This is the first Preliminary Economic Assessment report I have co-authored on the mineralproperty in question;

8) I co-authored an Environmental and Social Scoping Report for the mineral property in question, inFebruary 2011;

9) As of the date of this certificate to the best of my knowledge, information and belief, the TechnicalReport contains all scientific and technical information that is required to be disclosed to make thisreport not misleading;

10) I am independent of Lydian International Limited as defined by NI 43-101 regulations and haveprovided consulting services to Lydian;

2

11) I am responsible for the preparation of section 20, and jointly wrote the summary section of theTechnical Report dated July 29, 2011 entitled “Lydian International, Ltd. Development of AmulsarHeap Leach Facility Preliminary Economic Assessment”.

Dated this 29 day of July, 2011

John Maxwell Eyre FRICS MIMMM MIQ CEnvDirector,North Coast Consulting LimitedAssociate of Wardell Armstrong International Limited

Wardell Armstrong InternationalWheal Jane, Baldhu, Truro, Cornwall, TR3 6EH, United KingdomTelephone: +44 (0)1872 560738 Fax: +44 (0)1872 561079 www.wardell-armstrong.com

Wardell Armstrong International is the trading name of Wardell Armstrong International Limited,Registered in England No. 3813172

Registered office: Sir Henry Doulton House, Forge Lane, Etruria, Stoke-on-Trent, ST1 5BD, United Kingdom

UK Offices: Stoke-on-Trent, Cardiff, Edinburgh, Greater Manchester, Liverpool, London,Newcastle upon Tyne, Sheffield, Truro, West Bromwich. International Offices: Almaty, Beijing

ENERGY AND CLIMATE CHANGE

ENVIRONMENT AND SUSTAINABILITY

INFRASTRUCTURE AND UTILITIES

LAND AND PROPERTY

MINING, QUARRYING AND MINERAL ESTATES

WASTE RESOURCE MANAGEMENT

August 12, 2011

To: Lydian International, Ltd.British Columbia Securities CommissionAlberta Securities CommissionAutorité des marchés financiersTSX Venture Exchange


I, John M Eyre, FRICS. CEnv am a “qualified person” as defined in National Instrument 43-101 (NI43-101) and prepared or supervised the preparation of all or part of the technical report entitled“Lydian International, Ltd., Development of Amulsar Heap Leach Facility - Preliminary EconomicAssessment" dated 12 August 2011 (the “Technical Report”).

I hereby consent to the public filing of the Technical Report. I also consent to any extracts from, ora summary of, the Technical Report in the press release dated July 25, 2011 (the “Press Release”)of Lydian International.

I certify that I have read the Press Release and that it fairly and accurately represents theinformation in the sections of the Technical Report for which I am responsible.

Dated this 12th day of August, 2011.

____________________________John M Eyre FRICS CEnv

Lydian International - Amulsar Heap Leach Facility Preliminary Economic Assessment

K D Engineering Document No. Q439-03-028-01 12 August 2011 KDE FORM No. A263a-7/12/99

TABLE OF CONTENTS

Section 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 ....................... 23 6.0 HISTORY ............................................................................................... 26 7.0 GEOLOGICAL SETTING .............................................................................. 28 8.0 DEPOSIT TYPES AND MINERALISATION ................................................. 37 9.0 EXPLORATION ............................................................................................. 39 10.0 DRILLING ............................................................................................... 40 11.0 SAMPLE PREPARATION, ANALYSIS AND SECURITY ............................. 54 12.0 DATA VERIFICATION .................................................................................. 56 13.0 MINERAL PROCESSING AND METALLURGICAL TESTING .................... 78 14.0 MINERAL RESOURCE ESTIMATES ........................................................... 93 15.0 MINERAL RESERVE ESTIMATES ............................................................... 95 16.0 MINING METHODS ....................................................................................... 103 17.0 RECOVERY METHODS ............................................................................... 114 18.0 INFRASTRUCTURE ..................................................................................... 123 19.0 MARKET STUDIES AND CONTRACTS ...................................................... 131



TABLE OF CONTENTS (Continued)

Section Page 20.0 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT ..................................... 132 21.0 CAPITAL AND OPERATING COSTS ........................................................... 152 22.0 ECONOMIC ANALYSIS ................................................................................ 166 23.0 ADJACENT PROPERTIES ........................................................................... 187 24.0 OTHER RELEVANT DATA AND INFORMATION ........................................ 188 25.0 INTERPRETATION AND CONCLUSIONS ................................................... 189 26.0 RECOMMENDATIONS ................................................................................. 192 27.0 REFERENCES .............................................................................................. 196 28.0 APPENDICES ............................................................................................... 197 Appendix 1 - Design Criteria Appendix 2 - Equipment List Appendix 3 - Capital Cost Detail Appendix 4 - Drawings



List of Tables

Table 1.1 Testwork Results Summary ..................................................................................... 5 Table 1.2 Coarse Bottle Roll Cyanidation Leach Test Summary .......................................... 6 Table 1.3 Total Initial and Future Sustaining Project Costs .................................................. 12 Table 1.4 Cash Operating Cost ................................................................................................ 12 Table 1.5 Preliminary Economic Highlights (Base Case) ...................................................... 13 Table 1.6 Economic Evaluation ................................................................................................ 14 Table 1.7 Summary of Key Financial Parameters (Sensitivity to Gold Price) ..................... 14 Table 9.1 Summary of Main 2010 Exploration Activities on the Amulsar Gold Project ..... 39 Table 10.1 Selected Down Hole Gold Intercepts from 2010 Amulsar Drilling Program ....... 42 Table 10.2 Selected Down Hole Gold Intercepts from 2010 Amulsar Drilling Program ....... 42 Table 10.3 RC and DD Twinned Holes ....................................................................................... 47 Table 10.4 SG Values According to Rock Type ........................................................................ 49 Table 12.1 2010 Assay QA/QC Summary .................................................................................. 58 Table 13.1 Screened Metallics Analysis for Gold, Comp 1 ..................................................... 78 Table 13.2 Head Analyses of Composite 1 ............................................................................... 79 Table 13.3 Coarse Bottle Roll Leach Test Summary ............................................................... 80 Table 13.4 Testwork Results Summary ..................................................................................... 81 Table 13.5 Minus 75 µm Bottle Roll Leach Tests ..................................................................... 82 Table 13.6 Minus 2 mm Bottle Roll Leach Tests ...................................................................... 82 Table 13.7 Minus 38 mm Column Leach Tests ......................................................................... 83 Table 13.8 Minus 19 mm Column Leach Tests ......................................................................... 84 Table 13.9 Final Gold Recovery Summary by Test and Composite ....................................... 85 Table 13.10 Column Leach Test Results Summary ................................................................... 85 Table 13.11 Coarse Ore Bottle Roll Leach Test Results (Gold) ................................................ 87 Table 13.12 Coarse Ore Bottle Roll Leach Test Results (Silver) .............................................. 87 Table 13.13 Heap Leach Design Parameters .............................................................................. 90 Table 14.1 Mineral Resource Estimate at a 0.4 g/t Gold Cut-Off Grade ................................. 94 Table 15.1 Amulsar Gold Deposit – Au Resources Input to Whittle ....................................... 96 Table 15.2 Amulsar Au Deposit Whittle Input Parameters ...................................................... 98 Table 15.3 Whittle Optimization Results Optimal Pit Shell # with Physical Results for each Optimization Option Amulsar Au Deposit ... 99 Table 15.4 Whittle Optimization Results Optimal Pit Shell # with Financial Results for each Optimization Option Amulsar Au Deposit .. 99 Table 15.5 Conversion of Resources to PMM (In-situ Figures) .............................................. 99 Table 15.6 Design vs. Whittle Shell Comparison ..................................................................... 102 Table 16.1 Optimal Shells for Pit Design ................................................................................... 103 Table 16.2 Pit Design Parameters .............................................................................................. 103 Table 16.3 Amulsar Pit Design PMM Tonnes and Grade Inventory ........................................ 107 Table 16.4 Amulsar Waste Dump Volume Requirements ........................................................ 108 Table 16.5 Amulsar Waste Dump Design Parameters ............................................................. 108 Table 16.6 Amulsar Waste Dump Design Volumes .................................................................. 109 Table 16.7 Amulsar Production Schedule ................................................................................. 113 Table 20.1 Average Precipitation (mm) ..................................................................................... 132 Table 20.2 Humidity ..................................................................................................................... 133 Table 20.3 Water Monitoring Schedule ..................................................................................... 140



List of Tables (Continued)

Table 20.4 Water Quality Monitoring Parameters..................................................................... 140 Table 20.5 Income Distribution in the Study Area .................................................................... 145 Table 20.6 Priority Needs in the Study Area ............................................................................. 149 Table 21.1 Capital Costs Summary ............................................................................................ 152 Table 21.2 Operating Costs Summary ....................................................................................... 153 Table 21.3 Summary Process Plant Initial Capital Cost .......................................................... 154 Table 21.4 Summary Process Plant Sustaining Capital Cost ................................................. 155 Table 21.5 Process Plant Operating Cost Estimate – Summary ............................................. 159 Table 21.6 Operating Cost Estimate - Heap Leach Power & Energy ...................................... 159 Table 21.7 Operating Cost Estimate - Plant Labor (Phase I) ................................................... 160 Table 21.8 Operating Cost Estimate - Plant Labor (Phase II ................................................... 161 Table 21.9 Operating Cost Estimate - Yerevan Office Administration Labor ........................ 162 Table 21.10 Operating Cost Estimate - Site Administration Labor .......................................... 163 Table 21.11 Operating Cost Estimate - Heap Leach Consumables .......................................... 164 Table 21.12 Maintenance .............................................................................................................. 165 Table 21.13 Water .......................................................................................................................... 165 Table 22.1 Economic Analysis Summary .................................................................................. 166 Table 22.2 Economic Analysis Summary - US$ Pre-Income Tax Cash Flow ........................ 167 Table 22.3 Cash Flow Schedule (Owner Operating Case) ....................................................... 168 Table 22.4 Rate of Return Sensitivity ........................................................................................ 170 Table 22.5 NPV Sensitivity (US$ X 1000) ................................................................................... 171 Table 22.6 Economic Analysis Summary .................................................................................. 172 Table 22.7 Economic Analysis Summary - US$ Pre-Income Tax Cash Flow ........................ 172 Table 22.8 Cash Flow Schedule (Contract Mining) .................................................................. 173 Table 22.9 Rate of Return Sensitivity, Percent ......................................................................... 175 Table 22.10 NPV Sensitivity (US$ X 1000) ................................................................................... 176 Table 22.11 Economic Analysis Summary .................................................................................. 177 Table 22.12 Economic Analysis Summary - US$ Pre-Income Tax Cash Flow ........................ 177 Table 22.13 Cash Flow Schedule (Owner Operating with Erato) .............................................. 178 Table 22.14 Rate of Return Sensitivity, Percent ......................................................................... 180 Table 22.15 NPV Sensitivity (US$ X 1000) ................................................................................... 181 Table 22.16 Economic Analysis Summary .................................................................................. 182 Table 22.17 Economic Analysis Summary - US$ Pre-Income Tax Cash Flow ........................ 182 Table 22.18 Cash Flow Schedule (Contract Mining with Erato) ................................................ 183 Table 22.19 Rate of Return Sensitivity, Percent ......................................................................... 185 Table 22.20 NPV Sensitivity (US$ X 1000) ................................................................................... 186



List of Figures

Figure 4.1 Location of Amulsar Gold Project ........................................................................... 20 Figure 5.1 Physiography and Infrastructure in the Vicinity of the Amulsar Licenses .......... 24 Figure 7.1 Distribution of prospective Eocene-Oligocene arc-magmatic rocks and major known mineral occurrences draped on SRTM DEM. .................. 29 Figure 7.2 (L) - Siliciclastic marine sediments found to the north-west of the Saravan license area. (R) - Carbonate vein-style base metal mineralization ............... 30 Figure 7.3 A. Basaltic sill within volcano sedimentary units. B & C. Examples of andesitic-basaltic volcano sedimentary breccia units ................................. 31 Figure 7.4 Geological map of the Amulsar area showing locations of sections................... 32 Figure 7.5 Cross section showing geology from Erato to Tigranes ....................................... 32 Figure 7.6 Cross section showing geology from Artavasdes to the eastern and northern slopes of Tigranes ............................................................................................. 33 Figure 7.7 A - DDA-067, early ‘grey’ silica massive and porphyritic feldspar-augite andesite clasts within a second phase hydrothermal breccia with limonitic silica-clay cement. B – DDA-067, early ‘grey’ silica massive clasts within a second phase hydrothermal breccias ............................................................ 34 Figure 7.8 (R) - Oxidized massive sulphide cemented hydrothermal breccia with clasts of early ‘grey’ massive silica altered volcanics. (L) – Limonitic, silica altered flow-banded volcanics. ...................................... 35 Figure 8.1 Typical alteration patterns of the Borealis District Gold Deposit ......................... 38 Figure 10.1 Left Hole DDA-007 Showing Porous Limonite Breccia Zone ................................ 41 Figure 10.2 Right Hole DDA-07 Silicified Breccia Zone with Partial Limonite Matrix ............. 41 Figure 10.3 Map of Planned Drilling for 2011 at Amulsar .......................................................... 44 Figure 10.4 Channel cutting on site at Amulsar ......................................................................... 45 Figure 10.5 Channel sampling on site at Amulsar ..................................................................... 45 Figure 10.6 Map of Amulsar Area Showing Location of trenches in green and channels in red Drill hole collars are represented by small black dots .......................... 46 Figure 10.7 Probability plot of RC and DD assay grade within Structural wireframes ........... 48 Figure 10.8 Histogram of RC and DD assay Data within Structural wireframes ..................... 48 Figure 10.9 Histogram of SG for Breccia Samples ..................................................................... 50 Figure 10.10 Histogram of SG for Volcanic Samples ................................................................... 50 Figure 10.11 Histogram of SG for Prophyry Andesite Samples .................................................. 50 Figure 10.12 Correlation plot of SG vs Gold ................................................................................. 51 Figure 10.13 Plan View of SG samples With 0.2 Wireframes ..................................................... 51 Figure 10.14 RC sample splitting at Amulsar................................................................................ 52

Figure 12.1 Drilling being observed during the site visit (2011) .............................................. 57 Figure 12.2 RC Assay QA/QC - Blank Samples` ........................................................................ 59 Figure 12.3 RC Assay QA/QC - Duplicate Samples ................................................................... 59 Figure 12.4 RC Assay QA/QC - CRM G300-7 (1.00ppm) ............................................................ 60 Figure 12.5 RC Assay QA/QC - CRM G302-2 (4.15ppm) ............................................................ 60 Figure 12.6 RC Assay QA/QC - CRM G398-6 (2.94ppm) ............................................................ 61 Figure 12.7 RC Assay QA/QC - CRM G904-8 (5.53ppm) ............................................................ 61 Figure 12.8 RC Assay QA/QC - CRM GLG 304-1 (0.15ppm), including Outliers ..................... 62 Figure 12.9 RC Assay QA/QC - CRM GLG 304-1 (0.15ppm), zoomed View excluding Outliers 62 Figure 12.10 RC Assay QA/QC – CRM G307-2 (1.08ppm) ........................................................... 63 Figure 12.11 DDH Assay QA/QC - Blank Samples ....................................................................... 64 Figure 12.12 DDH Assay QA/QC - Duplicate Samples ................................................................ 65 Figure 12.13 DDH Assay QA/QC - CRM G 300-7 (1.00ppm) ........................................................ 65 Figure 12.14 DDH Assay QA/QC - CRM G303-2 (4.15ppm) ......................................................... 66



List of Figures (Continued)

Figure 12.15 DDH Assay QA/QC - CRM G304-1 (0.15ppm) ......................................................... 66 Figure 12.16 DDH Assay QA/QC - CRM G398-6 (2.94ppm) ......................................................... 67 Figure 12.17 Assay QA/QC - CRM G904-8 (5.53ppm) .................................................................. 67 Figure 12.18 DDH Assay QA/QC - CRM G307-2 (1.08ppm) ......................................................... 68 Figure 12.19 DDH Assay QA/QC - CRM G900-6 (2.56ppm) ......................................................... 68 Figure 12.20 Diamond drilling Laboratory Duplicates ................................................................ 70 Figure 12.21 Laboratory RC drilling Duplicates ........................................................................... 71 Figure 12.22 Amulsar location map detailing magnetic and resistivity surveys ...................... 72 Figure 12.23 3D Magnetic Data sliced by Rl ................................................................................. 74 Figure 12.24 3D Resistivity Data sliced by Rl ............................................................................... 75 Figure 12.25 3D Chargeability Data sliced by Rl .......................................................................... 76 Figure 13.1 Gold Leach Curves ................................................................................................... 80 Figure 13.2 Column Leach Curves (-38 mm) .............................................................................. 83 Figure 13.3 Column Leach Curves (-19 mm) .............................................................................. 84 Figure 13.4 Column Leach Curves (Various Crush Sizes and Cyanide Concentrations) ..... 86 Figure 13.5 Coarse Bottle Roll Leach Recoveries (-12 mm) ..................................................... 88 Figure 13.6 Effect of Head Grade on Gold Leach Recovery ..................................................... 89 Figure 13.7 Solution Application Rate vs. Leach Recovery ..................................................... 91 Figure 15.1 Amulsar Resource Depth Distribution .................................................................... 97 Figure 15.2 Amulsar Selected Optimal Whittle Pit Shells (#36) ................................................ 100 Figure 15.3 Whittle Optimization Sensitivity Spider Graph (Cash Flow) – Option 1 .............. 101 Figure 15.4 Whittle Optimization Sensitivity Spider Graph (Cash Flow) – Option 2 .............. 102 Figure 16.1 Pit Design Terminology ............................................................................................ 104 Figure 16.2 Amulsar Pit Design - Plan View ............................................................................... 105 Figure 16.3 Amulsar Pit Design - Oblique View Looking North ............................................... 106 Figure 16.4 Amulsar Waste Dump Location and Design .......................................................... 110 Figure 17.1 Amulsar Overall Flowsheet ...................................................................................... 115 Figure 18.1 Proposed Overall Site General Arrangement Layout ............................................ 126 Figure 20.1 Surface Water Sampling Locations in Relation to the License Boundaries ....... 134 Figure 20.2 HLP Potential Sites ................................................................................................... 135 Figure 22.1 Amulsar Gold Project Pre-Tax Sensitivity IRR (Owner Operator) ........................ 170 Figure 22.2 Amulsar Gold Project Pre-Tax Sensitivity NPV@5% (Owner Operator) .............. 171 Figure 22.3 Amulsar Gold Project Pre-Tax Sensitivity IRR (Contract Mining) ........................ 175 Figure 22.4 Amulsar Gold Project Pre-Tax Sensitivity NPV@5% (Contract Mining) .............. 176 Figure 22.5 Amulsar Gold Project Pre-Tax Sensitivity IRR (Owner Operating w/ Erato) ....... 180 Figure 22.6 Amulsar Gold Project Pre-Tax Sensitivity NPV@5% (Owner Operating w/ Erato) 181 Figure 22.7 Amulsar Gold Project Pre-Tax Sensitivity IRR (Contract Mining w/ Erato) ......... 185 Figure 22.8 Amulsar Gold Project Pre-Tax Sensitivity NPV@5% (Contract Mining with Erato) 186



Glossary

Bankable Feasibility Study ................................................................. BFS Canadian Institute of Mining ............................................................... CIM Celsius .................................................................................... C Certified Reference Material ............................................................... CRM Centimeter .................................................................................... cm Day .................................................................................... d Degree .................................................................................... ° Diamond Core Drill Hole ..................................................................... DDH Gram .................................................................................... g Grams per ton ................................................................................... g/t Greater than .................................................................................... > Hour .................................................................................... h Hours per day .................................................................................... h/d Kilogram .................................................................................... kg Kiloliter .................................................................................... kl Kilometer .................................................................................... km Kilovolt .................................................................................... kV Less than .................................................................................... < Life of Mine .................................................................................... LOM Liter .................................................................................... l Megawatt .................................................................................... MW Meter .................................................................................... m Metric ton (tonne) ................................................................................ t Millimeter .................................................................................... mm Micrometer .................................................................................... µm Million .................................................................................... M Million tons .................................................................................... Mt Million tons per year (annum) .............................................................. Mtpa Mineral Resource Estimate ................................................................. MRE Troy Ounce .................................................................................... oz Parts per million .................................................................................. ppm Percent .................................................................................... % Potentially Mineable Mineralization .................................................... PMM Volt .................................................................................... V Year (annum) .................................................................................... a

Lydian International - Amulsar Heap Leach Facility Preliminary Economic Assessment Page 1


1.0 SUMMARY Lydian International, LTD. (Lydian) has engaged K D Engineering (KDE), an independent engineering firm located in Tucson, Arizona, to prepare a Preliminary Economic Assessment (PEA) of the Amulsar heap leach facility. The KDE study was prepared to standards pursuant to Canada’s National Instrument 43-101 (NI 43-101) and was based on the following:

▪ KDE site visit conducted in May 2011

▪ Geological review, mineral resource estimates, estimates of potentially mineable mineral (PMM) and mine capital and operating cost estimates completed by CSA Global (UK)

▪ Heap leach facility design and cost estimates by Golder Associates Inc.

(Golder)

▪ Metallurgical testing by SGS Lakefield, SGS Cornwall and Wardell Armstrong International

▪ Environmental and Social Impact Scoping Study February 2011 plus a July

review of the content of this PEA by Wardell Armstrong International 1.1 Geology, Exploration and Resource Estimation The Amulsar Project is covered by two special prospecting licenses (SPL) No 41 and 42, along with a small mining license, No 14/588. The licenses are held by Geoteam CJSC, a 95 percent owned subsidiary of Lydian. Lydian has the option to purchase the remaining 5 percent. The SPL’s were granted in 2009 and are valid for 5 years with the potential to extend the period. The mining license is valid for 25 years.

The Amulsar high sulphidation gold deposit is situated in central Armenia and is hosted in the Upper Eocene to Lower Oligocene calc-alkaline magmatic-arc system that extends north-westward through southern Georgia into Turkey and south-eastwards into the Alborz-Arc of Iran. The volcanic and volcano-sedimentary rocks of this system comprise a mixed marine and terrigenous sequence that is interpreted to have developed as a near-shore continental arc along the southern margin of the Eurasian Plate and at the northern limit of the Neo-Tethyan Ocean. The Neo-Tethyan ocean closed and subduction ceased along this margin in the Early Oligocene when a fragment of continental crust known as the Sakarya continent, collided at the trench axis and accreted with the Eurasian plate.

Geological structures represent a very important control on gold mineralization

throughout the Amulsar deposit. Both NE and NW striking, steeply dipping, structures are the dominant trends at Tigranes and Artavasdes, whilst at Erato the dominant structural trend strikes at approximately 010 degrees and dips steeply to the west. There is no one obvious alteration facies that hosts the gold mineralization at Amulsar,



and this relates in part to the fact that the gold resides in late structures that cut lithologies and units that have been affected by multiple alteration events. Lydian has developed the following sequence of alteration events at Amulsar. Sequence of Alteration Events

▪ The emplacement of a high level intrusive into a volcanic pile resulted in extensive clay alteration, followed closely by silicification of ledges within the core area with surrounding clay alteration remaining.

▪ Silica-alunite alteration with the generation of second phase breccias and the

introduction of sulphide (dominantly pyrite) gold-silver ± copper mineralization.

▪ Later supergene processes caused the oxidation of sulphide material within

silica altered bodies to depths of over 200 m. The alteration observed at Amulsar appears indicative of being in the deeper parts of the high sulphidation epithermal system. Gold mineralisation at Amulsar is a late event in the development of the deposit. Gold occurs within limonitic zones after the oxidation of sulphide material, and three dominant controls on the localisation of gold mineralization have been defined. Sequence of Mineralizing Events

▪ The argillic altered (kaolinised) porphyritic andesite formed an impermeable barrier to mineralising fluids, and caused ‘ponding’ of higher grade mineralization along the contacts

▪ Faults and fractures provided conduits for the introduction of gold mineralising

fluids

▪ Porous and permeable lithological units (hydrothermal breccias, volcaniclastic breccias, and occasional leached/vuggy volcanics) allowed for the lateral migration of mineralising fluids away from structurally controlled conduits

The presence of silver mineralization is not well understood, and does not appear to correlate strongly with gold grades. Silver values average between 2.0 g/t silver and 5 g/t silver, although tenors can reach greater than 100 g/t within narrow intervals. Copper and other base metals are virtually absent from the areas explored to date. Mineral Resource Estimate The May 2011 Mineral Resource estimate has been utilized for this PEA. This was an update of two previous Mineral Resource estimates that were released in March 2010 and March 2009. The April 2011 Mineral Resource estimate was based upon the



results of diamond and RC drilling completed between 2008 and 2010 plus additional grade data from channel sampling undertaken on outcrops at Amulsar. During 2010 Lydian staff undertook a thorough review of the Amulsar geology, including surface mapping, logging of oriented drill core and closer spaced infill drilling which resulted in the generation of a more robust geological model, where lower grade strata-bound mineralization and higher grade structural bound mineralization are defined. Diamond drill recovery during 2010 was good, increasing from 93 percent in 2009 to 96 percent in 2010. Reverse circulation (RC) recovery increased from 72 percent in 2009 to 89 percent in 2010. Overall there does not appear to be a strong trend of greater than 0.2 ppm gold value samples (resource grade) that exhibit poor recovery. Bulk density has been determined by paraffin wax sealed method, and for the current mineral resource estimate the following bulk densities were assigned to the dominant lithologies.

▪ Breccias = 2.33 g/cm3 ▪ Volcanic = 2.41 g/cm3 ▪ Porphyritic Andesite = 2.24 g/cm3

CSA Global (UK) (CSA) conducted a thorough review of the assay results

collected from RC drilling and diamond drilling and believe these to be reliable for the purposes of resource estimation. RC and diamond assay datasets exhibit similar population distributions and are considered compatible for use in resource estimation. Assays on blank material inserted by Lydian demonstrate minimal contamination whilst duplicate samples show acceptable levels of accuracy and precision. CRM charts show acceptable levels of accuracy and precision, however there are a number of outliers which appear to be a result of incorrect labelling of CRM’s.

The mineral resource estimate comprised two sets of domain wireframes, the first based on 0.2 g/t gold to 0.3 g/t gold sub-horizontal strata-bound material and the second based on 0.3 g/t greater gold focused on mineralization that is structurally controlled. Domain wireframes were created by Lydian with input from CSA. The structural domains were then grouped according to their dominant strike orientation. A total of 12 domains were identified and used during mineral resource estimation. For the mineral resource estimation procedure, sample data was composited to one meter intervals prior to top cut analysis. Variography was performed on the larger domains, with these models being applied to the smaller domains where cross validation showed a positive correlation between the two populations. As a result of the geostatistical assessment Ordinary Kriging (OK) was used for the estimation of gold grade in all domains. Gold grade was estimated into a block model with cells 20 m x 20 m x 5 m (X x Y x Z), sub celled down to 2 m x 2 m x 0.5 m (X x Y x Z). Bulk density was assigned according to dominant rock type.

The May 2011 Amulsar Mineral Resource estimate has been classified in accordance with CIM guidelines A CIM compliant Indicated and Inferred Mineral Resource, reported at a 0.4 g/t gold cut-off totals 80.7 million tonnes at 0.97 g/t gold for 2.52 million contained ounces of which 32.4 million tonnes at 1.1 g/t gold for 1.1 million



ounces are contained within the Indicated resource and 48.3 million tonnes at 0.9 g/t gold for 1.4 million ounces contained within the Inferred resource (Ref CSA NI 43-101 May 2011, quoted here to one decimal place).

1.2 Mining

The selected Amulsar Whittle optimisation shell for the Indicated resource option contains 31.0 million tonnes of Potentially Mineable Mineralisation (PMM), allowing for 5 percent PMM loss and 5 percent waste dilution, at an average head grade of 0.98 g/t and 75.9 million tonnes of waste, resulting in an overall strip ratio of 2.45:1. The selected pit shell provides a “specified discounted cash flow” of US$ 263.75 million with two push backs. These figures exclude capital expenditure.

The selected Amulsar Whittle optimisation shell for the Indicated and Inferred resource option contains 53.6 million tonnes of PMM, allowing for 5 percent PMM loss and 5 percent waste dilution, at an average head grade of 0.94 g/t and 138.0 million tonnes of waste resulting in an overall strip ratio of 2.57:1. The selected optimum pit shell provides a “specified discounted cash flow” of US$ 367.08 million with two push backs. These figures exclude capital expenditure.

The optimisation work that has been completed, based on the input parameters provided by Lydian, demonstrates that the Amulsar gold deposit has the potential to be a financially viable operation.

The conversion factors from resources to PMM are favourable for both the

Indicated resources option and the combined Indicated and Inferred resources option.

The limits of the Whittle pit shell generated using the current mineral resource model, demonstrates that exploration should be extended both laterally and at depth.

This initial optimisation work demonstrates that a portion of the resources which are currently in the Inferred Category have the potential to be economically mineable. Further drilling should therefore be carried out within this portion of the Inferred resources to bring them to the Indicated Category.

The aforementioned scoping study is based on parameters determined at a

conceptual stage. Additional work and detail is required to:

▪ Convert Inferred resources at Tigranes and Artavasdes to Indicated resources to enable potential conversion to Probable Reserves. This will require additional drilling and geological investigation.

▪ Convert at least the first 3 years of potentially mineable mineralization to

measured resources allowing conversion to Proven Reserves. This will need additional drilling and detailed geological analysis.



▪ Carry out adequate hydrological test work to quantify if there are risks associated with water inflow into the pit including groundwater, snow and rainfall in order to achieve the required level of confidence for pit slope designs and life-of-mine planning.

▪ Carry out adequate geotechnical test work to enable accurate design of pit

slope angles based on both wet and dry pit walls. 1.3 Metallurgy To date scoping level testwork has been completed by SGS Lakefield/SGS Cornwall (SGS) and Wardell Armstrong International (WAI) on samples from the Amulsar project. Initial exploratory testwork carried out by SGS investigated the following process routes:

▪ Gravity ▪ Whole ore cyanidation leach (Carbon in Leach) ▪ Coarse ore cyanidation leach (Heap Leach)

Gravity tests showed that the Gravity Recoverable Gold (GRG) component at Amulsar is very low with 8.5 percent GRG recovered to a low grade sulphide concentrate. Whole ore cyanidation leach tests conducted at various grind sizes ranging from 75 to 150 µm resulted in gold leach recoveries of circa 97 percent, after 48 hours of leaching.

Coarse ore bottle roll leach tests conducted at various crush sizes ranging from 1/2 to 1/4 inch resulted in a gold leach recovery of 96 percent at the finest crush size, after 15 days of leaching. Based on the preliminary characterization testwork it was concluded that the Amulsar ore would be amenable to processing by heap leach technology.

Further testwork was carried out on composites generated by combining core

samples from the Tigranes and Artavasdes deposits and considered by Lydian geologists to be representative. Again, testwork explored conventional CIL versus heap leach technology. A summary of the testwork results is shown in Table 1.1 below.

Table 1.1

Testwork Results Summary Sample Crush/Grind

Size/Test Gold Recovery, % by Composite

Comp A Comp B Comp C Head Grade g/t Au 1.60 2.42 3.76

80% -75 µm 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

Table 1.1 shows that there is a direct correlation between sample crush/grind

size and gold leach recovery, i.e. the finer the size of the sample the higher the gold



leach recovery as a result of better gold liberation. These initial scoping testwork results indicated good recoveries and potentially attractive processing economics of the Amulsar project if the bulk mined low-grade ore is prepared using a 3-stage crushing facility and heap leached.

Lydian commissioned WAI to undertake an additional program of laboratory

testwork on samples from the Amulsar deposit. WAI conducted further bottle roll and column tests on the two composite samples originally tested by SGS. The testwork generally focused on leaching at finer crush sizes (-38mm, -19mm and -12mm) and using higher cyanide concentrations (0.025 to 0.075 percent NaCN) than were used in the SGS testwork.

It is concluded that, from the column test results, the optimum crush size for both

samples is probably 19 millimeter and the optimum cyanide concentration is 0.05 percent, although further work to determine the impact of cyanide concentration upon silver extraction will be required.

In late 2010, Lydian shipped whole core from three diamond drill holes representing each of the deposits; Tigranes, Artavasdes and Erato. Composite samples were also prepared representing each of the main four metallurgical rock types including: pervasive iron oxide, gossan, fault gouge and siliceous breccia.

The objectives of the testwork program were to determine variability of gold leach recovery arising from rock type, and crush size. WAI were commissioned to conduct both coarse bottle roll leach tests and column leach tests. Tests were conducted on individual drill hole intervals, at crush sizes of -19 mm and -12 mm.

Results of the coarse bottle roll cyanidation leach tests conducted for 14 days are

summarized in Table 1.2.

Table 1.2 Coarse Bottle Roll Cyanidation Leach Test Summary

Drill Hole Deposit Leach Recovery, %

Gold Silver -12 mm -19 mm -12 mm -19 mm

DDAM68 Erato 97.0 96.5 42.7 45.9 DDAM70 Artavasdes 86.5 82.8 44.0 36.3 DDAM71 Tigranes 90.7 89.2 42.8 48.7

The coarse bottle roll leach tests again showed that a higher gold leach recovery

was achieved with the finer crush size. Gold leach recovery at -12mm was 2 percent higher on average, compared to those achieved at a crush size of -19mm.

The coarse bottle roll leach tests also showed that there was no leach variability

arising from the different rock types represented at the Tigranes, Artavasdes and Erato deposits. The outcome of the coarse bottle roll leach tests was to prepare composite samples for column leach tests. Composite samples representing the four major rock types and each of the deposits were prepared and are currently being leached.



Particle size determinations conducted on each of the composite samples

prepared for the column leach tests indicated that the percentage of -38 µm was on average 5 percent by weight for all column samples. Hence it was decided not to conduct any percolation/agglomeration tests ahead of the column leach tests. Percolation tests will be conducted once the column has been terminated.

The test results confirmed that the Amulsar gold mineralization is amenable to

recovery using heap leach technology. Conventional three-stage crushing, followed by stacking, leaching and carbon-in-column (CIC) gold recovery is suggested.

Initial column leaching was exceptionally rapid for both samples with up to 80

percent gold recovery being recovered after seven days of leaching at a crush size minus 12 mm. Initial column leach testing indicate that greater than 85 percent gold leach extraction is attainable after 70 days of leaching.

Column leach tests also indicate that cyanide and lime consumption rates are

low at 0.4 kg/t and 1.0 kg/t respectively. 1.4 Recovery Methods

Run-of-mine ore will be hauled from the pit to the three stage crushing plant located in proximity to the mine. Haulage from pit to crushing is 1 kilometer or less. The crushing plant consists of primary crushing through a jaw crusher, secondary crushing through a cone crusher, and tertiary crushing through a pair of cone crushers. The circuit will reduce ROM ore from minus 700 millimeter top size to a product of 80 percent passing 12 millimeters and is designed to process ore at a rate of 5 Mtpa. In the first year of operation 3.75 Mtpa will be processed and 5 Mtpa in year two. Installation of a duplicate circuit ramps up production to 7.5 Mtpa in year 3 and to 10 Mtpa for Year 4 and through 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 initially eight, ultimately 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 heap leach pad will be constructed in three phases. The first phase, suitable

for the first three years of operation, will have an area of 347,430 m2 and will accommodate 13.4 Mt of ore in six 8-m lifts. The second phase will add 431,100 m2 of pad area and accommodate 28.9 Mt of ore in four 8-m lifts above the Phase 1 pad heap and the Phase 2 pad. The third phase will add 265,630 m2 of pad area and accommodate 32.7 Mt of ore in six additional lifts, for an ultimate ore heap amount of 75 Mt.



Pregnant leach solution, intermediate leach solution and storm ponds will be located east 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 dosed to contain 0.5 gpl sodium cyanide will be applied by drip emitters to the top of the ore heap at irrigation rates of 10 l/h/m2 and will operate to reduce evaporation in summer and allow leaching to continue in winter to maximize PLS gold concentration. These 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. The primary leach cycle will be 30 days and the secondary 80 days. 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, 140 total leach days, the predicted overall recoveries of 85 and 40 percent for gold and silver respectively, will be attained with an overall leach solution to ore ratio of 3 m3/t.

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

collected in a network of perforated drain 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 pregnant leach solution will be pumped from the pregnant pond into the ADR

plant. Precious metal will be adsorbed from solution onto activated carbon counter-currently in five adsorption columns of 4-t carbon capacity each. Five additional carbon columns will be installed for the Phase II expansion. Carbonate scale will be removed from four-tonne 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 solution where precious metal will be deposited onto steel mesh cathodes. Weekly, the deposited metal will be washed from the cathodes, dried in a retort to volatilize and collect elemental mercury, and smelted in an induction furnace.

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

off-site for sale and refining.

1.5 Infrastructure

The Amulsar Gold Project covers an area of 130 km2, 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. Since the proposed mine site is near Jermuk there is no need for a construction camp for this construction effort. Currently an exploration camp is available at site which utilizes a portable generator. 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



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. 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. Community relation issues are currently handled by HSEC manager and a community liaison officer and a good understanding of local issues and sensitivities has been established. The strategy for accommodating all construction personnel, employees and security personnel during the construction period will be detailed as part of the Bankable Feasibility Study (BFS). 1.6 Environmental and Social Impact

Wardell Armstrong International Limited (WAI) was instructed by Lydian

International Limited (Lydian) to undertake an Environmental and Social Impact Assessment (ESIA) for the Amulsar open pit gold project in southern Armenia. The internationally compliant ESIA will be undertaken in parallel with the BFS in order that each process can inform the other and to ensure that environmental and social issues are considered throughout the mine design, construction, operation and closure stages.

The scoping phase of the ESIA has been completed. The purpose of the ESIA Scoping Study is to set out the main project parameters and identify any potential environmental and social impacts. It also sets out the national and international legislative framework and assesses the project status against international standards. Actions required have been identified in detail in a separate Gap Analysis. The main findings of these reports are summarised here, however, for completeness this report should be read in conjunction with the full versions of the Scoping Study and Gap Analysis. The Scoping Study identified no prohibitive social and environmental issues for the project. A number of environmental and social considerations were ascertained through this exercise and associated public consultation events. These issues will form a focal point for further baseline studies in order to better characterise them and provide information on how they might be mitigated. Risks will also be directly addressed via assessments, consultation and the preparation and implementation of relevant management plans. The issues identified, together with the steps required to further quantify and/or mitigate them, include:

▪ Groundwater and surface water hydrology, quality and sustainability - further hydro-geological study, modelling and geochemical assessment via the installation of monitoring wells needs to be undertaken. Site-wide water balance calculations need to be determined and the sustainability of sourcing examined and managed;



▪ Visual impact – The visual impact of the project needs to be assessed from Jermuk, the public highway and village communities, with the results included in public consultations;

▪ Traffic - quantification of mining traffic and a traffic impact assessment is

required;

▪ Land disturbance and take - a land use survey, biodiversity assessment and archaeological study are required to inform an impact assessment and determine any compensation required for grazing lands and any required mitigation or management;

▪ Biodiversity - field surveys need to be undertaken on areas of potential

major disturbance. An ethnobotanical study is also required. The results will inform an overall impact assessment;

▪ Dust - dust dispersion modelling and impact assessment on flora and fauna

used by communities needs to be undertaken;

▪ Noise - Noise impact modelling, incorporating the operating specifics of mine fleet and machinery needs to be undertaken to inform the noise impact assessment on the nearest sensitive receptors;

▪ Cyanide and hazardous materials - As the project will use cyanide in its

heap leach process, the usage, transport, handling and disposal of cyanide will need to be managed via formal plans and procedures in accordance with international best practice, particularly the International Cyanide Management Code (ICMC);

▪ Waste rock - acid- and leachate-generating potential of waste rock types is

being assessed via a program of geochemical testwork. Environmental protection measures and/or management techniques will be incorporated in dump design, if needed;

▪ Local labour - expectations require careful management in an area of high

unemployment and a skills audit to assess future potential employment opportunities should be undertaken;

▪ Social Baseline - supplemental demographic surveys and a health baseline

assessment (to include seasonal visitors) need to be undertaken;

▪ Local Industry - further appraisal of the main (agricultural and manufacturing) sectors is required, including individual contributions to domestic income. An assessment of potential diversification and increase in access to markets with advent of mining is considered beneficial;



▪ Public disclosure and consultation - to be undertaken in line with international best practice, at all stages of the ESIA and mine development processes (to include Jermuk) to manage expectations, allay concerns, establish NGO partnerships, inform community development and spending and influence mine design;

▪ Closure - considerations need to be addressed from conceptual stage via the

preparation and continual update of a Mine Closure and Rehabilitation Plan (MCRP) (to include heap leach rehabilitation).

The findings of the Scoping Study and the terms of reference for continuing work

were presented, in May 2011, to the neighbouring communities. Key issues and observations were noted and are accommodated in the steps listed above. Details of the minutes from these consultation events can be made available on request. Further ESIA public consultation will be held in late 2011 and will include an indication of proposed mine design.

Since the issue of the Scoping Study, Lydian has commissioned several discrete studies in response to the above and has pledged support and resources to the actions identified. Issues, “red flags” were identified that requires detailed studying. A work plan covering all of the red flag items identified above has been developed between Geoteam and WAI and work is progressing against each of these topics. WAI will monitor and review the process of environmental and social baseline data collection by Geoteam, and general progression of the ESIA works, and provide continuing technical support, as required, to ensure that both national and international requirements are met. The baseline programme is continuing and the next ESIA deliverable will be a baseline progress report in third quarter 2011.

1.7 Capital Cost Capital costs for the project were estimated by CSA Global (UK) for mining, KDE for the processing plant and Golder for the leach pad and ponds. The capital expenditures for the Amulsar Project will occur in two phases; years 1 to 3 is Phase I and years 4 through the remaining years is Phase II. The initial and sustaining capital costs for the base case owner operator and contract mining cases are summarized in Table 1.3.



Table 1.3

Total Initial and Future Sustaining Project Costs

Item Initial cost Sustaining cost Total

US$ US$ US$ Owner Operated Mining (OOM) Case - Tigranes/Artavasdes

Mining Cost 9,068,273 136,859,913 145,928,186 Buy out Newmont Cost 15,600,000 15,600,000 Process Plant Direct Cost 133,703,205 43,403,137 177,106,342 Leach Pads 19,756,625 12,641,830 32,398,454 Closure and Reclamation 7,719,137 7,719,137 Total Initial and Future Sustaining Project Cost 162,528,102 216,224,018 378,752,120

Contract Mining (CM) Case - Tigranes/Artavasdes

Mining Cost 9,085,196 12,682,616 21,767,812 Buy out Newmont Cost 15,600,000 15,600,000 Process Plant Direct Cost 133,703,205 43,403,137 177,106,342 Leach Pads 19,756,625 12,641,830 32,398,454 Closure and Reclamation 7,719,137 7,719,137 Total Initial and Future Sustaining Project Cost 162,545,025 92,046,720 254,591,745

1.8 Operating Costs Operating costs for the project were estimated with input from KDE, CSA and Golder.

On site operating costs for the base case are estimated to be $10.58/t of ore leached for the OOM option and $12.60/t for the CM option of ore leached including mining, processing, general and administrative (G & A), and plant services. These costs are summarized in Table 1.4.

Table 1.4 Cash Operating Cost

Mining Option

Unit Cost (US$/t)

Owner Operator Contract Mining Mining 6.84 8.86 Processing 3.49 3.49 General and Administrative 0.25 0.25 Total Operating Cost 10.58 12.6 Total Operating Cost (US$/oz) 419.34 499.40



1.9 Economic Analysis

The preliminary economic highlights (base case) are summarized in Table 1.5.

Table 1.5 Preliminary Economic Highlights (Base Case)

Units Owner Operator

Contract Mining

Average mined gold grade g/t 0.92 0.92 Steady state annual gold production (Yr 1-3) oz 123,000 123,000 Steady state annual gold production (Yr 4-7) oz 256,000 256,000 Life of Mine from production start Yr 7 7 Planned Steady State Production Rate (Yr 1-3) tpd 15,000 15,000 Planned Steady State Production Rate (Yr 4-7) tpd 30,000 30,000 IRR Pre tax % 39.50% 45.4% NPV Pre tax (5% discount rate) US$M 493.6 514.5 Payback period from start of production Yr 3.2 2.6 NPV Pre tax (0% discount rate) US$M 747.3 759.8 Initial Capital Cost US$M 162.5 162.6 Total Capital Cost US$M 378.8 254.6 Cash Costs US$/oz 419.3 499.4 Metallurgical Recovery % 85 85 Total Mined Gold to Leach Pad Moz 1.64 1.64

The financials for the base case mining options are summarized in Table 1.6.



Table 1.6

Economic Evaluation

Owner Operator Mining Case US $ x 1000

US $/t Resource

US $/oz Gold

Mine Gate Value of All Resource Net of Transportation and Refining 1,666,157 30.15 1,194.75 Mining Operating Cost (378,238) (6.84) (271.22) Processing Cost (192,943) (3.49) (138.35) General & Administration (13,616) (0.25) (9.76) Cash Operating Cost (584,797) (10.58) (419.34) Cash Operating Cash Flow 1,081,360 19.57 775.41 Capital Cost including Pre-Production Development (378,752) (6.85) (271.59) Before Tax Cash Flow 702,608 12.71 503.82

Contract Mining Case US $ x 1000

US $/t Resource

US $/oz Gold

Mine Gate Value of All Resource Net of Transportation and Refining 1,666,157 30.15 1,194.75 Mining Operating Cost (489,893) (8.86) (351.29) Processing Cost (192,943) (3.49) (138.35) General & Administration (13,616) (0.25) (9.76) Cash Operating Cost (696,453) (12.60) (499.40) Cash Operating Cash Flow 969,704 17.55 695.35 Capital Cost including Pre-Production Development (254,592) (4.61) (182.56) Before Tax Cash Flow 715,113 12.94 512.79

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 1.7.

Table 1.7 Summary of Key Financial Parameters (Sensitivity to Gold Price)

Owner Operator Mining Gold Price, US$/oz 1,100 1,200 1,300 1,400 1,500 Pre-Tax NPV@ 5%, (000's) 392,754 493,594 594,433 695,273 796,113IRR, Pre-Taxes 33.8% 39.5% 44.9% 49.9% 54.7% Payback, Operating Years 3.5 3.2 2.9 2.4 2.1

Contract Mining

Gold Price, US$/oz 1,100 1,200 1,300 1,400 1,500 Pre-Tax NPV@ 5%, (000's) 413,688 514,528 615,368 716,208 817,048IRR, Pre-Taxes 39.4% 45.4% 51.0% 56.2% 61.1% Payback, Operating Years 3.1 2.6 2.2 1.9 1.8

With a project IRR at US$ 1200/oz gold price of 40 to 45 percent and upside

resource potential, there are sufficient reasons for Lydian to advance the Amulsar project towards preparing the BFS to be based on optimization investigations into:



▪ Additional drilling to enhance classification and extension of mineral resources.

▪ Owner operated mining versus contract mining. ▪ Leasing vs. Direct Capital Purchase of Owner Fleet. ▪ In pit waste dumping. ▪ In pit crushing-conveying versus remotely located crushing plant. ▪ Single primary gyratory crusher versus two parallel primary jaw crushers. ▪ Metallurgical optimization on representative sample to fix design criteria and

costs to BFS standards and to enhance silver recovery. ▪ International ESIA compliance work. ▪ Plans for Project operating and project execution. ▪ Risk and Reward assessments.

1.10 Conclusions and Recommendations

This engineering study qualifies as a PEA level document according to Canadian National Instrument 43-101. The economic models utilized in this report indicated a range of before tax rate of return from 39.5 to 46.0 percent at an assumed gold price of US$ 1,200 per ounce. If the gold price used in the economic evaluation, which excludes inferred resources, is increased to current levels (US$ 1,500 per ounce in July 2011), then the before tax rate of return ranges from 54.7 to 61.8 percent.

The studies to date have not indicated any major difficulties or costs in

establishing the necessary infrastructure facilities to support the operation. Several opportunities were proposed by the client and their consultants to further reduce the overall capital or operating costs including the following:

▪ Project development of a phased approach to the construction effort.

▪ Additional drilling and metallurgical testing to define the Amulsar deposit.

▪ Further develop the trade off studies including contract vs. owner operated

mining and a crusher equipment layout and location optimization.

Based on the optimization work that has been completed, using the input parameters provided by Lydian, the Amulsar gold deposit has the potential to be a financially viable operation and it is recommended that this be advanced to a BFS.



2.0 INTRODUCTION This report forms an update to the report titled “Mineral Resource Estimate, Lydian International Limited, Amulsar Gold Project. 43-101 Technical Report, Armenia” dated 19 May 2011, prepared by Galen White and Nerys Walters of CSA Global Ltd (“CSA”), and filed on SEDAR, detailing recent resource development work carried out over the Amulsar Project, culminating in an updated Mineral Resource Estimate (“MRE”) for the project. This Technical Report has been prepared by Mr Joseph M. Keane, P.E. of K D Engineering (KDE) located in Tucson, Arizona, USA 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 through 13, 17 through 19, the process portions of Section 21, 22 through 24, 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. Galen White, BSc (Hons), FGS, MAusIMM, of CSA, is the qualified person for all matters relating to the mineral resource estimate and is responsible for Sections14 and 15

Mr. Richard Kiel, P.E., of Golder, visited the property in June 2011, and is the Qualified Person for all matters relating to the leach pad/pond design, portions of Sections 17 and 18 and the geotechnical portions of Section 21.

Mr. Kent Bannister, MAusIMM CP, of CSA, is the qualified person for all matters relating to open pit and mine design and is responsible for Section 16 and the mining portions of Section 21.

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.



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

This independent Report has been prepared following resource development activities over the project during 2010 and the first six months of 2011. The following information has been utilized during the preparation of this report. Post-December 2009 technical data, documents, reports and information supplied by Lydian, including:

▪ Mineral Resource Estimate, Lydian International Limited, Amulsar Gold Project. 43-101 Technical Report, Armenia, May 2011, prepared by CSA.

▪ ESIA Scoping Study by Wardell Armstrong International, dated 4 August

2011

▪ Process engineering designs completed in July 2011 to support this report by KDE.

▪ Geotechnical designs concerned with hydrological issues and leach pad

designs produced for this report by Golder, reported on July 19, 2011 (Golder).

▪ Preliminary mining leach pad, and process capital and operating cost

estimates produced by CSA, KDE, and Golder.

▪ Reports and data in the public domain.



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 two 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.



4.0 PROPERTY DESCRIPTION AND LOCATION

The following property description and location is a summary taken from the 19 May 2011 NI 43-101 report mentioned in the Introduction of this report and from Lydian documentation.

4.1 Summary The Amulsar Gold Project covers an area of 130 km2, located in south central Armenia (Figure 4.1). The project area is covered by two Special Prospecting Licenses (SPL). The Saravan license (No 42) and the Gorayk license (No 41). A small 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. A 17 km2 extension to the Saravan license was awarded in May 2011. The Armenian Ministry of Energy and Natural Resources approved an initial C1+C2 reserve of 16.38 t of gold metal in balance and 3.6 t of gold off-balance for the Tigranes site at Amulsar on February 23, 2009. These soviet-style reserves are not CIM compliant. Subsequently an application was submitted for a Mining License which was granted on 4th April 2009. Mining License No 14/588 is valid for 25 years, although under the approved work programme the resource would be exploited in 14.4 years. Lydian has submitted an application to the Armenian Ministry of Energy and Natural Resources, in July 2011, to update the approved resources and to cover both the Tigranes and Artavasdes areas. Following approval of the updated resources Lydian intends to apply for a revised Mining License covering the enlarged resource area. 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).



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. There is also a 5-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 41 and 42) were awarded to Lydian at auction and were granted under the 2007 mining law. The World Bank is advising the Armenian Government on revisions to the existing mining laws, including royalties, and a revised mining code is expected to be finalized by the end of 2011.



4.2 Armenian Minerals Legislation

A Prospecting License is an exclusive license granted for a term of three years without extension. It is activated by execution of a Concession Contract which regulates the licensee’s activities under the law. If exploration results prove positive, and if there is an indication of economically significant resources, a Prospecting License must be converted within its term to a Special Prospecting License or to a Mining License. A Special Prospecting License is current for a period of between three and five years, but with up to two extensions possible for periods of two years each. On application for a Special Prospecting License the licensee can conclude a Stabilization Contract which is a written agreement with the Republic of Armenia regulating the licensee’s advanced exploration activities under the law. A Mining License is valid for a period of up to 25 years but the actual term is based on the mining plan submitted as part of the Mining License application and the time required to exploit the resource. In the case of Mining License No 14/588 the resource will be exploited in 14.4 years based on the submitted mining rate of 1.5 Mtpa.

Application for a Special Mining License can be submitted at any time but usually follows granting of a Special Prospecting License. Special Prospecting License holders have the entitlement to convert the license to a Special Mining License upon application. Special Mining Licenses are current for a period from twelve to twenty five years, but can be extended upon application by the licensee. Under the 2007 regulations the Prospecting License stage has been annulled for newly awarded ground and only Special Prospecting Licenses, current over 5-year terms, are granted at auction. The execution of a Concession Contract is still required. Concession Contracts for all three Lydian licenses have been executed. 4.2.1 Royalties Armenian Mining royalty tax comprises two components, a Natural Reserves Depletion Fee and Royalty. These are calculated as follows:

▪ Natural Reserves Depletion Fee This is payable on quarterly basis before the 1st day of the second month of the following quarter. The Natural Reserves Depletion Fee is calculated on 1.5 percent of the quantity of the officially approved precious metal reserves depleted during the quarter, using metal prices announced by the Armenian government (based on international market prices).

▪ Royalty Royalties are payable if profitability of a mining license holder exceeds 25 percent. The calculation of the royalty and payment is done on a quarterly basis. Profitability is calculated by the following formula (R-C)/R, where; R - is revenue from mined reserves sold



C - is operational expenses for the same period For each 0.8 percent increase exceeding 25 percent profitability, the payable royalty is 0.1 percent of the revenue amount. 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.

The royalties are payable before 30th of the month following the reportable period. 4.2.2 Armenia - Newmont Joint Venture Agreement

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 is due at the end of 2011 and the final tranche is to be paid either at the end of 2012 or 90 days after completion of a BFS and all permissions to move into production are in place, whichever occurs earliest. 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. 4.2.3 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 20.7.



5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

The following property accessibility, climate, local resources, infrastructure and physiography descriptions are taken from the NI 43-101 report dated the 19th May 2011 mentioned in the Introduction of this report. 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. Until all season roads are constructed, snow plows are purchased and routine maintenance is ongoing access is generally possible only from March to November and access for heavy machinery is confined to the period from May to October/November. 5.3 Resources and Infrastructure Infrastructure in the main areas of 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 centres.



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

Green- Amulsar Mining License; Red-Saravan license; Blue-Gorayk license 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 country lies within a seismically active zone, with some areas having a high risk of major earthquakes. The last major quake was in 1988 in northern Armenia. The Amulsar area is also seismically active with the most recent destructive earthquake being recorded in September 1931 at the township of Sisian some 30 km to the southeast of the prospect area. Destructive earthquakes have also been recorded in the ancient city of Vayk, some 20 km to the west. The core of the Amulsar area comprises a mountainous terrane, an approximately 7 km long northwest-southeast trending ridge which reaches a maximum altitude of 2,988 m. The mountainous area around Amulsar is quite rounded and, with



vegetation limited to wild grasses and isolated scrub, access is relatively straightforward outside of the winter period.



6.0 HISTORY

The following history descriptions are taken from the NI 43-101 report dated 19th May 2011 mentioned in the Introduction of this report. 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, prior to the recent downturn in metal prices. 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 have 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 companies active in Armenia include Global Gold, Caldera Resources and Anglo African Minerals. 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 m3 of trenching were also completed during this time, mostly on the north-eastern side of the ridge. 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.



7.0 GEOLOGICAL SETTING The following geological setting descriptions are taken from the NI 43-101 report dated 19th May 2011 mentioned in the Introduction of this report. 7.1 Regional Geological Setting

The following section is taken from an internal Lydian report detailing the geological setting at Amulsar.

The Amulsar high sulphidation gold deposit is situated in central Armenia and is hosted in the Upper Eocene to Lower Oligocene calc-alkaline magmatic-arc system that extends north-westward through southern Georgia into Turkey and south-eastwards into the Alborz-Arc of Iran.

Volcanic and volcano-sedimentary rocks of this system comprise a mixed marine

and terrigenous sequence that is interpreted to have developed as a near-shore continental arc along the southern margin of the Eurasian Plate and at the northern limit of the Neo-Tethyan Ocean. The Neo-Tethyan closed and subduction ceased along this margin in the Early Oligocene when a fragment of continental crust, known as the Sakarya continent, collided at the trench axis and accreted with the Eurasian plate. The location of Amulsar within this arc is shown below in Figure 7.1.



Figure 7.1 - Distribution of prospective Eocene-Oligocene arc-magmatic rocks and major known mineral occurrences draped on SRTM DEM. Blue circles indicate copper-molybdenum resources,

yellow circles indicate gold dominant resources. 7.2 Local Geologic Setting

The oldest rocks within the Saravan and Gorayk license areas are siliciclastic marine sediments found in the north-west of the Saravan License area and interpreted to be Upper Palaeocene in age. These comprise westerly dipping (approx. 30°) coarsely bedded, polymictic, matrix supported conglomerates with occasional Bivalve fossils preserved. Clasts are sub-rounded and composed of variable intermediate composition volcanics. The lithology is interpreted as having formed in a near shore, relatively high-energy marine shelf environment (Figure 7.2A).

These conglomerates are overlain by a thin (less than 20 m thick) tuffaceous unit that hosts at least two separate known zones of carbonate vein-style base metal



mineralization. Veins are relatively narrow (30 mm to 200 mm wide), and the largest identified zone of veining is less than 5 m in width, the along-strike and down-dip extent of which is unknown. The dominant ore mineral is copper-bearing azurite and minor lead-bearing galena (refer to Figure 7.2 right hand image).

Figure 7.2 - (Left Image) - Siliciclastic marine sediments found to the north-west of the Saravan

license area. (Right Image) - Carbonate vein-style base metal mineralization (Sample No. AR2777: 1805 g/t Ag, 14% Cu, 1.94% Zn).

These conglomerates and tuffs are uncomformably overlain by an Upper Eocene

age, thick succession of northward dipping (approx. 10°) dark-grey, monolithic, sub-aqueous, basalt-andesite composition volcano-sedimentary breccia units that grade into increasingly coherent extrusive volcanic lava flows which exhibit autoclastic "quench" textures higher in the sequence, formed as a result of extrusion in a sub-aqueous environment (Figure 7.3).

There are numerous basaltic sills less than 2 m in thickness that can be traced

for greater than 1 km along bedding planes within the sequence. These volcanic/volcano-sedimentary units cover a large area and can be seen surrounding Saravan Village and to the north-east of Gorayk Village.



Figure 7.3 A. Basaltic sills within volcano sedimentary units.

B & C. Examples of andesitic-basaltic volcano sedimentary breccia units.

The bedded lithologies detailed above were subsequently intruded into by a large, hypabyssal andesite dome that dominates the Amulsar Mountain, with associated lava flows that overlie the bedded units at lower elevations to the north and west. The intrusion has a fine to medium grained porphyritic crystalline texture, with an aphanitic groundmass and plagioclase phenocrysts ranging from less than 1 mm to 3 mm.

Lydian has created a geological map of the license area and two schematic cross

sections, Figures 7.4 through 7.6.



Figure 7.4 - Geological map of the Amulsar area showing locations of sections

Figure 7.5 - A-A’ cross section showing geology from Erato to Tigranes.



Figure 7.6 - B - B’ cross section showing geology from Artavasdes to the eastern and northern

slopes of Tigranes.

The Amulsar gold deposit itself is situated along the ridge-top of the mountain, hosted in a sequence of Eocene-Oligocene volcanic andesites and volcaniclastic rocks that overlie the lithologies detailed above and dip at approximately 20° to the north east. These rocks have commonly been affected by multi-stage acid-leaching hydrothermal alteration events which have produced widespread texturally destructive silicification and hence it is often very difficult to identify any primary composition or texture and there is no known unaltered examples. Poorly sorted, coarse grained (5 mm to 200 mm), matrix supported, sub-angular clasts are characteristic of the volcaniclastic units which are most commonly found lower on the flanks of the deposit at Erato and North Erato; flow-banding is a common feature within the volcanic units that dominate the Tigranes and Artavasdes areas. These units were later intruded by a large, irregular sill-like feldspar-augite porphyritic andesite, with feldspar phenocrysts less than 5 mm within an aphanitic matrix.

Episodic fault activity and high level magmatism has been responsible for the

development of multi-stage phreatic and hydrothermal breccias that form large bodies throughout the deposit. The character of these breccias is variable across the deposit, although it appears that they formed predominantly during two separate events. The relative ages of these breccias is best determined by the alteration of matrix and their component clasts, and cement type. The oldest breccias are pre-mineralization, altered to massive silica to silica clay, grey in color, and monomictic to polymictic; they are also relatively competent with poor permeability due to the strong silica massive cement. The second phase of brecciation and alteration appears to have occurred shortly after the intrusion of the feldspar-augite porphyritic andesite sill, and was potentially contemporaneous with gold mineralization. These breccias can occur as monomictic where found on the margins of the porphyritic andesite sill, yet are predominantly polymictic, poorly sorted with rounded ‘milled’ clasts, although jigsaw breccias are occasionally present to a much lesser extent. The discriminating feature of these second stage breccias is the hydrothermal fluid derived cement composed of hematite dusted silica and having a ‘honey’ brown color, there is also a close association with gold mineralization and silica-alunite alteration. Clasts of the early ‘grey’ silica massive altered lithologies and occasionally clasts of porphyritic andesite can be found within these second phase breccias. Associated with this second brecciation was the formation of massive sulphide cemented hydrothermal breccias and veins with clasts of



early ‘grey’ silica altered volcanics. These have been entirely oxidized by supergene processes, altering the massive sulfides to hematite-jarosite cement.

The porphyritic feldspar-augite andesite is interpreted to have been emplaced post the first silica alteration and breccia event but prior to the second and gold mineralization. This inference can be made because clasts of the early ‘grey’ silica altered lithologies are found within the porphyritic andesite (Figure 7.7 Left Image), yet it is brecciated by the second stage breccias and localisation of higher grade gold mineralization often occurs at its boundaries due to its capacity as an impermeability barrier to hydrothermal fluids.

At lower elevations than the outcropping Amulsar deposit, unaltered Quaternary basalt flows seen to the north, east and south, overlay the Lower Tertiary volcano-sedimentary and intrusive rocks.

Figure 7.7 – (Left Image) - DDA-067, early ‘grey’ silica massive and porphyritic feldspar-augite andesite clasts within a second phase hydrothermal breccia with limonitic silica-clay cement.

(Right Image) – DDA-067, early ‘grey’ silica massive clasts within a second phase hydrothermal breccia.



Figure 7.8 – (Left Image) - Oxidised massive sulphide cemented hydrothermal breccia with clasts

of early ‘grey’ massive silica altered volcanics. (Right Image) – Limonitic, silica altered flow-banded volcanics.

7.3 Structure

Structures represent a very important control on gold mineralization throughout the Amulsar deposit, hence surface and down-hole mapping of structures has been carried out across the project area, and has been integral in the generation of the geologic and resource models. From Lydian’s work it is clear that both northeast and northwest striking, steeply dipping trends show most strongly at Tigranes and Artavasdes, although the orientations are not tightly constrained and show a continuation of possible strike orientations between northeast and west-northwest. At Erato the dominant structural trend observed in approximately striking 010° and dipping steeply to the west, however down-hole and surficial structural data is limited at present. High grade structures are seen as faults, veins, and fractures, with no significant orientation controlling higher grades than the other. 7.4 Alteration

There is no one obvious alteration facies that hosts the gold at Amulsar. This relates in part to the fact that the gold resides in late structures that cut lithologies and units that have been affected by multiple alteration events. Silica is a universal component of the gold mineralized intervals, but the high grade gossans often lack any alteration characteristics that can be recorded other than quartz and alunite, hence the strong alunite association.



7.5 Sequence of Alteration Events

Lydian proposes the following sequence of alteration events at Amulsar;

▪ The emplacement of a high level intrusive into the volcanic pile and generation of copious amounts of hot water resulted in extensive clay alteration, followed closely by silicification of ledges within the core area with surrounding clay alteration remaining.

▪ Silica-alunite alteration with the generation of second phase breccias and the

introduction of sulphide (dominantly pyrite) gold-silver ± copper mineralization. The porphyritic feldspar-augite andesite underwent argillic alteration in which the plagioclase feldspar phenocrysts are replaced by kaolinite, the augite phenocrysts by pyrite and the groundmass by fine-grained quartz. Outwards from the silica – silica-alunite ledges, the groundmass quartz content of the kaolinised rock decreases, forming an impermeable barrier and causing ‘ponding’ of mineralization along the contacts.

▪ Later supergene processes caused the oxidation of sulphide material within

silica altered bodies to depths of over 200 m, converting sulphides to hematite and jarosite and liberating gold and silver. The massive quartz bodies are normally a product of wholesale silicification. Where the quartz displays a vuggy texture, it is commonly due to supergene oxidation and removal of pyrite grains rather than to the hypogene leaching of silicate minerals. The sulphur released is believed to have caused the leaching of copper, removing it from the system and explaining the complete lack of copper mineralization within the area drilled so far.

The alteration observed at Amulsar appears indicative of being in the deeper parts of the high sulphidation epithermal system. This hypothesis is supported by the presence of steeply dipping, structurally controlled silica ledges, and the lack of shallowly inclined silicified horizons generated at paleo-water tables and/or powdery advanced argillic alteration formed in the steam-heated environment between paleo-water tables and paleosurfaces. Also the observed presence of wormy (elsewhere termed gusano texture) pyrophyllite replacement textures that is formed at temperatures of over 240°C and is common near the base of silicic lithocaps, and in turn over the top of shallow porphyry systems (e.g. Yanacocha, Tucari, and Tantuatay).



8.0 DEPOSIT TYPES AND MINERALISATION The following deposit types and mineralization descriptions are taken from the NI 43-101 report dated 19th May 2011 mentioned in the Introduction of this report. Gold mineralisation at Amulsar is seen to have been a late event in the development of the deposit; hence the spatial distribution was therefore controlled by all alteration events and structures. Gold occurs within limonitic zones after the oxidation of sulphide material, and three dominant controls on the localisation of gold mineralisation have been defined:

▪ The argillic altered (kaolinised) porphyritic andesite formed an impermeable barrier to mineralising fluids, and caused ‘ponding’ of higher grade mineralisation along the contacts, often appearing as thick, leached and ‘gossanous’ zones. This is present throughout the deposit: in the Tigranes area the contacts are irregular yet generally steeply dipping, whereas in parts of Artavasdes and at Erato, the argillic altered porphyritic andesite forms an overlying ‘cap’ which results in the localisation of sub-horizontal mineralisation within siliceous units below.

▪ Faults and fractures provided conduits for the introduction of gold mineralising

fluids, resulting in high grade ‘gossanous’ veins and clay-gouge dominated faults. These often form broad corridors of numerous, closely spaced high-grade structures, and can also be seen to cross-cut the argillic altered porphyry andesite, and allow for discrete zones of mineralisation; such as between Tigranes and Artavasdes where mineralisation is seen to be continuous between the two areas along such structures.

▪ Porous and permeable lithological units (hydrothermal breccias, volcaniclastic

breccias, and occasional leached/vuggy volcanics) allowed for the lateral migration of mineralising fluids away from structurally controlled conduits, and formed large tonnages of lower grade material. This is seen particularly well in Erato, where weaker silicification has preserved the intrinsic permeability of the original lithologies and resulted in broad areas of consistent gold mineralisation.

The presence of silver mineralisation is as yet poorly understood, and does not appear to correlate strongly with gold grades. Silver values at Amulsar average 2 to 5 g/t silver, although can reach greater than 100 g/t silver within narrow intervals. Copper and other base metal mineralisation is virtually absent from the areas drilled so far, although this is present within carbonate veins from lower stratigraphic elevations in the west of the Amulsar license area. A small silver mining project adjoins the Amulsar licence 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.



Lydian have indicated that genetic similarities exist between the mineralisation at Amulsar and Gryphon Gold Corporation’s Borealis gold project in Nevada, USA, however the alteration patterns differ in detail. Both are high sulphidation epithermal gold deposits which have higher grade, structurally controlled mineralisation surrounded by a lower grade, strata bound halo of mineralisation. A generic type section of the Borealis deposit can be found in Figure 8.1.

Figure 8.1 - Typical alteration patterns of the Borealis District Gold Deposit (Gryphon Gold Corp)



9.0 EXPLORATION

The following exploration information is taken from the NI 43-101 report dated 19th May 2011 mentioned in the Introduction of this report.

The work program at Amulsar has been designed to define the scale of the epithermal system and the extent and grade of the gold mineralization within that system. 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. This work is summarised in the Table 9.1 below.

Table 9.1 Summary of Main 2010 Exploration Activities on the Amulsar Gold Project

Item Unit Quantity RC drilling Meter 17,036 (127 holes)

Diamond drilling Meter 7,502 (49 holes) Channel Sampling Meter 1,092 ( 43 Channels)



10.0 DRILLING The following drilling information is taken from the NI 43-101 report dated 19th May 2011 mentioned in the Introduction of this report 10.1 Drilling Lydian has carried out two types of drilling at Amulsar during 2010. Reverse circulation (RC) and diamond (DDH) drilling was undertaken. A total of 49 core holes and 127 RC holes drilled, totalling 24,539 m, comprising 7,502 m of diamond drilling and 17,036 m of RC drilling. Drill-Ex International was responsible for all of the drilling on site during this time. Drill holes were angled and were drilled either towards the northwest or southeast (in the northern area) and southwest or southeast (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 40 m apart, with holes generally drilled at 40 m 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 to 250 m but most holes were drilled to approximately 140 m down hole depth. The diamond drill holes were usually drilled at PQ diameter for the first 20 m to 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, no deviation issues were identified. After logging of rock chips or drill core, samples were taken, generally at one-meter intervals throughout the entire hole, and dispatched for multi-element analysis and gold assay. A total of 7,082 samples were assayed from drill core, as well as 16,962 rock chip samples from RC holes. Analysis for silver was also undertaken; however samples were in some instances composited up to 6 m to reduce costs and limited QA/QC data was acquired during analysis for silver. If Lydian is interested in including silver assays as part of future MREs then CSA recommend re-sampling at original intervals, with representative QA/QC sample support. Silver results were not reported as part of this MRE. Drill results, lithology interpretations and assay results have further defined the two previously identified zones of gold mineralization, Tigranes and Artavasdes. Increased drilling to the north has defined an additional area of gold mineralisation, Erato. Many of the holes show a pattern of relatively widespread low grade mineralization (generally less than 1g/t gold) within silica-altered zones (sometimes with



iron oxide matrix), with shorter higher grade intervals generally associated with structurally controlled higher grade iron oxide filled faults and fractures (Figure 10.1 & Figure 10.2).

Figure 10.1 - Left Hole DDA-007 Showing

Porous Limonite Breccia Zone 20m Depth (>7g/t gold)

Figure 10.2 - Right Hole DDA-07 Silicified Breccia Zone with Partial Limonite Matrix

43mDepth(approx.1g/t gold) Based on drilling to date, the Amulsar deposit remains open at depth. The Artavasdes area remains open along strike and to the south. The area to northeast of Tigranes has untested potential on the extensions of the main northeast trending mineralised faults present on Tigranes. The Erato area remains open in all directions and limited drilling has occurred at the Erato North target. Further infill drilling at Tigranes and Artavasdes is required in order to increase the geological and grade continuity such that currently defined resources can be upgraded to higher resource categories. Step-out drilling along strike and to the east of Artavasdes area should be considered in order to investigate mineral continuity in areas of sparse drilling. Infill drilling is required at Erato in order to adequately confirm geological and grade continuity such that the defined resource can be upgraded to a higher resource category. Drilling to further investigate mineral potential at Erato north along with other exploration targets should be undertaken to assess the likelihood of augmenting the current resource base with resources from these areas.



Highlights from the 2010 drilling program include the following resource infill and extension holes at Artavasdes, Tigranes and Erato contained in Table 10.1.

Table 10.1 Selected Down Hole Gold Intercepts from the Lydian 2010 Amulsar Drilling Program

Area Hole ID From (m) To (m) Downhole Interval Grade g/t Au Artavasdes DDA-047 88 112 24 7.4 Artavasdes DDA-066 107 175 67 1.8 Artavasdes RCA-191 1 52 51 1.8 Artavasdes RCA-915 83 167 84 4.9 Artavasdes RCA-208 102 197 95 1.3 Artavasdes RCA-212 86 172 86 2.7 Artavasdes RCA-301 121 161 40 5.6

Tigranes DDA-036 8 93 85 2.9 Tigranes DDA-076 11 100 89 1.9 Tigranes RCA-177 0 112 112 1.0 Tigranes RCA-193 0 74 74 1.4 Tigranes RCA-198 98 149 51 2.3 Tigranes RCA-226 0 91 92 1.2 Tigranes RCA-258 0 69 69 1.7

Erato RCA-233 36 131 95 1.7 Erato RCA-280 85 170 85 1.2 Erato RCA-281 14 200 186 1.1

*Down hole interval. DDA = diamond drill holes; RCA – Reverse Circulation Holes NB: cut-off 0.2g/t Au, Maximum down-hole internal Dilution 10m. These are down-hole intervals and not true widths; true thickness can be up to 50% less than the down-hole intervals.

Notable exploratory and step out results are detailed in Table 10.2.

Table 10.2 Selected Down Hole Gold Intercepts from the Lydian 2010 Amulsar Drilling Program

Area Hole ID From (m) To (m) Downhole

Interval Grade g/t Au

Arshak DDA-034 30 64 34 1.0 Arshak RCA-276 72 125 53 1.0

Artavasdes W RCA-284 64 93 29 1.0 Tigranes E DDA-058 49 89 40 1.2 Tigranes E RCA-220 84 121 37 1.2 Tigranes N DDA-061 227.4 251 23.6 1.1 Tigranes N RCA-292 35 49 14 1.6

*Down hole interval. DDA = diamond drill holes; RCA – Reverse Circulation Holes NB: cut-off 0.2g/t Au, Maximum down-hole internal Dilution 10m. These are down-hole intervals and not true widths; true thickness can be up to 50% less than the down-hole intervals.

CSA consider these results to be appropriate for the type of deposit and all selection criteria are clearly detailed in company press releases.



10.2 Planned Drilling for 2011 Lydian has planned drilling totalling 45,000 m and aims to complete at least 30,000 m during the 2011 drill season, comprising 15,000m of DD and another 15,000 to 30,000 m of RC drilling. Planned collar positions, compared to historic drilling are displayed in Figure 10.3. This total excludes holes planned for metallurgical testwork, hydrological monitoring and pump testing, geotechnical investigations, and sterilization drilling planned as part of the planned BFS. Lydian states their drilling objectives as;

▪ Infill drilling - to close off gaps within the current Indicated resource and to upgrade Inferred portions of the resource that are located within the current pit design. A significant amount of this drilling will be in boulder field areas which have not been accessed previously for drilling.

▪ Step out drilling - to test the lateral extensions of the mineralised faults that

have been identified in the current model and to upgrade the classification of Exploration Potential Material.

▪ Deep drilling - to test the depth extents of mineralisation below current drilling. ▪ BFS drilling - to obtain core for geotechnical, hydro-geological and

metallurgical requirements for the BFS. Lydian consider the planned drilling, along with improvements in wireframe models and planned improvements to QA/QC procedures should allow the upgrade of a portion of the current resource to Measured Resource, to be considered as part of the next resource update. CSA offers the following comment on planned drilling;

▪ The current plan is ambitious, but should be achievable as long as Lydian can source enough drill equipment and that the equipment doesn’t suffer from any major breakdowns, or unexpected delays.

▪ Planned infill drilling is considered sufficient to support an extended Indicated

Resource within the current planned pit.

▪ Step out drilling is sufficient to further define current Exploration Potential material, with the possibility for some to be upgraded into an Inferred category.

▪ Exploration holes are focussed on defined targets from region exploration

work and should investigate the potential for mineralisation in these areas.



CSA cannot say that the planned drilling will increase, or upgrade the current resource estimate, or its classification. This is entirely dependent upon the results of the planned drilling.

Figure 10.3 - Map of Planned Drilling for 2011 at Amulsar

10.3 Channel Sampling A total of 1,092 m collected from 43 channels from across the Amulsar area. Channels were cut by a two bladed angle grinder to approximately 2 cm depth, Figure 10.4. Perpendicular cuts were made to facilitate sampling which was undertaken using a hand chisel and a hammer. Samples were collected below the channel in a sample bag, Figure 10.5. Samples were delineated according to the geological and alteration styles identified.



Figure 10.4 - Channel cutting on site at Amulsar

Figure 10.5 - Channel sampling on site at Amulsar

All analysis was undertaken under the same procedures as for drill samples. Historic trench data is also available at Amulsar. This data was not used during this resource estimate as Lydian was concerned about the reliability of data collected from these trenches.



Figure 10.6 - Map of Amulsar Area Showing Location of trenches in green and channels in red

Drill hole collars are represented by small black dots 10.4 RC and DDH Sample Recovery Investigation During the 2008 drill campaign core recovery averaged approximately 91 percent, with RC recovery close to 100 percent. RC and DD recovery during 2009 decreased to 72 and 93 percent respectively. Recoveries from the 2010 drill campaign have increased with diamond core recovery at 96 percent and RC recovery at 89 percent. CSA undertook a review of RC sample recovery versus assay grade to assess this relationship and to determine the level of confidence that can be applied to the representivity of samples within mineralized zones. It was observed that the majority of samples within the interpreted domains exhibited a range of recovery values between 60 and 120 percent. Very low sample recoveries dominantly occur in near-surface areas, to a depth of 10 m. Sample recoveries exceeding 100 percent may occur at down-hole depths corresponding to the end of a drilled run, and may be due to sample accumulation in the cyclone.



Correlation plots of grade versus recovery within each of the 12 interpreted domains are contained in the May 19, 2011 43-101 Technical Report mentioned in the Introduction. It was observed that a broad spread of recoveries occurred within each domain. However the majority of material was collected with recoveries greater than 60 percent. Overall there does not appear to be a strong trend of >0.2ppm gold samples (resource grade) exhibiting poor recovery. However, clearly samples that have been affected by poor sample recovery are present within the mineralized zones, and to some extent the representivity of sample grades may be brought in to question and some grades may be overstated or understated, to an unknown degree. As part of future resource development, Lydian should undertake a detailed investigative study of the RC sample recovery versus grade relationship, focusing on an analysis of spatial representivity and drilling practices, as well as investigating whether sampled material could have been preferentially lost during drilling. Overall CSA considers that the effect of poor sample recovery in mineralised domains to be minimal. 10.5 Twin Hole Investigation Lydian under took a twin drill hole investigation during exploration and resource drilling at Amulsar. A total of 14 RC and DD holes were twinned, see Table 10.3.

Table 10.3 RC and DD Twinned HolesRC DD

RCA-040 DDA-002 RCA-038 DDA-007 RCA-039 DDA-009/DDA-044 RCA-037 DDA-011 RCA-253 DDA-017 RCA-224 DDA-035 RCA-187 DDA-138 RCA-181 DDA-041 RCA-144 DDA-042 RCA-156 DDA-045 RCA-021 DDA-046 RCA-084 DDA-047 RCA-252 DDA-048 RCA-060 DDA-050

A review of the global data populations for RC and DD data was undertaken and is shown in the following P-Plot (Figure 10.7) and Histogram (Figure 10.8) of gold distribution within the 0.3 g/t wireframes supplied by Lydian. A review of the local data, using twinned drill hole data supplied by Lydian was also undertaken. Diamond drill hole data was composited to 1m intervals (or 2 m intervals in the case of holes DDA-042 vs. RCA-144 and DDA-045 vs. RCA-156 where RC data was collected in 2 m samples).



Figure 10.7 - Probability plot of RC and DD assay grade within Structural wireframes

Figure 10.8 - Histogram of RC and DD assay Data within Structural wireframes

CSA believes that there is no conclusive evidence, nor any suggestion, that RC assay data consistently understates or overstates gold grade as compared to diamond core assay data, on a global or local scale. There are no specific grade ranges that show difference. CSA conclude that one or other sample type does not introduce bias. When the twin hole data (RC V DDH) was reviewed and, accounting for shift where appropriate when comparing intervals between holes, do not see any significant understating of RC as compared to DDH. High grade intervals in RC are replicated in twin DDH and low grade too. Any observed differences, which we don't consider to be significant, are more than likely related to natural grade variation and the nugget effect. Given that there appears to be no observed bias with either DDH and RC data, and recent average RC sample recovery of 89 percent appears not to understate or overstate grade, CSA considers both data sets, exhibiting the average recovery stated above, to be acceptable for use in resource estimation work.



10.6 Sample Density Analysis Density determination at Amulsar has progressed through a number of phases. Density prior to 2009 was estimated using data from two studies, and a total of 55 samples. The first data set was comprised of 13 half core samples from 2 diamond drill holes representing a number of lithologies and alteration styles. SG was determined using hydrostatic weighing and waxing of the samples, testwork was undertaken by Geoid, an Armenian company specializing in mineralogical studies. Results of this testwork produced SG’s that ranged between 2.17 g/cm3 and 2.54 g/cm3, with an average of 2.38 g/cm3. The second study, undertaken prior to 2009, comprised 42 half core samples of a uniform 10cm length, collected at various depths from 13 holes. Samples covered a range of lithologies and alteration types encountered at Amulsar and improved spatial representivity. SG was determined by ALS Chemex using procedure OA-GRA08 which, depending on the nature of the sample utilised paraffin wax as a sealant. The specific gravities obtained from this work ranged from 1.96 g/cm3 to 2.68 g/cm3, with an average of 2.41 g/cm3. Testwork in 2009 was undertaken by Newmont on a total of 241 samples from a range of material types. Samples were selected from the 2008 and 2009 diamond core, and were grouped according to logged alteration and lithology. SG was determined using the paraffin wax sealant method, producing a weighted average SG for each alteration type. Lydian undertook a review of SG samples during 2010 as the geological model was developed, grouping them into dominant rock type; breccias, volcanics and porphyry andesite, and producing average SG values according to rock type, Table 10.4.

Table 10.4 SG Values According to Rock Type Rock Type Bulk Density (g/cm3)

Breccia 2.33 Volcanic 2.41

Porphyritic Andesite 2.24 CSA reviewed the SG data available and concluded;

▪ The data was not affected by any outliers (Figures 10.9, 10.10 and 10.11). ▪ There was no correlation between SG and gold grade (Figure 10.12). ▪ SG data collected adequately covers the resource area at Amulsar (Figure

10.13).



Figure 10.9 - Histogram of SG for Breccia Samples

Figure 10.10 - Histogram of SG for Volcanic Samples

Figure 10.11 - Histogram of SG for Prophyry Andesite Samples



Figure 10.12 - Correlation plot of SG vs Gold

Figure 10.13 - Plan View of SG samples

With 0.2 Wireframes demonstrating the coverage of sampling



SG values were assigned to the blocks within the mineral resource estimate according to rock type. Lydian has not collected any SG samples following the Newmont review in 2009. The current SG dataset is well spaced throughout the deposit and represents the dominant rock types, however when compared to the total sample population the data set is small. CSA recommend that Lydian continue to collect SG information at a 1:20 ratio during drilling to further refine the SG value for the deposit. 10.7 Sampling Method and Approach 10.7.1 RC Drilling RC samples are taken every one metre 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 (Figure 10.14). Samples are split to produce two 1.5 to 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 m 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.

Figure 10.14 - RC sample splitting at Amulsar



10.7.2 DD 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 m. Current practise dictates that core is pushed to one end of the core box and sample marks are made onto the core tray. CSA considers this practise to be inadequate, allowing for human error when sampling and would recommend that sample intervals be marked on core prior to cutting and sampling, with either a china graph pencil or white paint marker. Lydian systematically photographs all core prior to sampling. Historic logging has been of a high quality, however due to the nature of the mineralisation it is difficult to identify the two mineralisation styles from rock type alone. CSA would recommend that logging of percent porosity should be undertaken to assess whether this can be used as an indicator of mineralisation style along with structural measurements and may go some way towards identifying mineralisation types during logging and prior to the receipt of assay results.



11.0 SAMPLE PREPARATION, ANALYSIS AND SECURITY The following sample preparation, analysis and security descriptions are taken from the NI 43-101 report dated 19th May 2011 mentioned in the Introduction of this report. 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. In mid-September 2010 the crusher suffered a mechanical breakdown and was not in operation during CSA’s May 2011 visit. Lydian has ordered a new containerised sample preparation unit from ALS Chemex in Johannesburg which is due for delivery in June 2011 and will replace the existing unit. The new sample preparation laboratory will contain two crushers and a rotary splitter divider which will allow Lydian to reduce the sample size shipped for assay to 200 g to 250 g which will significantly reduce sample shipping costs. 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 2 mm. Samples are split using a riffle splitter and a 1,000 g split is then pulverised to 85 percent passing 75 µm. A 50 g 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 returning grades 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 prepares its samples and dispatches these to ALS Chemex with the inclusion of blanks, duplicates, and standards of different concentrations with random numbering.



12.0 DATA VERIFICATION The following data verification is taken from the NI 43-101 report dated 19th May 2011 mentioned in the Introduction of this report. 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 are directly related to massive clayey limonite zones, while wider, lower-grade sections relate 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 (Figure 12.1) were undertaken on the lower slopes of Amulsar. CSA considers Diamond drill practices to be of a high standard with a well organized and safe work environment, drilling was undertaken by well-trained competent drillers and offsiders. 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.



Figure 12.1 - Drilling being observed during the CSA site visit (2011)

In preparation for the forthcoming BFS, 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). 12.1 QA/QC Introduction Industry standard assay QA/QC control was implemented by Lydian during exploration and resource development activities in 2008, these practices are still current.



In addition to the QAQC data reviewed below, at the time of reporting Lydian were about to submit 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 0.2g/t gold and 1 in 40 (2.5%) for assays less than 0.2 g//t gold. A similar programme will be undertaken at the end of the 2011 drill programme. CSA has not had access to this data and as a result cannot comment on the representivity of the sample batch. 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.

Table 12.1 2010 Assay QA/QC Summary

During the 2010 RC DD sampling program a total of 3,646 QA/QC samples (comprising 2% blanks, 3% duplicates and 5% certified reference samples) were generated, which is considered representative of drilling and spatially representative. CSA believes the proportion of blanks and duplicates to be acceptable. 12.2 RC QA/QC Control Plots for RC QA/QC data collected during the 2010 drilling campaign is contained in Figures 12.2 to 12.10 below,



Figure 12.2 - RC Assay QA/QC - Blank Samples

Figure 12.3 - RC Assay QA/QC - Duplicate Samples



Figure 12.4 - RC Assay QA/QC – CRM G300-7 (1.00ppm)




Figure 12.8 - RC Assay QA/QC - CRM GLG 304-1 (0.15ppm), including Outliers

Figure 12.9 - RC Assay QA/QC - CRM GLG 304-1 (0.15ppm), zoomed View excluding Outliers




12.3 RC QA/QC - Conclusions Results from the analysis of blank reference samples shows less than 1 percent of samples (2/470) returning grades greater than 3× the detection limit (greater than 0.015 ppm 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 mineralized. CSA recommends that during future drilling campaigns the primary laboratory be audited to ensure adequate sample preparation practices 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). The scatter graph in Figure 12.3 shows that duplicate results exhibit acceptable levels of accuracy and precision. 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 percent, 100 percent, 98 percent, 100 percent, 99 percent, 98 percent 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 under-reporting 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 labeling 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 minimize 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 minimize these errors in future. Current blank samples are sourced from alluvial sand for RC samples. Although no significant contamination has been identified CSA would recommend that Lydian source certified blank material to align their procedures with industry best practice. 12.4 DDH QA/QC A total of 47 diamond holes have been drilled in 2010. Control Plots for DDH QA/QC data collected during 2010 is contained in Figures 12.11 to 12.19 below.

Figure 12.11 - DDH Assay QA/QC - Blank Samples



Figure 12.12 - DDH Assay QA/QC – Duplicate Samples

Figure 12.13 - DDH Assay QA/QC – CRM G 300-7 (1.00ppm)



Figure 12.14 - DDH Assay QA/QC – CRM G303-2 (4.15ppm)





Figure 12.17 - Assay QA/QC – CRM G904-8 (5.53ppm)





12.5 DDH QA/QC - Conclusions 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.



Adjacent samples were not mineralized. CSA recommends that during future drilling campaigns the primary laboratory be audited to ensure adequate sample preparation practices 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). The scatter graph in Figure 12.12 shows duplicate results to exhibit acceptable levels of accuracy and precision at grade ranges less than 2 ppm gold. At elevated grades, less levels of precision are observed. It should be noted that the presence of 7 outliers (7/193, or 4 percent of total population) does influence the correlation coefficient. 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 percent, 100 percent, 98 percent, 100 percent, 97 percent, 92 percent 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 labeling of the CRM and Blank samples, resulting in some confusion in values. CSA recommends that Lydian reviews systems currently in place that manage the insertion of CRM samples to minimize 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 plans to improve QA/QC monitoring during the next phase of drilling, which should address this. 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 CSA would recommend that



Lydian source certified blank material to align their procedures with industry best practice. 12.6 Laboratory 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 12.20 and 12.21. 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 12.20 - Diamond drilling Laboratory Duplicates



Figure 12.21 - 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. CSA would recommend that Lydian undertake a lab audit at the facility to review procedures and ensure that Lydian samples are being prepared and analyzed to industry best practice. 12.7 Geophysical Surveying Geophysical surveying was undertaken across the Amulsar prospect by Newmont during the now finished Joint Venture agreement with Lydian. Geophysical surveying during 2007/2008 was supervised by Mark Goldie, a Newmont geophysicist. Mr Goldie was also bought in during the interpretation and modeling work as a consultant for Condor Consulting during 2010. Approximately 150 line km of ground magnetic data was collected on a series of east west lines in 2006/2007 and 2009. Line spacing in 2006/2007 was 100 m increasing to 200 m during the 2009 survey. A total of 54.6 line km of induced polarization (IP) and resistivity (100m dipole-dipole) was completed over the Amulsar area during 2007 and 2008, Figure 12.22.



Figure 12.22 - Amulsar location map detailing magnetic and resistivity surveys

CSA reviewed two reports written by Mark Goldie, “Report on IP/Resistivity Surveys performed at the Amulsar Project, Armenia 2007 & 2008” and “Lydian Head Office Trip Summary: November 15 to 19, 2010”, contained in Appendix E of the May 19, 2011 43-101 Technical Report mention in the Introduction. These reports include the following conclusions; IP/Resistivity values, principally derived from the 2007 survey over the main area of interest, of several thousand to tens of thousands of ohm-m correspond well with the zones of known silicification, while the corresponding low chargeability reflects the depth of oxidation. Conversely, the 2008 survey area generally exhibits low resistivity, and high chargeability values in the order of 30 to 40 msec. This could represent an area of argillic alteration that contains 5 to 10 percent disseminated sulphides and may be related to a distal porphyry style system (Goldie 2008). The response of the magnetic data at Amulsar is clearly illustrating the contrast between the areas of known alteration (absence of magnetism) with the surrounding unaltered volcanics. Some lineaments can be seen in the derivative products that may represent structure (Goldie, 2010). The following recommendations were made as part of the 2010 report:

▪ Selected core and surface rock samples be collected and sent to a laboratory for physical property measurements. Susceptibility, IP, resistivity, and density information derived from this exercise will help the interpretation of the geophysical data sets and may help plan any future surveys. It is suggested



that a series of samples from each member of a suite of rock and alteration types be considered.

▪ Airborne magnetic/EM survey be considered to cover the Amulsar license. It

is recommended that a VTEM platform be used to fly approximately 1500 line km in an east-west direction with 100 m line spacing. The results of this survey could help in the prioritization of additional targets outside of the known area on known mineralization. The resistivity information could help identify silicification related to high sulphidation mineralization under the basalt cover, while the magnetics and resistivity could help identify other possible targets such as porphyries or base metals.

▪ Follow-up of any interesting anomalies based on the airborne survey should

be field checked and sampled with possible consideration for additional IP/resistivity to assist in the targeting. A conventional IP/resistivity approach will probably be sufficient to follow-up, but if there is a need to improve the resolution and depth of investigation of the original data sets, a Titan24 IP/resistivity approach should be considered

If there is interest in learning more about the underlying sediments and their possible relation to the mineralization at Amulsar, a ground gravity orientation reconnaissance survey could be considered at Amulsar. It is assumed that there will be a density contrast between the volcanics and sediments and that this boundary can be mapped with the gravity. Ideally the survey should extend out to any known outcropping sediments and some samples from this rock type should be sent out for physical property measurements. As a first pass, 300 to 400 stations could be collected along areas of vehicle access within the Amulsar license, such as roads and tracks, at approximately 500 m spacing. There are a number of geophysical contractors that could run this survey, and perhaps there is some capability within Armenia (i.e. through the government survey).

▪ An effort should be made to track down the raw IP/resistivity data files from Newmont (instrument dump files). These files contain important information related to the quality of the field measurements that may be of future use to Lydian.

▪ Although low priority, it may be worthwhile at some time to re-model the

IP/resistivity data in 2D with the recently acquired detailed topographic data. During 2010 Lydian staff and Mark Goldie created 3D geophysical models, generating point data with corresponding X, Y, Z locations. Lydian staff then used these 3D models to generate the following plots using a simple nearest neighbor estimation of values in Datamine, Figures 12.23 to 12.25.



Figure 12.23 - 3D Magnetic Data sliced by Rl



Figure 12.24 - 3D Resistivity Data sliced by Rl



Figure 12.25 - 3D Chargeability Data sliced by Rl

Lydian interprets the strong resistivity anomalies as being due to the silica alteration. There is a close relationship between the existing drilling and the resistivity anomalies which conforms this correlation. At deeper RL’s the resistivity identifies strong anomalies at Erato, to the north and east of Tigranes and to the south of Artavasdes. These areas have largely been untested by drilling and may also represent areas of silicification. The chargeability plots show small scattered highs at the higher RL levels which are interpreted as being associated with the pyrite rich porphyry andesite. At deeper RL levels a pronounced chargeability high occurs to the south and south west of Artavasdes. Lydian has embarked on an exploration drilling programme in this area which should identify the source of this chargeability anomaly. It is apparent that the bulk of the existing drilling is in areas of low chargeability and these areas are interpreted to represent oxidized silicified volcanics.



The 3D magnetics are dominated by gravity lows with small magnetic highs in the Tigranes, Artavasdes and Erato areas associated with areas of know mineralization. Sharp changes between magnetic highs and lows may represent fault structures. The 3D geophysics has been incorporated into Lydian’s drill planning for 2011.



13.0 MINERAL PROCESSING AND METALLURGICAL TESTING The following is a discussion of the metallurgical testing aspects of the proposed Amulsar Project. 13.1 SGS Lakefield Research (2008) In September 2008 a gold recovery test program was undertaken at SGS Lakefield in Canada on a 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-kg 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 S, S2- 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 Calc. Head Grade Au,

g/t

+150 mesh Fraction -150 mesh Fraction Gold Distribution, % Au, g/t

Au, g/t

Au, g/t

Au avg., g/t

+150 Mesh

-150 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



Table 13.2

Head Analyses of Composite 1 Element Units Comp 1 Element Units Comp 1

S % 0.08 S2- % 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 <20 Ba g/t 110 P g/t 94 Be g/t <0.1 Pb g/t 120 Bi g/t 68 Sb g/t 120 Ca g/t 390 Se g/t <30 Cd g/t <2 Sn g/t <25 Co g/t <4 Sr g/t 66 Cr g/t 22 Ti g/t 8,400 Cu g/t 98 Tl g/t <30 Fe g/t 37,000 U g/t <20 K g/t 620 V g/t 8 Li g/t <5 Y g/t 0.8

Mg g/t 330 Zn g/t 80 Mn g/t 9.6

13.1.2 Mineralogy A 1-kg sample of Composite 1 was submitted for qualitative mineralogical evaluation. The standard “rapid mineral scan” examination was applied, which includes general mineral assemblage with manual grain counting. An X-ray diffraction analysis was included for the identification of the major minerals. The rapid mineral scan evaluates the bulk petrography and general mineral types, abundances and grain size ranges. The rapid mineral scan indicated that the Composite 1 sample consisted of the following crystalline mineral assemblage phases:

▪ Major - quartz ▪ Minor - hematite

- goethite - potassium-feldspar - rutile

▪ Trace - chlorite - mica - alunite

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 was 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

Feed Size

Leach Final Residue g/t Au

Calc. Head Grade Au, g/t

Gold Recovery, % NaCN CaO Added

kg/t Cons. kg/t

Added kg/t

Cons. kg/t

1 day

3 days

7 days

10 days

15 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.

70

75

80

85

90

95

100

1 day 3 days 7 days 10 days 15 days

%Au Re

covery

Leach Time

Coarse Bottle Roll Leach TestsGold Leach Curves

1/2in. 1/4in. Figure 13.1 - Gold Leach Curves



13.1.4 Testwork 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 Testwork Results Summary

Liberation Size P80 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 NaCN 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 obtained for the two process routes, 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 NaCN 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 testwork, focusing on coarser fractions and lower cyanide concentration solution concentrations than previous testwork 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. The composites were differentiated by alteration style and by gold and multi-element 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.

Cyanidation bottle roll leach tests were conducted at two size distributions, minus

75 microns and minus 2 mm.

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 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,

(h) Gold Recovery, (%)

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, (d)

Gold Recovery, (%) 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,

d Gold Recovery, %


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.

0

10

20

30

40

50

60

70

80

90

100

0 20 40 60 80 100 120 140 160

%Au Re

covery

Leach Cycle (Days)

Column Leach Curves (‐38mm)

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,

d Gold Recovery, %


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.

0

10

20

30

40

50

60

70

80

90

100

0 10 20 30 40 50 60 70 80

%Au Re

covery

Leach Cycle (days)

Column Leach Curves (‐19mm)

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

Size and Test Gold Recovery, %

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 testwork 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.

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 testwork generally focused on leaching at finer crush sizes and using higher cyanide concentrations than were used in the SGS testwork. 13.3.1 Column Leach Tests Column testwork 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/m2/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, mm

NaCN Concentration, %

Gold Recovery, %

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



0.00%

10.00%

20.00%

30.00%

40.00%

50.00%

60.00%

70.00%

80.00%

90.00%

100.00%

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

% Au Re

covery

Leach Time (Days)

Column Leach Recovery Curves

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 sample the optimum crush size may be 19

mm and the optimum cyanide concentration 0.05 percent, although, further work will be required to substantiate this. 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 testwork. 13.4.1 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 minus 12 mm and 19 mm.



Results of bottle roll leach tests are shown in Table 13.11 for gold and Table 13.12 for silver.

Table 13.11Coarse Ore Bottle Roll Leach Test Results (Gold)

Deposit DDH# Rock Type Description Sample

(m)

Head Assay

Au (g/t) % Gold Recovery % Var

Au Rec

Avg. Deposit % Au Rec

-12 mm -19 mm -12 mm -19 mm

Erato DDAM 068

FeOx-banded volcanics, slight leaching 62.7-64.2 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 97.0 96.5

Artavasdes DDAM 070

Heavily leached + FeOx gossan 85.1-86.4 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 86.5 82.8

Tigranes DDAM 071

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 SM volcanics, numerous small-scale (<10mm) fracs, patchy FeOx 52.4-54.1 2.37 85.3 82.8 2.5 Fault zone with clay gouge, host silicified bx 88.8-89.8 2.72 98.3 96.1 2.2 SM volcanics, numerous small-scale (<10mm) fracs, patchy FeOx 108.0-109.0 0.59 92.2 90.1 2.1 Same as above (5) yet also faults with clay gouge 109.0-110.2 0.25 89.5 80.5 9.0 SM bx, patchy FeOx, no fractures/faults 110.2-112 90.7 89.2

Average 3.08 90.33 88.25 2.1

Table 13.12Coarse Ore Bottle Roll Leach Test Results (Silver)

Deposit DDH# Rock Type Description Sample

(m)

Head Assay

Ag (g/t) % Silver Recovery % Var

Ag Rec

Avg. Deposit % Ag Rec

-12 mm -19 mm -12 mm -19 mm

Erato DDAM 068

FeOx-banded volcanics, slight leaching 62.7-64.2 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 42.7 45.9

Artavasdes DDAM 070

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 44.0 36.3

Tigranes DDAM 071

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 SM volcanics, numerous small-scale (<10mm) fracs, patchy FeOx 52.4-54.1 1.60 41.3 57.5 -16.2 Fault zone with clay gouge, host silicified bx 88.8-89.8 3.70 57.6 61.1 -3.5 SM volcanics, numerous small-scale (<10mm) fracs, patchy FeOx 108.0-109.0 1.45 20.3 44.8 -24.5 Same as above (5) yet also faults with clay gouge 109.0-110.2 1.30 26.9 16.5 10.4 SM bx, patchy FeOx, no fractures/faults 110.2-112 42.8 48.7

Average 2.21 43.15 44.11 -1.0

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 at by crushing from minus 19 mm to minus 12 mm. The gold recoveries for the minus 12 mm tests are plotted against sample in Figure 13.5.



75

77.5

80

82.5

85

87.5

90

92.5

95

97.5

100

0 2 4 6 8 10 12 14

%Au Leach Re

covery

Sample

Coarse Bottle Roll Leach Tests (‐12mm)Amulsar Deposit Types

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 effect of head grade on 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.



75

77.5

80

82.5

85

87.5

90

92.5

95

97.5

100

0.00 2.00 4.00 6.00 8.00 10.00 12.00 14.00 16.00

%Au Leach Re

covery

Head Grade (g/tAu)

Coarse Bottle Roll Leach TestsEffect of Head Grade on Leach Recovery

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 than customarily used for gold. Future work will address silver extraction parameters. 13.4.2 Current Testing - Column Leach Tests At the time of writing this report drill hole intervals for each column leach test had been selected to prepare columns for each of the rock types:

▪ Leached and Iron Oxide Rich Gossan, Silicified ▪ Medium Iron Oxide, Siliceous ▪ Fault Gouge Zone ▪ Siliceous Breccia

Plus three columns to represent a master composite for each of the drill holes for:

▪ Tigranes ▪ Artavasdes ▪ Erato



Based on the coarse bottle roll leach tests the optimal crush size for the column leach tests to be carried out will be 100 percent -12 mm.

Particle size determinations conducted on each of the composite samples indicated that the percentage of minus 38 µm material was on average 5 percent by weight. Hence it was decided not to conduct any percolation/agglomeration tests ahead of the column leach tests. Percolation tests will be conducted once the columns have been terminated. 13.5 Future Testwork

As part of the 2011 field season there are plans to drill specific metallurgical diamond drill PQ holes. The purpose of these diamond drill holes will be to prepare composite samples that represent both depth and spatial across the Tigranes and Artavasdes open pit shells. Particular focuses will also emphasize on drilling within the starter pit to prepare metallurgical samples for column leach testing during the payback period i.e. first three years.

13.6 Heap Leach Design

The test results confirmed that the Amulsar gold mineralization is amenable to

recovery using heap leach technology. Conventional three-stage crushing, followed by stacking, leaching and gold recovery will be required. The design parameters for the leach stage are summarized in Table 13.13.

Table 13.13

Heap Leach Design Parameters Parameter Units Design Lift Height m 8 Solution Application Rate l/m2/h 8-10 Primary Leach Cycle days 30 Secondary Leach Cycle days 80 Solution:Ore Application kl/t 3.0

The design solution:ore application ratio of 3.0 kl/t was derived to attain an

overall gold leach recovery of 90 percent, as shown in Figure 13.7. The solution:ore application rate was derived from a column leach test conducted at a crush size of minus 12 mm and feed solution cyanide concentration of 0.5 g/l. These are the same parameters used for the current column leach tests being carried out by WAI.

In order to achieve a solution:ore application rate of 3.0 kl/t, a total leach cycle

time of 140 days is required at 10 l/m2/h. The design total leach cycle for a single lift is 110 days therefore the remaining gold will be leached in subsequent lifts. The primary leach cycle is designed to recover circa 70 percent of the gold, whilst an additional 10 percent will be recovered in the secondary leach cycle. The remaining 5 percent gold recovery will be leached in subsequent multiple lift leaching.



Laboratory column leach test recovery is downgraded by 5 percent to reflect the attainable full scale heap leach recovery due to losses on the side of the heap and due to channeling effects.

The main reason for limiting the leach cycle is to ensure that the pregnant leach

solution (PLS) grade to the adsorption circuit is maintained as high as possible. The current design PLS grade is circa 0.52 ppm gold. During the operations stage the operations staff will make the call when to terminate the leach cycle on each of the cells.

Figure 13.7 - Solution Application Rate vs. Leach Recovery

13.7 Metallurgical Recovery & Reagent Consumption Predictions

Initial column leaching was exceptionally rapid for both samples with up to 70 to

75 percent gold recovered after seven days of leaching; at crush size minus 12 mm. Initial column leach testing would indicate that circa 90 percent gold leach recovery is attainable after 140 days of leaching. For the PEA an overall gold recovery of 85 percent is used.

Column leach test recoveries are typically downgraded 2 to 5 percent units in the

full scale heap leach to take into account gold losses on the side of the heaps and due to inefficiencies arising in solution percolation within the heap.

0.00%

10.00%

20.00%

30.00%

40.00%

50.00%

60.00%

70.00%

80.00%

90.00%

100.00%

0 0.3 0.4 0.5 0.6 0.8 1.1 1.2 1.4 1.5 1.6 2.0 2.2 2.3 2.3 2.4 2.7 2.8 3.0 3.1 3.3 3.7 4.2 4.7 5.2 5.7 6.4

%Au Recovery

Solution Application Rate (m3/t)

Solution Application Rate12 mm 0.5 g/l

12mm 0.5gpl



Initial indications of cyanide and lime consumptions are that they will be low. Cyanide consumption is in the range 0.2 to 0.4 kg/t, whilst lime consumption is around 1 to 1.2 kg/t. Reagent consumptions in the laboratory are much higher than those obtained in the full scale heap, due to the recycling of leach liquor in a heap, as opposed to a static column which is in open cycle. The full scale plant reagent consumptions are typically half of those recorded in the laboratory.

These initial scoping testwork results suggest attractive processing economics of

the Amulsar project. Bulk mining of low-grade ore with a leach operation requiring three stage crushing is feasible.



14.0 MINERAL RESOURCE ESTIMATES

The text in Section 14.0 is a summary from the CSA Mineral Resource Estimate Report R210, dated 18 May 2011. Reference should be made to that report for the full details of the project geology and resource estimates.

During 2010 Lydian staff undertook a thorough review of the Amulsar geology,

including surface mapping, logging of oriented drill core and closer spaced infill drilling. This has resulted in the generation of a more robust geological model, where lower grade strata-bound mineralisation and higher grade structural bound mineralisation are defined.

The work program at Amulsar has been designed to define the scale of the

epithermal system and the extent and grade of the gold mineralization within that system. 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.

Work completed by Lydian since the production of the Mineral Resource

Estimate (MRE) in December 2009 comprised RC drilling, diamond drilling, trenching, channel sampling and sample assaying.

The mineral resource estimate comprised two sets of domain wireframes, the

first based on 0.2 to 0.3 sub-horizontal strata-bound materials the second based on 0.3 g/t Au and greater focused on mineralization that was structurally controlled. Domain wireframes were created by Lydian with input from CSA. The structural domains were then grouped according to their dominant strike orientation. A total of 12 domains were identified and used during mineral resource estimation.

The generation of distinct strata-bound and structural domains has greatly

reduced grade mixing that was highlighted as an issue during the previous resource estimates. Further refinement of these wireframes should reduce the minimal amount of grade mixing that is observed in the current domains.

Gold grade was estimated into a Block model with cells 20m x 20m x 5 m (X x Y

x Z), sub celled down to 2m x 2m x 0.5m (X x Y x Z). Bulk density was assigned according to dominant rock type.

The Amulsar Mineral Resource Estimate (MRE), dated 4 August 2011, has been

classified in accordance with CIM guidelines using the following criteria;

▪ Interpolation criteria based on sample density, variography, search and interpolation parameters.



▪ Assessment of the reliability and confidence that can be given to geological, sample, survey and bulk density data used in modelling. CSA have classified the Amulsar Mineral Resource Estimate as both Indicated

and Inferred Resources. As of 18 May 2011 the CIM compliant Indicated and Inferred Mineral Resource,

reported at a 0.4 g/t cut-off totalling 80.7 Mt @ 0.97 g/t gold for 2.52 million ounces, of this total 32.4 Mt @ 1.1 g/t gold for 1.1 million ounces is contained within the Indicated Resource and 48.3 Mt @ 0.9 g/t gold for 1.4 million ounces is contained within the Inferred Resource was delineated.

Table 14.1

Mineral Resource Estimate at a 0.4 g/t Gold Cut-off Grade

Resource Classification Million Tons Au g/t

Au Mozs

Density g/cm3

Indicated 32.38 1.07 1.11 2.37 Inferred 48.34 0.91 1.41 2.37

Total 80.72 0.97 2.52 2.37 1. The CSA Mineral Resource was estimated by construction of a block model within constraining

wireframes. 2. The resource is reported at lower cut-off grade of 0.4 g/t Au, which defines the continuous/semi-

continuous mineralised zone potentially amenable to the low grade, bulk tonnage mining scenario currently being considered by Lydian.

3. Apparent differences may occur due to rounding errors. The resource reported has been rounded to reflect the fact that it is an approximation.

4. Mineral resources are reported in accordance with the CIM Code. 5. Mineral resources which are not mineral reserves do not have demonstrated economic viability. The

estimate of mineral resources may be materially affected by environmental, permitting, legal, marketing, or other relevant issues.



15.0 MINERAL RESERVE ESTIMATES

The mine PMM estimates, pit optimization study and mining costs have been investigated by CSA Global (UK) and are described in detail in the CSA report number R268.2011 prepared for Lydian in July 2011. The following is a summary from this report.

In April 2011, Lydian International Pty (“Lydian”) commissioned CSA Global (UK)

Ltd (“CSA”) to undertake the mining portion of a Preliminary Economic Assessment (PEA) review of the available information relating to the Amulsar Gold Project, located in Armenia, with the objective of developing a concept for the successful development of the project.

The initial stage of this work was to carry out pit optimisation work based on

physical and financial parameters agreed between CSA and Lydian. CSA has not carried out independent review of these parameters at this initial stage of the study and has relied on the information provided by Lydian.

CSA completed most of this preliminary optimisation work during April and May

2011. The purpose of this work was to test a series of options involving combinations of Inferred and Indicated resources and pit wall slopes. All other physical and economic parameters for each scenario remained unchanged. In particular, the initial optimisation work was carried out to test whether the resources which are currently in the Inferred Category would be economically mineable if further drilling were to be carried out to bring them to the Indicated Category.

The outcomes of the optimisation runs have been combined during June/July 2011

into an economic analysis to provide a preliminary indication of the viability of the project and generate sensitivity studies to changes in key economic and operational parameters particularly the option of a contractor mining verses owner mining operation. Study Approach

The following work was undertaken:

▪ Preliminary estimation of operating costs for Amulsar gold deposit

▪ Pit optimizations to determine practical pit limits, whilst maximizing the project value, using Whittle Four-X software

▪ Sensitivity analysis to determine the influence of changes in various factors on

the project ▪ Pit designs on selected optimisation shells ▪ Reporting of inventories within the Amulsar pit designs ▪ Determination of waste dump volume requirements and waste dump designs



▪ Generation of mining schedules using the open pit inventory information ▪ Generation of Scoping Study level operating and capital cost estimates

The gold resource used in the Whittle optimiser is detailed in Table 15.1. It

should be noted that a lower cut-off grade of 0.2 g/t compared to the computed resources at a cut-off grade of 0.4 g/t was used.

Table 15.1 Amulsar Gold Deposit – Au Resources Input to Whittle (0.2g/t Au Cut-off Grade)

Area Mineralization Style 4X CODE Classification

TotalTonnes

(Mt)

Total Au Content

(kg)

AuGrade (g/t)

Artavasdes Structurally Controlled AHO Indicated 19.46 18,352.22 0.94 Artavasdes Strata Bound ALO Indicated 1.82 623.12 0.34 Tigranes Structurally Controlled THO Indicated 18.60 17,894.28 0.96 Tigranes Strata Bound TLO Indicated 1.06 269.75 0.25 Total Indicated 40.95 37,139.37 0.91 Artavasdes Structurally Controlled AHI Inferred 24.48 19,445.88 0.79 Artavasdes Strata Bound ALI Inferred 1.98 563.69 0.28 Tigranes Structurally Controlled THI Inferred 14.08 10,893.01 0.77 Tigranes Strata Bound TLI Inferred 2.38 723.09 0.30 Total Inferred 42.92 31,625.66 0.74 Total Indicated and Inferred 83.86 68,765.04 0.82

1. The CSA Mineral Resource was estimated by construction of a block model within constraining wireframes. 2. The resource stated above is reported at lower cut‐off grade of 0.2g/t Au. 3. Apparent differences may occur due to rounding errors. The resource reported has been rounded to reflect the fact that it is

an approximation. 4. Mineral resources are reported in accordance with the CIM Code. 5. Mineral resources which are not mineral reserves do not have demonstrated economic viability. The estimate of mineral

resources may be materially affected by environmental, permitting, legal, marketing, or other relevant issues.



‐

5.00

10.00

15.00

20.00

25.00

‐

500,000

1,000,000

1,500,000

2,000,000

2,500,000

0 20 40 60 80 100 120 140

Maxim

um Parcel Au Grade

g/t

Resource ton

nes

Z Block No

Amulsar ResourceDepth Distribution

Ind tonne Ind‐Inf tonne Ind Max Au g/t Ind‐Inf Max Au g/t

Option 1 (Indicated) Optimal Pit has a basal block at 72

Option 2 (Ind‐Inf) Optimal Pit has a basal block at 33

Figure 15.1 - Amulsar Resource – Depth Distribution

Figure 15.1 above provides a guide to the level of influence on the depth of the pit optimisation of anomalously high grades. In neither of the runs does this appear to be a major factor. Whittle Optimization Input Parameters The following four options were optimised:

▪ Option 1 - Indicated resources only, pit slopes 40o ▪ Option 2 - Indicated and Inferred resources, pit slopes 40o ▪ Option 3 - Indicated resources only, pit slopes 45o ▪ Option 4 - Indicated and Inferred resources, pit slopes 45o



Table 15.2 Amulsar Au Deposit Whittle Optimisation Input Parameters

Input Parameter Amulsar Au Deposit Whittle Optimisation Input Parameters

Options 1 and 2 Options 3 and 4 Mining cost ($/t mined) Ore - $2.10/t Waste - $2.00/t Ore - $2.10/t Waste - $2.00/t Processing/ore costs ($/t milled) Including all Fixed costs

5Mtpa (HL) - $7.30/t

10Mtpa (HL) - $5.97/t

5Mtpa (HL) - $7.30/t

10Mtpa (HL) - $5.97/t

Overall open pit slope 40o 45o Metal Price US $ 1,000/oz US $ 1,000/oz Exchange Rate Royalties 1.5% 1.5% Sell Cost $144.68/kg Au $144.68/kg Au Plant capacity 5Mtpa 10Mtpa 5Mtpa 10Mtpa Mining Limit No Limit No Limit Mining recovery 95% 95% Mining Dilution 5% 5% Process recovery 80% 80% Discount Rate 5% 5%

Whittle Optimization Results

A balance of DCF and maximum PMM tonnes was used to select the preferred pit shells. Options 3 and 4 were immediately excluded, as a decision was made by Lydian to use the flatter pit slope angles due to a lack of geotechnical data. The optimal pit shell #36 was selected from the Option 1 optimisation. Cut back positions have been selected at Pits #7 and #11. The optimal pit shell #36 was also selected from the Option 2 optimisation. Cut back positions have been selected at Pits #8 and #23.

The decision was made to produce a pit design for Amulsar on the chosen optimal

pit shell using pit shell #36 based on the Option 2 Whittle optimisation.



Table 15.3

Optimal Pit Shell Physical Results for all Options Amulsar Au Deposit

Optimisation Option

Pit Shell #

PMM in the ground

PMM processed (incl. dil. and ore loss)

Waste tonnes

Stripratio

Gold to Heap Leach

RecoveredGold

AveragePlant

Recovery(Mt) Au (g/t) (Mt) Au (g/t) (Mt) t/t (%) Au (kg) (%)

Option 1 36 31.1 1.03 31.0 0.98 75.9 2.45 30,329.3 24,262.7 80 Option 2 36 53.8 0.99 53.6 0.94 138.0 2.57 50,528.4 40,421.5 80 Option 3 36 32.5 1.01 32.4 0.96 73.1 2.25 31,182.3 24,946.5 80 Option 4 36 54.5 0.98 54.3 0.94 131.0 2.41 50,909.3 40,727.5 80

Table 15.4 Optimal Pit Shell Financial Results for all Options

Amulsar Au Deposit

Optimisation Option

Cost of Mining

Cost of Processing

Revenuefrom Au

Cash flow(Undisc.)

Discounted Cash flow LOM Best Specified Worst $M $M $M $M $M $M $M years

Option 1 213.8 208.0 764.8 343.0 267.8 263.8 242.8 4.60 Option 2 383.2 345.4 1,274.1 545.5 391.7 367.1 333.9 6.86 Option 3 211.0 196.9 786.3 378.4 347.8 337.4 337.4 3.24 Option 4 370.7 329.9 1,283.7 583.2 520.6 486.0 486.0 5.43

The level of confidence in the technical studies complete to date would not support

the presentation of a Reserves Statement for the Amulsar Project. All resources captured by the optimisations cannot be converted to Reserves, but can be referred to as Potentially Mineable Mineralisation (“PMM”).

The percentage of resource to PMM conversion is favourable for both the indicated resources only option and the indicated and inferred resources option as illustrated in Table 15.5 below.

Table 15.5

Conversion of Resources to PMM (In-situ Figures)

Indicated Resources Only Indicated and Inferred Resources Mineralised

tonnes (Mt)

In-situ Au content

(kg) Grade (g/t)

Mineralised tonnes

(Mt)

In-situ Au content Grade

(kg) (g/t) Resource 40.95 37,139 0.91 83.86 68,765 0.82 PMM (Insitu) 31.05 31.92 1.03 53.75 53,187 0.99 Percent Conversion 75.83 85.96 64.09 77.35 The surface extent of the final pit shells for Option 1 (Pit #36) and Option 2 (Pit #36) are illustrated in Figure 15.2 below.



Figure 15.2 - Amulsar Selected Optimal Whittle Pit Shells (#36)

Options 1 and 2 Sensitivity Analysis The following sensitivities were analysed for the Amulsar Option 1 and Option 2

Optimizations. Sensitivities of -10% and +10% were analysed for the Amulsar Option 1 and Option 2 Optimisations on:

▪ gold price ▪ discount rate ▪ mining dilution ▪ mining recovery ▪ mining costs ▪ processing costs ▪ process recoveries



A further sensitivity analysis examined the effect of adding the inferred material in Erato area to the Option 2 resource to illustrate the potential upside that the addition of this portion of the resource would have on this project should future exploration convert these resources to the measured or indicated category.

The studies indicated that the project is fairly insensitive to fluctuations in process

recoveries of the strata bound mineralization. However, it is sensitive to fluctuations in gold price, mining recovery and process recoveries of the structurally controlled mineralization.

Figure 15.3 and Figure 15.4 illustrate the results of the sensitivity analysis. Note that all sensitivities are based on the undiscounted cash flow.

Figure 15.3 - Whittle Optimization Sensitivity Spider Graph (Cash Flow) – Option 1



Figure 15.4 - Whittle Optimization Sensitivity Spider Graph (Cash Flow) – Option 2

Table 15.6 Design vs. Whittle Shell Comparison

Description Ore Kt

Grade Au g/t

Au Content kg

Waste kt

Strip Ratio t:t

Optimisation Shell #36 53,617 0.94 50,400 138,002 2.57 Pit Design Shell 55,268 0.92 51,030 135,969 2.46 Variance (%) 3.08 -1.77 1.25 -1.47 -4.42



16.0 MINING METHODS

The following is a summary from CSA Report No R268.2011 submitted to Lydian in July 2011.

16.1 Pit Design Parameters

In designing the final pit for Amulsar, the following criteria have been taken into

account.

• Statutory and safety measures • Policy decisions • Equipment dimensions

Design ore production rates for Amulsar start at 3.75 Mtpa in year 1 ramping up to

10 Mtpa in year 4. From the optimisation studies, the following pit shells have been chosen on which

to base the pit designs. A pit design was only completed on the final pit shell.

Table 16.1 Optimal Shells for Pit Design

Deposit Optimisation Run Pushback Final DesignAmulsar Indicated and Inferred resources excluding Erato Pit Shell #8 and #23 Pit Shell #36

Table 16.2 details the geotechnical parameters, as agreed by Lydian and CSA,

used to design the Amulsar pit. Figure 16.1 illustrates the terminology used in pit wall design.

Table 16.2 Pit Design Parameters

Batter Angle (degrees)

Batter Height (m)

Berm Width (m)

Ramp Grade (1 : x)

Ramp Width (m)

65 10 5 10 25

Features of the pit design include:

▪ 25m wide ramps to allow for the safe passing of the selected haulage trucks with an allowance for a bund wall on the open side of the ramp.

▪ A ramp gradient which allows the selected haulage trucks to maintain maximum

production.

▪ A minimum pit base width of 20m to allow for load and haul operations.



Figure 16.1 - Pit Design Terminology

16.1.1 Pit Designs

Figure 16.2 and Figure 16.3 below are plan and oblique views of the pit design of Amulsar based on the Amulsar Indicated and Inferred optimisation. The ramp daylights on the north eastern edge of the pit, close to the areas allocated for waste dumping. The final pit utilises a clockwise spiral ramp from the pit exit to access 2,688 mRL in the southern end of the pit with a switchback at 2,758 mRL to access 2,728 mRL in the northern end of the pit.

Due to the mountainous topography, the area of the waste dump is expansive

and straddles multiple valleys extending some distance from the pit, thus increasing haul distances. The open pit is also expansive with a North-South length of some 1.3 km which also contributes to increased haulage distances.

Table 16.3 below summarise the total rock tonnes, insitu and diluted ore tonnes

and grades, waste tonnes and insitu and diluted gold content within each of the mining areas of the Amulsar pit design reported by 10 metre bench.



Figure 16.2 - Amulsar Pit Design - Plan View



Figure 16.3 - Amulsar Pit Design – Oblique View Looking North



Table 16.3

Amulsar Pit Design PMM Tonnes and Grade Inventory

BENCH

Total Tonnes

(kt)

Insitu PMM

Tonnes (kt)

Insitu PMM Au

Grade (g/t)

Insitu PMM Au

Content (kg)

Heap LeachFeed PMM

Tonnes (Including 5% Loss and 5%

Dilution) (kt)

Waste Tonnes

(kt)

Strip Ratio (t/t)

Heap LeachFeed PMM Au

Content (Including 5% Loss and 5%

Dilution) (kg)

Head Grade

Au (g/t)

2980-2990 20.3 0.3 0.60 0.2 0.3 20.0 65.04 0.2 0.57

2970-2980 211.1 64.5 0.61 39.3 64.3 146.8 2.28 37.3 0.58

2960-2970 889.2 354.2 0.74 260.6 353.3 535.8 1.52 247.6 0.70

2950-2960 2,121.5 814.1 0.73 594.1 812.0 1,309.4 1.61 564.4 0.69

2940-2950 3,581.2 1,199.7 0.70 844.1 1,196.7 2,384.6 1.99 801.9 0.67

2930-2940 4,932.5 1,350.4 0.70 940.6 1,347.0 3,585.5 2.66 893.6 0.66

2920-2930 6,993.8 1,891.0 0.80 1,521.8 1,886.2 5,107.5 2.71 1445.7 0.77

2910-2920 8,561.3 2,339.2 0.85 1,997.8 2,333.3 6,228.0 2.67 1898.0 0.81

2900-2910 9,616.9 2,633.8 0.86 2,278.1 2,627.3 6,989.6 2.66 2164.2 0.82

2890-2900 10,536.2 2,863.5 0.93 2,656.4 2,856.3 7,679.9 2.69 2523.6 0.88

2880-2890 11,418.3 3,034.6 0.94 2,842.5 3,027.0 8,391.3 2.77 2700.4 0.89

2870-2880 11,848.6 3,061.5 0.91 2,789.7 3,053.9 8,794.7 2.88 2650.3 0.87

2860-2870 11,881.6 3,056.3 0.92 2,798.7 3,048.7 8,832.9 2.90 2658.7 0.87

2850-2860 12,040.3 3,164.2 0.90 2,861.0 3,156.3 8,884.0 2.81 2717.9 0.86

2840-2850 11,816.8 3,063.4 0.94 2,890.9 3,055.7 8,761.0 2.87 2746.4 0.90

2830-2840 11,621.6 3,131.7 0.93 2,913.0 3,123.9 8,497.7 2.72 2767.3 0.89

2820-2830 11,327.2 3,082.8 1.01 3,115.1 3,075.1 8,252.1 2.68 2959.4 0.96

2810-2820 10,867.7 3,023.4 1.07 3,229.5 3,015.8 7,851.9 2.60 3068.1 1.02

2800-2810 9,801.6 2,832.4 1.19 3,359.7 2,825.3 6,976.2 2.47 3191.7 1.13

2790-2800 8,645.9 2,661.0 1.14 3,028.7 2,654.4 5,991.5 2.26 2877.3 1.08

2780-2790 7,355.7 2,297.5 1.23 2,824.9 2,291.8 5,064.0 2.21 2683.6 1.17

2770-2780 6,352.8 2,095.1 1.19 2,500.3 2,089.9 4,263.0 2.04 2375.3 1.14

2760-2770 5,536.0 1,909.7 0.99 1,884.8 1,905.0 3,631.0 1.91 1790.6 0.94

2750-2760 4,523.7 1,666.8 1.00 1,674.7 1,662.6 2,861.1 1.72 1591.0 0.96

2740-2750 3,426.0 1,379.0 1.07 1,475.5 1,375.5 2,050.5 1.49 1401.8 1.02

2730-2740 2,348.7 1,084.5 1.08 1,171.9 1,081.8 1,266.9 1.17 1113.3 1.03

2720-2730 1,526.0 711.2 0.98 697.7 709.5 816.5 1.15 662.8 0.93

2710-2720 807.1 343.2 0.78 268.3 342.3 464.8 1.36 254.8 0.74

2700-2710 454.3 209.8 0.83 173.7 209.3 245.0 1.17 165.0 0.79

2690-2700 163.6 86.6 0.95 82.1 86.4 77.2 0.89 78.0 0.90

2680-2690 8.8 0.7 0.60 0.4 0.7 8.1 11.71 0.4 0.57 Total 191,236.3 55,406.3 0.97 53,716.2 55,267.8 135,968.5 2.46 51030.4 0.92



16.2 Waste Dump Location and Design

The waste dump for Amulsar is located to the east and northeast of the main pit, in the least visible location from the majority of the communities. The average from the pit was assumed to be 1 km. The viability of in-pit dumping will be investigated as part of the BFS to potentially reduce the footprint of the surface dump, if possible.

Due to the mountainous topography, the area of the waste dump is expansive

and straddles multiple valleys extending some distance from the pit, thus increasing haul distances. This area has been nominated by Lydian as the area least likely to be mineralised in the vicinity of the pit.

If necessary, mine waste will be segregated and/or appropriately managed to

mitigate any potential Acid Rock Drainage (ARD) in accordance with the results of the ESIA ARD assessment programme. The design of the waste dump will incorporate appropriate drainage and environmental protection measures, as necessary.

Based on a swell factor of 35%, the total volume required for waste dumping for

Amulsar is 78.6 Mlcm. The swell factor has been assumed by CSA for this study and requires confirmation by future testwork. Table 16.4 provides a breakdown of the waste dump volume requirement.

Table 16.4 Amulsar Waste Dump Volume Requirements

Description Waste Production (kt)

Insitu Bulk Density g/cm3

Volume Requirement kbcm klcm

Amulsar Au Deposit 135,969 2.34 58,209 78,581

Table 16.5 details the general criteria that were used for the waste dump designs.

Table 16.5 Amulsar Waste Dump Design Parameters Swell Factor 35%

Overall Slope Angle Average 24o Batter Angle Average 30o Berm Width 10m

Bench Height 20 m Ramp Width 25 m

Ramp Gradient 1:10



Table 16.6 and Figure 16.4 provide details of the waste dump design.

Table 16.6 Amulsar Waste Dump Design Volumes

Interval Waste Dump Volume, klcm 2780 - 2800 2,160.4 2760 - 2780 1,959.6 2740 - 2760 1,810.3 2720 - 2740 1,770.4 2700 - 2720 1,845.2 2680 - 2700 1,968.0 2660 - 2680 2,071.6 2640 - 2660 2,108.0 2620 - 2640 2,162.3 2600 - 2620 2,232.8 2580 - 2600 2,200.2 2560 - 2580 2,060.0 2540 - 2560 2,010.0 2520 - 2540 1,941.1 2500 - 2520 1,856.3 2480 - 2500 9,560.9 2460 - 2480 8,787.9 2440 - 2460 7,785.7 2420 - 2440 6,693.1 2400 - 2420 5,670.9 2380 - 2400 4,520.5 2360 - 2380 3,315.4 2340 - 2360 2,018.8 2320 - 2340 862.0 2300 - 2320 171.5 2280 - 2300 4.4

Total Volume 79,547.5



Figure 16.4 - Amulsar Waste Dump Location and Design



16.3 Mining Methodology 16.3.1 Introduction

This study investigates two mining options at Amulsar, an owner operator mining operation and a contractor mining operation. All mining costs have been derived from typical mining costs for an operation of similar type and size.

The mining operation has been based on the following assumptions; ▪ Surface mining only

▪ Drilling and blasting of waste on 10 m benches

▪ Bulk mining of ore from the high grade structurally controlled and low grade

strata-bound ore resources

▪ A relatively small pool of local skilled and semi-skilled labour

▪ A large fleet of 130 t and 90 t rear dump mechanical drive haul trucks selected since they are in current use or on order by mining companies in Armenia and are supported by local dealers

16.3.2 Mining

The mining method assumed for the Amulsar deposit considered in this study will be conventional open pit mining, with drilling and blasting of waste on 10m benches and drilling and blasting of ore on 5m or 10m benches, depending on the ore demarcation required for grade control purposes. Load and haul will be done on flitches varying between 5m and 10m in height. The general process of loading ore and waste will use backhoe or shovel type excavators with the excavator sitting on the production bench while the haulage trucks are loaded on the level below.

The mining sequence will include removal of waste to dumps and the mining of

ore and its delivery either to the specified ROM pads or direct tipping into the conveyor bin in close proximity to the open pit. Ore from the ROM will then be transported from the pit to the Heap Leach (HL) plant by conveyor.

Ore extraction will be by a bulk mining method. The cost estimates include an allowance for grade control as there will be a requirement to mine ore and waste with a minimum of dilution and maximise recovery.

Standard ancillary equipment will be used where necessary for smooth graded floors and face clean-ups.



16.4 Open Pit Production Equipment 16.4.1 Selection Criteria

Mining operations have been assumed to work two 12 hour shifts per day six days per week. Mining equipment selection is based on optimal utilisation of the selected equipment. 16.4.2 Drilling

Drill rigs such as Sandvik Pantera DP1500 drills capable of drilling holes with diameter between 89 mm and 150 mm and Tamrock Ranger DX 700 drills capable of drilling holes with diameters between 64 mm and 115 mm would be sufficient for this operation. These one man operated diesel hydraulic drills have a stated capacity of up to 15 m per hour in hard strata.

Up to five drill rigs will be required to meet peak production for in pit ore and waste grade control and blast hole drilling at Amulsar, including allowance for downtime. 16.4.3 Loading Equipment

Loading equipment planned for Amulsar consists of 180 t hydraulic shovels such as an RH-90 class primarily for loading of ore and 250 t hydraulic shovels such as an RH-120 class for loading of waste.

Up to three RH-90 and two RH-120 hydraulic shovels will be required to meet peak production for in pit ore and waste loading at Amulsar, including allowance for downtime.

Provision has been made in this study for wheel loaders to load approximately 40 percent of ROM ore onto the conveyor for transportation to the HL plant, with the remainder being direct tipped by haul truck into the conveyor bin. 16.4.4 In Pit Haulage Trucks

A fleet of Cat 777 90-t haul trucks is planned primarily for ore transportation and a fleet of Cat 785 130-t haul trucks is planned primarily for waste transportation. Maximum truck size is constrained by equipment availability in Armenia.

Up to twenty one 777 haul trucks and sixteen 785 haul trucks will be required to meet peak production for in pit ore and waste haulage at Amulsar, including allowance for downtime. Haul distances from pit base to ramp exit and surface haulage to RoM and waste dump will be up to 2 km one way, requiring the large trucking fleets.



16.4.5 Ancillary Equipment

Ancillary equipment will include the following: ▪ Graders ▪ Dozers ▪ Service trucks ▪ Explosives transport ▪ General Purpose trucks ▪ Personnel transport ▪ Light vehicles ▪ ROM loader

16.5 Amulsar Production Schedule

A Life of Mine (LoM) mining schedule has been developed for the Amulsar open pit. The aim of the LoM schedule was to find a balance between mining high grade material as early as possible in order to get maximum returns, whilst minimising the use of stockpiles in order to keep re-handling costs to a minimum. Stockpiles were used in the schedule to smooth out annual PMM and waste mining. Strip ratios increase in later years in the schedule which mean that waste mining has to increase to maintain Heap Leach Feed. To mitigate this effect, stockpiles are created to supplement Heap Leach feed when strip ratios increase.

Table 16.7

Amulsar Production Schedule

Year

Mining

Total PMM

Mined (Mt)

Total PMM Stockpile

Start (Mt)

Total PMM Stockpile

End (Mt)

Total PMM Input to

Heap Leach (Mt)

Grade Input to

Heap Leach (g/t)

Total Waste Mined (Mt)

Strip Ratio (t:t)

1 4.16 0.00 0.41 3.75 1.16 5.00 1.20 2 5.03 0.41 0.44 5.00 0.86 10.00 1.99 3 7.42 0.44 0.36 7.50 0.87 17.31 2.33 4 10.03 0.36 0.39 10.00 1.00 27.19 2.71 5 10.00 0.39 0.39 10.00 0.90 27.27 2.73 6 9.64 0.39 0.02 10.00 0.97 26.59 2.76 7 8.99 0.02 0.00 9.02 0.79 22.61 2.51

Total 55.3 55.3 0.92 136.0 2.46



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 for 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 3 are the following Flowsheets for Phase I and Phase II as well as a General Arrangement drawing. A proposed overall site plant is presented in Section 18. 10-F-01 Flowsheet Primary Crushing 13-F-01 Flowsheet Secondary Crushing 15-F-01 Flowsheet Tertiary Crushing 17-F-01 Flowsheet Lime Addition

19-F-01 Flowsheet Ore Stacking 20-F-01 Flowsheet Heap Leach 23-F-01 Flowsheet Carbon Adsorption 25-F-01 Flowsheet Stripping and Refining 27-F-01 Flowsheet Carbon Reactivation 30-F-01 Flowsheet Utilities and Reagents 00-L-001 Site Plan Overall Project Area The Overall Flowsheet is shown below in Figure 17.1.



Figure 17.1 Amulsar Overall Flowsheet



17.1 Crushing Plant and Stacking Ore is processed through three stages of crushing. 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 is 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. 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 recirculates 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 a second tertiary vibrating screen and belt feeder and two additional tertiary cone crushers with belt feeders. 17.1.4 Stacking The overland conveyor brings the ore to the vicinity of the leach pad and runs along the north side of the pad. Pebble lime is added to the ore on the overland conveyor by way of a storage silo and screw feeders with the rate of lime addition varying with tonnage. The ore with lime is discharged onto a network of up to 8 portable conveyors terminating at a radial stacking conveyor. Ore is discharged from the stacking conveyor onto the heap leach pad to an ore height of 8 meters. The overland conveyor, lime silo, portable conveyors and stacking conveyor are all sized and installed in Phase I to accommodate the increase in throughput required in Phase II, although, an additional 16 portable conveyors will ultimately be purchased as the overall height of the pad increases. 17.2 Heap Leach

Golder completed and submitted to Lydian a separate document detailing a conceptual design and cost estimate for the leach pad and ponds. Heap leaching consists of stacking the crushed ore on the leach pad in lifts and leaching each individual lift to extract the gold. Barren leach solution (BLS) containing approximately 1 g/l sodium cyanide is applied to the ore heap surface using drippers at an application rate of 10 l/h/m2. 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 m3/t 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 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 (RLS) 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 at Site 13 will be constructed in three phases with the ultimate ore heap amount of 75 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 in the graded valley bottom and rise up the surrounding valley slopes enough height to accommodate the Stage 1 heap of approximately 13.4 Mt, which will be stacked on the Phase 1 pad in six 8 m thick lifts during the first three years of operation. The Phase 1 pad will have an area of 347,430 m2 and the Stage 1 heap will have a top surface elevation of 2,396 m.

The Phase 2 pad will be an expansion to the Phase 1 pad up the valley slopes and will have an area of 431,100 m2. The Phase 2 pad will allow the stacking of the Stage 2 heap of approximately 28.9 Mt in four additional lifts above the Stage 1 heap for an approximate cumulative heap amount of 42.3 Mt with a heap top surface elevation of 2,428 m. Stacking of the Stage 2 heap will continue through Year 6 of Operation.

The Phase 3 pad will again be an expansion to the Phase 2 pad up the valley slopes to the Ultimate pad limits. The Phase 3 pad will have an area of 265,630 m2, and the Ultimate (Phases 1, 2 and 3) pad area will be 1,044,160 m2. The Phase 3 pad will allow the stacking of the Stage 3 heap of approximately 32.7 Mt in six additional lifts above the Stage 2 heap for an Ultimate heap amount of 75.0 Mt with a heap top surface elevation of 2,476 m. Stacking of the Stage 3 heap will continue through Year 9 of Operation.

If additional leachable ore resources are identified, the ore may be stacked higher

on Site 13 and/or on a leach pad constructed in the western portion of Site 11 and solution from this pad would be piped to the Site 13 collection ponds.

The pad will have a composite liner system consisting of 2.0 mm LLDPE geomembrane underlain by 0.3-m minimum compacted thickness of low-permeability cohesive soil layer. The geomembrane will be smooth in most areas and may be double-side textured in a strip along the downgradient toe of the pad to enhance heap stability. An interlift liner of similar design may be constructed beneath the Stage 3 heap if deemed necessary due to insufficient percolation of the 80 m high leached ore of Stages 1 and 2 heap, to be determined in the BFS.

The drain pipe network above the pad liner will be embedded within a 0.6-m minimum loose lift thickness liner cover drain fill composed of free-draining, hard, and durable granular material. Solution and storm runoff flows collected by the drain pipe network in each pad cell will be routed via header pipes through the pad spillways to the process ponds, and will be directed by valve control to either the pregnant or intermediate ponds.



17.2.2 Collection Ponds

The collection ponds will consist of process (pregnant and intermediate) ponds, a storm pond, and an underdrain collection and monitoring pond. The collection ponds will be constructed during the Phase 1 pad construction east of the pad downgradient toe in the lowest portion of the Site 13 valley. The ponds will have a crest elevation approximately 8 m higher than the lowest existing ground surface to match the elevation of the pad spillways to the process ponds and minimize site grading cuts. The ponds layout generally matches the terrain in the area to minimize overall site grading.

Solution and storm runoff flows from the leach pad cells will be routed to either the pregnant or the intermediate ponds. A common divider berm will be constructed between the pregnant and intermediate ponds for solution and storm runoff overflow from the pregnant pond to the intermediate pond. A spillway will be constructed between the intermediate pond and the storm pond for storm runoff overflow from the intermediate pond to the storm pond.

The process ponds will accommodate the solution operational and draindown

storage requirements for the Ultimate pad, and the storm pond will accommodate the design storm runoff from the Ultimate pad and pond areas. The required size of the process ponds will double after the first three years of operation as the solution flow rate to the pad is doubled in Year 4 with the doubling of the ore processing tonnage.

The underdrain collection and monitoring pond will collect flows from the underdrains constructed beneath the leach pad and the process and storm ponds. Collected flows meeting acceptable water quality requirements will be discharged into the natural drainage. Collected flows not meeting the requirements will be pumped to the pregnant pond and/or the process plant circuit. The underdrain collection pond will be sized in the BFS to store the anticipated flows with an appropriate safety factor. The pond size shown on Drawing 2 is based on experience and should suffice for the PEA Conceptual Design purposes.

The process ponds and the underdrain collection pond will have a composite liner with a double geomembrane underlain by 0.3-m minimum compacted thickness of low-permeability cohesive soil layer, and a leak detection system between the geomembranes. The bottom (secondary) geomembrane will be 2.0 mm smooth LLDPE and the top (primary) geomembrane will be 2.0 mm single-side textured HDPE with texturing at top for traction. The leak detection system between the geomembranes will consist of a geocomposite connected to a leak detection sump and well system. The storm pond will have a composite liner consisting of 2.0 mm single-side textured HDPE with texturing at top for traction, underlain by 0.3-m minimum compacted thickness of low-permeability cohesive soil layer. 17.3 Process Plant The process plant consists of an ADR Plant, electrowinning cells, a refinery and reagent handling equipment. For the Phase II expansion, essentially a duplicate ADR



plant and electrowinning cells will be installed; the refinery and reagent handling facilities will be initially sized to accommodate the increase in metal production. The entire process plant designed by Scotia International of Nevada, Inc. treats four 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 refining.

The sourcing, transportation, handling, use and disposal of any hazardous substances will be regulated in accordance with relevant framework management plans prepared in accordance with international best practice to support the ESIA submission. 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 four tonnes of carbon. Hydrochloric acid, diluted to approximately 3 to 5 percent, recirculates through the carbon bed for a period of one to two hours. 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. 17.3.3 Carbon Stripping Metal is desorbed from the carbon in the strip vessel. The strip vessel holds four tonnes of carbon and operates under conditions of elevated temperature and pressure. Barren strip solution flows up through the bed of carbon, strips gold from the carbon, and then 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 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 the safety shower, wire samplers on the barren and electrowinning feed solutions and the 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 650 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. In the Phase II expansion, an additional kiln, kiln dewatering screen, and quench tank is installed. 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 attrited 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. 17.3.6 Electrowinning and Refining 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 ultimately is packaged into 50-gallon drums for off-site shipment and processing. The doré is packaged for off-site shipment. The refining 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 is 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 are added to the caustic mix tank to a 25 percent concentration. The caustic/mix transfer pump recirculates 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 the safety shower and the sump and pump reporting to the barren solution surge tank.

Metering pumps and lines deliver antiscalant directly from 50-gallon drums to the

barren solution surge tank. A pump and line deliver concentrated hydrochloric acid from carboys to the dilute

acid tank.



18.0 INFRASTRUCTURE

This section describes the proposed support facilities for the project. It includes the entire fixed infrastructure for mining operations, as well as the administration and plant operations.

18.1 Existing Infrastructure and Services 18.1.1 Location

The Amulsar Gold Project covers an area of 130 km2, located in south central

Armenia.

18.1.2 Site Access and Roads The Amulsar area is located some 170 km by sealed road from the capital city of

Yerevan, and 15 km by gravel track from the town of Jermuk. The license area straddles the boundary between Vayots-Dzor and Syunik provinces and incorporates part of the main highway south from Yerevan into Iran.

18.1.3 Buildings

An exploration camp owned by Geoteam is established at the site. The camp has

capacity for 15 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. Geoteam has also established an exploration sample preparation and core/sample storage facility in the village of Gorayk.

There is no allowance for a construction camp. During the construction period the

workers will be housed in Jermuk and travel to and from the work site.

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. Figure 18.1 shows a regional plan giving perspective of where the Amulsar project sits relative to Jermuk and Gorayk as well as the capital city of Yerevan.

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 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 type cabins or skid mounted equipment.



18.1.5 Communications

The exploration camp is currently serviced by satellite disk 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

Geoteam has employed a pool of field assistants during the various exploration

programs. It is intended to hire from this pool of known employees as a first priority for the operations phase. Community relations issues are currently handled by HSEC Manager and a Community Liaison Officer and a good understanding of local issues and sensitivities has been established. 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. The owners intend to have this in operating in third quarter of 2011.

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 at the 35 kV national grid, stepped down at AEN owned substations and reticulated as required to consumers.

Except for power consumption costs some capital costs to build power lines and

substation must be considered. The high tension power lines that stand along the main highway are 110 kV, however half of the year the line operates under 35 kV. The distance of the potential heap leach site is roughly 13 to 14 km from the highway. The design and construction of the lines for up to 12 km is a fixed cost. If the distance is longer than 12 km, every 250 meter will be an additional cost. The design and the construction of the power lines are done solely by AEN.

The construction and the design of the sub-station, which is an obligation to build,

is a licensed activity in Armenia. There are very limited companies that hold the licenses. An option to construct a 35 kV sub-station should be considered (as the 110 kV is

not available all year around). The 35 kV line will then be stepped down to 6 kV for reticulation around the site. From there the power would be distributed to the crushing plant and the ADR building.



Another option is to bring the power in at 35 kV to the main sub-station. Then from the sub-station distribute the power to the crushing and ADR plants, which will both have their own dedicated transformers. Here the 6 kV power would be stepped down to 220 V.

Thus the overall site power requirements are one transformer to step down from 35

kV to 6 kV and two transformers to step down from 6 kV to 220 V.

18.2 Site Development The Project will require development at six major locations: ▪ The mining areas. ▪ The mine surface facilities, including the mine administration building, mine

workshops, refuelling area, mine control areas and explosives yard. ▪ The crushing plant area including administration and contract laboratory. ▪ The ADR plant, leach pad and storage ponds. ▪ Accommodations. ▪ Road and site Access.

The ESIA team provides input into the site selection and design decisions for all

major infrastructure to ensure that environmental, 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.



Figure 18.1 - Proposed Overall Site General Arrangement Layout


K D Engineering Document No. Q431-02-028 12 August 2011 KDE FORM No. A263a-7/12/99

18.2.1 Mine Area A nominal area of 40,000 m2 has been earmarked for the mine services area and

is based on similar sized mining operations. The infrastructure required for the mining operation will be provided by the successful mine contractor and typically includes the refuelling station, go-line and mine control facilities for the mining fleet and administration buildings. The chosen site is located on reasonably gently sloping ground to the north of the plant site. The site will be cleared, grubbed and capped with laterite to provide pads for the facilities. A spur from the main haul road will provide mine vehicle access.

A normal area of 10,000 m2 has been earmarked for the explosives magazine. The

magazine area will be located away from the plant site and approximately 1 km north of the mine services area, on a spur off the haul road.

No large, heavily loaded structures are proposed in these areas. Based on similar

previous installations, the allowances made for bulk earthworks in the capital estimate are sufficient to accommodate any minor ground improvements which may be required subsequent to any detailed geotechnical investigations. These would be undertaken at the outset of the Project. 18.2.2 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 contractors services area will include the following facilities: ▪ Mine Administration Office. ▪ Ablutions. ▪ Mine Mess. ▪ Heavy vehicle workshop / store and washdown bay. ▪ Light vehicle workshop. ▪ Heavy vehicle fuelling station. ▪ Light vehicle fuel station. ▪ Fuel storage. ▪ Magazine.

18.2.3 Crushing Plant

The ROM pad and crushing plant will be located to the North of the main open pit. Sub-surface conditions will be established by test pit excavation and core drilling

supervised by Golder.



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. 18.2.4 Leach Pad & Ponds

The Adsorption-Desorption-Regeneration (ADR) treatment plant will be located to

the East of the heap leach process and storm ponds. Site preparation for the Adsorption/Desorption/Regeneration plant and associated

ponds consist of removing of all economic timber products from the basin area, removing of vegetation, clearing and grubbing of the embankment footprints and under-drainage areas.

A description of the design and construction of the leach pad and ponds is

contained in Section 17.

18.2.5 Accommodation The strategy for accommodating all construction personnel, employees and

security personnel during the construction period will be detailed as part of the BFS.

The town Jermuk has numerous run-down apartment buildings and hotels which could be upgrades for use as staff accommodation negating the need to build a permanent accommodation camp at the Amulsar site. 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. ▪ Mine haul roads.



▪ Borefield access road.

18.5 Administration and Plant Site Buildings The location of the various site buildings is shown in Appendix 3. All offices, the

plant control room and the MCC buildings will be air conditioned using individual split system units.

The following buildings have been allowed for: ▪ Administration building ▪ Plant administration ▪ First Aid clinic ▪ Laboratory ▪ Security ▪ Mess room ▪ Ablutions/Changeroom ▪ Plant Warehouse/Workshop ▪ Reagents Day Shed ▪ Chemical and hazardous substance storage facility ▪ Goldroom ▪ Sub-station and MCC buildings Current exploration assay requirements are met by the contract laboratory located

at ALS Chemex laboratories in Rosia Montana, Romania and Vancouver, Canada. The plant laboratory will be able to process Amulsar plant samples. The on-site

laboratory will comprise sample preparation, fire assay, an atomic absorption spectrometer (AAS) for assaying plant and elution solution samples, and facilities for minor metallurgical testing (screen analyses, coarse bottle roll and column leach tests).

The laboratory will consist of an insulated metal clad shed adjacent to the

goldroom building.

18.7 Water Supply

A site-wide water balance study and hydrological assessment will be conducted as part of the BFS/ESIA studies. Water sourcing and the impact on other users will also be assessed. 18.7.1 Potable Water Supply

Potable water will be used for drinking water, ablution water, laboratory water and

safety showers. Potable water is not required for any process purpose.



The design potable water demand is 115 m3/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 m3/d will be required within the plant (ablutions, laboratory, safety showers, etc). Accordingly, a supply of 370 m3/d has been allowed and will be purchased from local community supplies.

18.7.2 Raw Water

Subject the findings of the water balance study and hydrological assessments, the

primary source of raw water will be from the Vorotan River. An abstraction permit will be required in accordance with Armenian legislation; however, there is no reason to suspect that this will not be granted.

Rainfall for the model was based on 40 years of daily rainfall records collected at

the Vorotan pass Climate Station located approximately 4 km from the site. Raw water will be used in the following areas at an estimated rate of 0.2 tonnes of

water per tonne of ore: ▪ Elution ▪ Reagent make-up ▪ Cooling water ▪ Process water make-up ▪ Dust suppression 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 little pond on the western side of the pit. 18.8 Process Water Supply

The primary source of process water will be the Vorotan River. Process water

quality will be monitored and, provided it is acceptable, will be used in the following areas:

a. Leach pond make-up water b. Screen sprays c. Carbon transfer d. Process plant washdown



19.0 MARKET STUDIES AND CONTRACTS

Not applicable at this time.



20.0 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL IMPACT

Wardell Armstrong International Limited (WAI) was instructed by Lydian

International Limited (Lydian) to undertake a Scoping Study as part of the Environmental and Social Impact Assessment (ESIA) for the Amulsar open pit gold project in the southern part of the Republic of Armenia (RoA).

The scoping phase of the ESIA has been completed. Scoping requires the

gathering of preliminary information on the environmental context and the potential effects of the project. This enables the assessor to identify critical issues and provides an opportunity for early consultation with the local community and statutory authorities. The output of the scoping exercise is the ESIA Scoping Study report.

The purpose of the ESIA Scoping Study is to set out the main project parameters and identify any potential environmental and social impacts. It also sets out the national and international legislative framework and assesses the project status against international standards. Actions required have been identified in detail in a separate Gap Analysis. The main findings of these reports are summarised here, however, for completeness this report should be read in conjunction with the full versions of the Scoping Study and Gap Analysis (WAI Report No. EO-52-0088-1, dated February 2011).

The information presented below provides an overview of the environmental and

social conditions within the project area. It also identifies areas where extra work is required and planned.

20.1 Climate

The area is characterized by a temperate climate of long cold winters and short relatively cool summers. Climate data for the Vorotan River area are available from the Armenian State Hydro-meteorological Service (for the period of 1980-2010).

The maximum temperature in the summer is reported to reach greater than 30°C

and the minimum winter temperature is minus 28°C, with the coldest month frequently being January. April and May generally receive the most precipitation, with summer rainfall events capable of exceeding 60 mm/d. Evapo-transpiration rates are highest in July, with relative humidity being consistently elevated in winter months (December to March).

Table 20.1

Average Precipitation (mm) Month Jan Feb Mar Apr May Jun Jul Aug Sept Oct Nov Dec Annual

Average Monthly 45 56 75 91 108 71 49 32 29 58 42 50 706 Maximum Daily 22 44 32 32 67 45 72 63 41 41 31 21 72



Table 20.2 Humidity

Month Jan Feb Mar Apr May June July Aug Sept Oct Nov Dec AnnualEvapo-transpiration (hPa) 3.2 3.3 4.6 6.9 9.7 12.4 14.9 14.2 11.3 8.0 5.6 3.9 8.2 Relative humidity (%) 79 79 79 73 75 74 72 71 67 71 76 79 75 Maximum humidity 5 93 92 93 88 86 89 88 82 84 93 93 95 Minimum humidity 64 56 66 60 58 53 47 59 49 56 62 62 47

It should be noted that the preceding data has been averaged and sourced from a

lower elevation than the deposit. Site-specific data at a closer elevation to the open pit will be generated by the meteorological station recently installed at the open pit area and by a station established near HLP site 13.

Monitoring results indicate that annual air emissions did not exceed the maximum

allowable emissions standard established for the given source. Air emissions data collected throughout the year will continue to be documented into official logbooks. In addition, visual inspection of the bag-housings, and other control devices at the mine should be part of regular operational procedures and will be documented and reviewed by the Environmental Engineer as appropriate. 20.2 Water Resources & Quality

The project area contains numerous surface water features situated in the catchments of either the River Vorotan or Arpa. These comprise seasonal melt water rivers and permanent tributaries. The Vorotan flows through the license area to its outlet in the Spandaryan reservoir to the south.

At least one lake has been identified in the Exploration License area. The largest,

Moraine Lake, is currently abstracted to facilitate drilling activities. Wetland areas have been observed in the foothills and some of the proposed potential HLP sites.

A drinking water supply pipe and an irrigation surface channel divert water from the mountains and Vorotan River, through the Exploration License, to Gndevaz village. Groundwater resources in the project area have not been defined.

Samples from the Vorotan, Arpa and Darb Rivers and their tributaries, together with the Spandaryan reservoir and Gorayk drinking water supply are taken as part of the environmental monitoring and baseline data collection program (Refer to figure 20.1). Samples are tested for a wide range of physical, chemical and microbiological parameters.

Sporadic instances of nitrate, sulfate, Cu, Mn, Zn, Cr, Fe and V levels exceeding the maximum permitted concentrations (MPCs) have been noted. Other concentrations are predominantly below the detection limits, or 'limit of record' (LOR). The microbiological quality of the Gorayk water supply is poor for drinking water purposes, but a chlorinator has been supplied to upgrade it and will be supplemented by other treatment systems, as appropriate. The surface water sampling locations are shown in Figure 20.1.



During open pit mining there is potential for water inflow from two sources: precipitation and snow melt. Rainfall is generally more prolific during March to June. Seasonal water inflow from snow melts in spring, March to May.

Analysis of environmental baseline climate data from the on-site station will further characterize the likelihood and quantities of the above.

Figure 20.1 - Surface Water Sampling Locations in Relation to the License Boundaries

20.3 Hydrogeology

The spa town of Jermuk, to the north of the project produces bottled mineral water from groundwater resources.

Prior to fall 2010, no boreholes had been drilled at the project for the purposes of

groundwater monitoring. During the third and fourth quarters of 2010, three boreholes were drilled at the open pit site and each of three possible HLP sites and will be used to monitor groundwater characteristics and quality in these areas, refer to Figure 20.2.



Figure 20.2 – Potential HLP Sites



Six springs have been identified in the area of open pit. Five of the six springs are located in the Amulsar watershed and on the eastern slopes. Groundwater and spring water were tested for physical-chemical and chemical parameters.

This testing will be continued in 2011. Archived data on groundwater (flow, pH, major components) has also been sourced from the Republic of Armenia Geological Fund.

Historically, state exploration boreholes are believed to have been dry to depth in

excess of 70 m below ground level at the open pit location, however, water is added to facilitate drilling, hence groundwater strikes are not recorded.

A spring-fed lake (currently used for exploration drilling purposes) is present in the Mining License. Although groundwater is not expected to be encountered by the open pit, other areas of the license may have a shallower groundwater table. There is currently no definitive physical or compositional data on the groundwater system within the Exploration License. These resources remain to be fully defined. A hydrocensus and groundwater impact assessment will be required in both the open pit and designated HLP areas. The impact assessment will consider other users and sustainability of sourcing. Surface and groundwater will therefore be the subject of further detailed hydrogeological and hydrological studies for the ESIA and BFS in order to characterize the regime, assess potential impacts and allow these to be negated, designed-out or otherwise managed as the project develops.

20.4 Land and Landscape Quality

The footprint area of mine development infrastructure remains to be finalized,

however, it is considered very unlikely that the land taken for the project will materially affect local landowners, occupiers or tenants given the high elevation of the project. Nevertheless, land use will be assessed in the context of flora and fauna, ethno botany, and economic usage of grazing pastures.

Given the remote location of the site and the limited population of villages in the

vicinity, there are not considered to be significant numbers of residents that would experience visual intrusion. However, the elevation of the property extends the range of visual effect, which increases the number of vantage points from which an operation may be visible. Jermuk is a tourist destination and consequently, particular attention will be placed on the analysis and reduction of visual impact of the project development. Photomontage will be used to assess potential impacts and reduction measures.

Many internal haul roads are already in place and the effect of widening of these

will also be considered. A traffic Impact assessment will also be carried out, to assist in the identification of approach routes and network assessment, as part of the ESIA.

Archaeological and biodiversity field studies are currently being done for summer

2011.



Soil samples have been taken and analyzed for physical, chemical and biological baseline characterization purposes. Geotechnical and soil classification and distribution studies are continuing throughout 2011. 20.5 Potential In-Pit Water Flow During open pit mining there is potential for water inflow from two sources: precipitation and snow melt. Rainfall is generally more prolific during March to June. Seasonal water inflow from snow melts in spring, March to May. Baseline climate data from the on-site station will further characterize the likelihood and quantities of the above.

20.6 Seismicity Armenia is situated within the highly active Caucasus region in the vicinity of the Alpine-Himalaya seismic belt. The project is situated in an area of Armenia which is seismically active. Recorded earthquakes in the Zangezur region occurred in 1138, 1309, 1314, 1622, 1658, 1931, and 1968. The 1931 earthquake was registered as 8 on the Richter Scale and there were 50 fatalities in Syunik Marz. There is no detailed information on tremors in the project area. Detailed seismic assessments studies are planned for 2011 to aid in project assessment and design as part of the BFS.

20.7 Licenses and Permits

Subsequent to the exploration phase, and prior to development of the mine, several permits and licenses are likely to be required, including:

▪ Mining Licence ▪ Rock Allocation Area and change of land use status ▪ Technical Safety ▪ Water abstraction and discharge licence ▪ Air emission permit ▪ Change of land use in case the land is outside the RAA (the change takes up to

12 months for land outside the rock allocation area) ▪ Contract with Explosives Company, which has the permit to store, transport and

blast ▪ Contract with cyanide supplier and transportation companies that are fully

ICMC compliant. ▪ Construction and Architecture permits ▪ Gas and power use designs and construction expertise and permits ▪ Waste Passports ▪ Ecological Passport ▪ Environmental Impact Assessment to meet both national and international

requirements



20.8 Acid Rock Drainage (ARD) A Phase 1 ARD Characterization Program with respect to Amulsar waste rock has been carried out by WAI on 12 samples of potential waste rock, received from the Amulsar project in accordance with the agreed ESIA program. Waste rock types were identified by Lydian geological staff from batches of PQ drill core supplied to WAI for metallurgical testwork in 2010.

The screening assessments undertaken to date indicate that one waste rock, andesite porphyry, has acid-generating potential, and whilst the majority can be classified as non-acid generating, further studies and testwork will be required to verify these findings and further characterize the andesite porphyry and any borderline rock types.

The leachability testwork performed on these samples and the relatively low

leachability of all the inorganic components under TCLP conditions would indicate that dissolved metals would be below typical and regulatory levels (US EPA and Armenian Drinking Water Standards and accepted Toxicity Characteristic Regulatory Levels) should only low levels of acidity prevail.

A bespoke program of static analyses for the purposes of the ESIA waste rock

ARD impact assessment is underway and initial results, together with the requirement for kinetic testing and/or material management, will be available in Q3 2011. Detailed recommendations to enhance the ARD assessment to BFS level will also be provided. 20.9 Environmental Monitoring Program

Following the completion of the Scoping Study the terms of reference for baseline data gathering and monitoring have been further refined. Monitoring of environmental media will continue using a methodology designed by WAI and Lydian, informed by Armenian and International lender requirements. This program has been developed for the baseline stage of the ESIA. The submission of the draft ESIA will reflect the completion of the BFS, and is currently anticipated for Q2 2012. At construction stage and prior to operation, Lydian will transpose and modify this program so that it becomes adequate for on-going operational purposes.

The primary purpose of the current monitoring program is to assure that the project

remains in compliance with Armenian operating permits and environmental regulations and the IFC’s Performance Standards and to evaluate the effectiveness of the environmental mitigation measures. The results of the monitoring program are reviewed by project management on a periodic basis. If adverse environmental changes occur as a result of the project, appropriate remedial measures will be implemented to reduce or eliminate project-related effects. Specific details of any mitigation measures associated with unforeseen project-related effects will be developed based on the results of the monitoring. The current monitoring program contributes information to the environmental criteria for the baseline study in addition to monitoring on-site exploration and field activities.



An on-site Environmental Officer is employed at the project to oversee the current baseline monitoring and mitigation programs, assure that permit stipulations are met, direct worker environmental training programs, and serve as the contact for environmental status reports and site inspections. In addition, there is a Health, Safety, Environmental and Community team currently engaged with the project. The current monitoring program and team provide a valuable platform for rolling-out a formal Environmental Management and Monitoring Plan (EMMP) to guide further development of project operations. Such a plan will be developed in parallel with the ESIA process and then taken on and extended to form the basis of a construction and operation relevant EMMP. 20.9.1 Meteorology/Air Quality

Historic climatic data for monitoring purposes can be obtained from the nearby Vorotan Pass weather station; this can be supplemented by the recently installed Lydian Meteorological Stations. Data obtained from this meteorological station will provide information such as precipitation data, to aid in the refinement of the project water balance for the mine pit and process facilities.

A noise and air quality monitoring program will continue to inform the baseline study and will then extend into the monitoring of construction and full-scale operations. Air quality monitoring will involve both the collection of ambient air samples, and daily inspection of engineered air quality controls on various facilities. For the collection of ambient air samples, specific sampling stations will be identified, so that they can be used year after year to help improve the consistency of the results.

Monitoring results should indicate that annual air emissions from the past year did not exceed the maximum allowable emissions standard established for the given source. Air emissions data collected throughout the year will be documented into official logbooks. In addition, visual inspection of the bag-housings, and other control devices at the mine should be part of regular operational procedures and will be documented and reviewed by the Environmental Engineer as appropriate. 20.9.2 Groundwater Monitoring

Monitoring of the groundwater wells commenced in Q3 2011 and will be conducted as outlined in Table 20.3, Water Monitoring Schedule. On a monthly basis, the groundwater monitoring wells will be measured for water level elevations. In addition, field water quality parameters will be determined, including temperature, pH, electrical conductivity, turbidity, dissolved oxygen and salinity. On a quarterly basis, groundwater monitoring wells will be measured for water level elevation, field parameters, and the parameters listed in Table 20.4, Water Quality Monitoring Parameters.

Additional groundwater monitoring wells will be established around the project area

for evaluating the groundwater hydrology and water quality at the site. The location, spacing, and depth of these wells will be such as to comply with existing Armenian regulations regarding the monitoring of groundwater flows around heap leach pads or other industrial facilities.



A number of monitoring wells will be located around the Amulsar project site.

Specific locations for groundwater monitoring wells may be modified based on the results of more detailed hydrogeological evaluation of the project area.

Water quality data will be maintained on a database. This will allow for easy retrieval of the data, in addition to statistical evaluation of trends and parameters of interest.

Table 20.3

Water Monitoring Schedule

Monitoring Stations Field Parameters1

Laboratory Analysis3

Waste Rock Dumps Monthly Quarterly Heap Leach Pad Monthly Quarterly Solution Ponds Monthly2 Quarterly 1 Field measurements for groundwater wells include ground water level and temperature,

pH, and electrical conductivity. 2 WAD cyanide and pH. 3 Laboratory analyses shown on Table 18.2, Water Quality Monitoring Parameters.

Table 20.4 Water Quality Monitoring Parameters

No. Parameter No. Parameter 1 pH 16 Arsenic 2 Total dissolved solids 17 Bromine 3 Total suspended solids (surface water only) 18 Cadmium 4 WAD cyanide 19 Chromium 5 Alkalinity (total/CaCO3) 20 Copper 6 Bicarbonate as HCO3 21 Iron 7 Calcium 22 Lead 8 Magnesium 23 Manganese 9 Potassium 24 Mercury 10 Sodium 25 Selenium 11 Chloride 26 Silver 12 Fluoride 27 Zinc 13 Nitrate and N 28 Total petroleum hydrocarbons (TPH) 14 Ammonia + Ammonium 29 Radionuclides 15 Sulfate

20.9.3 Heap Leach Pad A monitoring program for the planned valley fill heap leach pad will be maintained to evaluate and document the performance of the leaching facility. This monitoring program is divided into the following parts:

▪ Geotechnical observations ▪ Hydrologic monitoring ▪ Water balance review



Geotechnical Observations

During operations, regular visual inspections of the heap pad and facilities will be made to check the condition of the embankments, liner, pipelines, and sprinkler facilities. Observations will be recorded in a field diary. Special attention will be given to any gullying and erosion, distinct changes in vegetation growth, plugged pipelines or drains, and the actual operation of any monitoring instrumentation. A record of such data will be maintained by the Environmental Engineer on-site. Hydrological and Water Quality Monitoring During operations the standard monitoring program will include hydrologic analysis for the following areas:

▪ groundwater ▪ leach collection system piping ▪ collection ponds

The general groundwater monitoring program for various facilities at the mine site was addressed above in Section 20.9.2, Groundwater. The groundwater monitoring program around the heap leach facilities will be designed to measure any changes in existing groundwater conditions down-gradient of the leach pad. Solutions in the collection ponds will be analyzed weekly for pH and cyanide as set forth in Table 20.3, Water Monitoring Schedule. Water Balance Review

A heap leach facility water balance will be developed on an annual basis as based on variations in operations and meteorological conditions. The heap leach pad and processing facility should be maintained using a closed circuit water supply. Monitoring of the water balance will serve to indicate if any significant and unexplained water losses or gains in this system are occurring. Such losses or gains could indicate seepage or other problems along the closed circuit water supply system.

20.9.4 Waste Rock Drainage

The ESIA acid rock drainage (ARD) impact assessment for Amulsar waste rock types will identify the potential acid- and leachate-generating rock types which can then be segregated and managed, if necessary, to prevent soil, surface water and groundwater contamination. Furthermore, to evaluate the chemical composition of the runoff and potential infiltration to groundwater from ARD from the waste rock dumps, monitoring groundwater quality down-gradient of the waste rock disposal or ore stockpiling areas will be conducted.

Acid rock drainage will be addressed for the waste rock dumps, ore stockpiles, and

mine pits. The parameters considered to be of greatest concern in potential waste rock



drainage will be: acid generating potential, blasting residuals (ammonium nitrate and fuel oils), and total suspended sediments (TSS).

During mining, water quality samples will be collected from groundwater monitoring wells, mine workings, and any springs or seeps that develop in the toe areas of the waste rock dumps or ore stockpiles. Field pH and conductivity will be monitored on a monthly basis in waters collected from groundwater monitoring wells down-gradient of these facilities. On a quarterly basis or as determined by seasonal constraints, detailed laboratory analyses will be obtained for these sites. The schedule for monitoring is shown in Table 20.3, Water Monitoring Schedule. Laboratory analyses to be conducted are presented in Table 20.4, Water Quality Monitoring Parameters. As part of the spring-and-seep surveys, a visual inspection of soils will be conducted for evidence of vegetation damage and/or mineral staining. Such surface expressions could be related to ARD and should be photographed and monitored for changes. Before it is placed in the dump, waste rock will be characterized by its mineralogy. Lithologies with a high potential to oxidize and create acid conditions will be encapsulated in materials with a high potential to neutralize acid, enclosures will be constructed, or other measures taken to minimize the potential for acid rock drainage. 20.10 Reclamation and Closure For a mining project to have a positive contribution to the sustainable development of a community or region, closure objectives and impacts must be considered from project inception. Closure and reclamation objectives include:

▪ Future public health and safety are not compromised; ▪ Environmental resources are not subject to physical and chemical deterioration; ▪ The after-use of the site is beneficial and sustainable in the long term; ▪ Any adverse socio-economic impacts are minimised; and ▪ All socio-economic benefits are maximised.

In accordance with international best practice for the mining industry and the

environmental policy of Lydian International Ltd, a conceptual MCRP will be developed for the Amulsar site during the BFS and environmental and social impact assessment (ESIA) phase of project design. Additionally, further work will be required in the detailed engineering design stage to refine the concepts and produce a Final MCRP, to ensure that the site is properly rehabilitated. At this stage of the project development, it is possible to state a commitment to:

▪ progressive rehabilitation of areas as they are worked, ▪ decontamination of affected sites; ▪ provide stability of physical and chemical environmental factors and; ▪ development of a detailed plan for eventual closure and implementation of

rehabilitation as the study proceeds.

Steps will need to be taken to ensure that funding is available to rehabilitate the mine upon closure and contribute to the long-term sustainable development of the area.



According to the Concession law of the Republic of Armenia, the design and cost calculation of reclamation works should be part of the designs and the environmental bonds should be paid, with 15% of the costs during the first year after the license is granted and the rest of the sum is divided onto the remaining years of the project. The Environmental bonds are paid on the special account in the Central Bank of Armenia.

Since the basic objective for the rehabilitation of any mineral operation is to achieve a long-term sustainable after-use of the site, Lydian International will develop a plan in accordance with the following seven principle objectives:

▪ minimise the area of land to be used for mining, processing and waste disposal which will require rehabilitation;

▪ ensure that opportunities for progressive rehabilitation of leach pads, waste

dumps, pits and ponds are identified, planned and programmed;

▪ construct landforms that provide long-term physical and chemical containment and stability of mine and process wastes that are acceptable to the regulatory authorities and are in-line with international good practice;

▪ avoid aspects of the mining and leaching operation and associated activities

that would make rehabilitation more difficult, costly or less effective;

▪ ensure that post-operation landforms are provided in a suitable form for the proposed final after-use;

▪ make certain that post-closure groundwater and surface water quality is

acceptable to the regulatory authorities and meet international standards;

▪ follow a consultative process, involving all interested parties, to determine the preferred after-use for the operational area and associated facilities.

Although the MCRP will primarily address closure on exhaustion of the mineable

resources and completion of processing of stockpiled ore, it must also provide for orderly decommissioning and rehabilitation should premature closure be required.

Decommissioning procedures for cyanide facilities will be integrated into the overall MCRP. Decommissioning and closure of cyanide facilities will entail the removal or detoxification of unused sodium cyanide and the clean-up of cyanide residues in process tanks and equipment. The specific measures to accomplish these tasks will be included in the MCRP, as should cost estimates and information on financial assurance for decommissioning and closure, in accordance with the International Cyanide Management Code (ICMC). All cyanide process tanks and piping systems should be triple flushed with water to remove residual cyanide, and the effluent routed to a detoxification circuit for reduction of residual cyanide concentrations to below EU standards, or other appropriate guidelines for cyanide in tailings. The decommissioned process plant tanks and piping systems should then be cut up for disposal or recycling.



Any decant water remaining in the solution ponds after process plant decommissioning should be pumped to the secondary cyanide treatment plant and treated. It is suggested that the water quality discharge standard be less than 0.1 mg/l total cyanide. However, approvals from the appropriate Armenian government will need to be agreed.

The MCRP for Amulsar will be integrated with annual action plans, especially with regard to environmental and socio-economic management issues. On-going reviews of the closure objectives and design are necessary to allow for changes in political, legislative, physical or socio-economic conditions. If planning is delayed, it may affect which mine closure objectives can be met. Guidance may therefore be required in carrying out periodic studies of closure options, to reduce the uncertainties and ensure that the MCRP remains adequate, realistic, appropriate and integrated within the whole operation life plan. Post-closure environmental monitoring and after-care will be required, in accordance with the established Environmental Management and Monitoring Plan (EMMP). This includes the monitoring of all sensitive receptors (including air quality, soil, groundwater and surface waters) in the project area and has been designed in accordance with international standards. Post-closure monitoring of groundwater in the vicinity of the heap leach facility will be of particular importance. The cost of environmental mining should be incorporated into the costed MCRP and will need to continue for a minimum of 5 years after rehabilitation of the site, in line with international best practice. The Armenian legislative requirements for post-closure monitoring will also be incorporated into the MCRP, as directed. 20.11 Social Context The project is located on the border of two Marzes (Vayots-Dzor and Syunik). Three villages with a total population of approximately 2 000 are potentially affected by the Project. These include the communities of Saravan (consisting of Saravan, Saralanj and Ughedzor (which is only inhabited during the summer months)) and Gndevaz both of which are in Vayots Dzor Marz some 5 to 9 km to the west of the deposit and Gorayk, located in Syunik Marz, 6 km southwest of the deposit. The closest city is Jermuk (which includes the village of Kechut), located 13 km from the project. With a population of approximately 6,200 people, the city is endowed with natural hot springs and waterfalls and is home to several national and international renowned health resorts and spas. Marzes are separated into different administrative centers, with Gorayk reporting into Sisian town in Syunik Marz and Gndevaz, Saravan and Jermuk reporting to Vayk in Vayots Dzor Marz. The nature of the anticipated impact varies for each community, as is summarized below. Detailed engineering for the project is ongoing, and as such, the final location for some facilities has not yet been confirmed. The impacts currently anticipated include:



▪ Gorayk - Areas of seasonal pastureland will need to be acquired by the project to locate some facilities. Significant employment opportunities will be available to this village.

▪ Saravan - Packages of land will need to be acquired by the project to locate

some facilities. It is likely that the village of Saralanj will be able to see the project activities on the mountain. Significant employment opportunities will be available to this village

▪ Gndevaz - The project activities will likely be visible from Gndevaz. Significant

employment opportunities will be available to this village.

▪ Jermuk - Studies are being undertaken to determine the visual amenity impacts of the project upon Jermuk and mitigation plans will be developed once these studies are complete. Jermuk will also be the likely residence of many of the workers on the project as it develops.

20.11.1 Standard of Living, Infrastructure and Services In order to assess the standard of living in the study areas certain indicators are presented below: Income Table 20.5 shows the relative distribution of income brackets in the study area.

Table 20.5 Income Distribution in the Study Area

Percentage of Households Gorayk Saravan Gndevaz Less than US$ 191.001 per month 70.9 79.6 58.2 Between US$ 191,00- 273 per month 17.7 11.1 16.3 Between US$ 274 - 409 per month 7.6 1.9 14.9 More than US$ 510 per month 3.8 7.4 10.6 Source: MPG consultant survey, 2009 1 Conversion 1 US Dollar = 366.30 AD (Armenian Drams)

From the table it can be seen that in Gndevaz incomes are the highest. Agricultural

produce is the major source of income, accounting for 54.6 percent in Gndevaz, 69.6 percent in Gorayk and 75.9 percent, Saravan. Housing Most houses are made from stone with an asbestos roof. The majority consist of 3 or 4 rooms with a kitchen, bathroom and toilet in Saravan community outside the main structure. The dwellings in Gndevaz look to be in generally better condition than in the other villages.



Infrastructure All communities have access to electricity, whilst only Gndevaz and Gorayk have access to piped gas, which they use for cooking. Saravan uses gas canisters or wood. All communities mostly use wood and dried manure for heating. All communities have access to piped water; however the quality of water in Gorayk and Saravan is sub-standard. The information of poor water quality of Saravan is based on people’s concerns, while the Gorayk water was tested and proved to be of poor quality.

There is no domestic waste removal or sewerage system in the communities except for Gorayk, where the waste removal is organized by Geoteam. Solid waste is mostly burned by households and what cannot be burned is disposed of by crude landfill in the valleys surrounding the villages. 20.12 Historical and Social-Cultural Situation In order to contextualize and provide a better understanding of the current communities of the study area a short historical description of each community is provided1: 20.12.1 Jermuk Ruins of an ancient fortress and an 8th century basilica testify to early human settlement in Jermuk. The foundation of the modern town of Jermuk, however, only took place in 1940, when the hot springs were discovered and the first sanatorium was opened to the public. Development programs were implemented to turn Jermuk into a modern resort for all Soviet nationals. In 1980 the population of Jermuk reached 10,000. After the collapse of the Soviet Union and as a result of the economic crisis in Armenia during the 1990s, the population dropped to 6000 (including the village of Kechut, which was included administratively into Jermuk in 1996) in the beginning of the 21st century. Recently however, the Armenian authorities took steps in order to further develop the Jermuk town-resort. As a result of those new resolutions the town has experienced unprecedented success as a tourist destination. 20.12.2 Gorayk The exact date of the foundation of Gorayk could not be established. However during the 18th century the village was populated by Russians (Malakans) who were resettled from Russia due to religious issues. During the period 1912 to 1914 Armenians from surrounding villages came to settle in Gorayk and by 1951 all Russian inhabitants had left. During 1965, under the soviet system, the original Gorayk village was flooded to create a reservoir (Spandaryan) and people were moved to the area where the village is currently located (2 km away from original spot). 1 Data for the historical background is gleaned from Wikipedia and discussions with local village chiefs.



The current position however is not favored by the population. It is regarded as a swampy and damp area, unfavorable for agriculture and lacking clean water. 20.12.3 Gndevaz There is archeological evidence (caves) of human habitation from 2000 BC. In 1300, a village was founded, but in 1602 the population was chased away by an Iranian Shah. Despite this, in 1800 the village started to regrow with families from surrounding villages. An old church, Gndevank is located in the gorge next to the village which is found in 10th Century BC. The Church is regularly visited by tourists most of which are based in Jermuk.

20.12.4 Saravan

Saravan is a much more recent settlement founded in 1950. It was subsequently inhabited by Azerbaijanis in 1980.

After 1988, all Azerbaijanis left and Armenians returned. Currently 45% of the population consists of Armenian refugees who fled Azerbaijan during the war. 20.12.5 Socio - Cultural Aspects

To further provide insight into the social life of the local population, some pertinent socio-cultural aspects characterizing the communities are listed below:

Family life and family allegiance is the cornerstone of the local community. Often family units consist of different generations, with sons bringing their wives into the family home. Mother and daughter in law relationships are paramount, with the mother in law firmly managing the household assisted by daughters and daughters in law. Although women have an important role in the household, men are regarded as the head of the family and community affairs are predominantly managed by men. There seems to be little evidence of generation conflict, with young men and women performing their roles within the extended household. The lack of jobs however, forcing young people to search for jobs outside the village may change the generation dynamics.

Armenia prides itself as being the first country to declare itself a Christian state and local communities are all adherents of the Armenian Apostolic Church. The religion can be both of a cultural lifestyle and a strict doctrine with rules and regulations.

20.12.6 Health

In discussion with local residents and health professionals it was indicated that the main health problems in the area are: high blood pressure and related heart problems, osteoporosis, cancers, diabetes, urinary tract infections and seasonal health problems such as influenza in winter and bee stings and diarrhea (due to impure water) during summer. It was also indicated that TB was on the increase.



Clinics in the villages are first tier facilities, with the health post in Gorayk referring to the hospital in Sisian for more major health problems and Saravan, Gndevaz and Jermuk referring to Vayk. There are no pharmacies in the villages, the health posts however stock a number of basic medicines which can be obtained if the resident pays a quarterly adherence fee to the clinic and for free as well. Certain categories of patients (certain chronic ailments) are provided with free medication. The clinic staff is remunerated by the government and patients that are not included in special assistance groups are required to pay a consultation fee.

Armenian rural people have the reputation of being reluctant to visit a health facility and generally wait until problems have become serious health threats before seeking medical help. Minor ailments are often dealt with at home with herbal medicines. 20.12.7 Education

All three villages have a primary school and schooling is free. The number of students at schools is however low for the capacity of the buildings (Gorayk 70, Gndevaz 133, Saravan 30) and numbers are still decreasing, threatening the schools with closure. 20.12.8 Communications

Transport All villages can be reached by tar road. Gorayk and Saravan are located on the M2 highway (silk route), Gndevaz and Jermuk are on the H43 - H42. There is a regular bus service, connecting with Yerevan and other urban areas. Roads within the villages are un-tarred, dirt roads which are often in bad condition. A substantial number of households possess private cars. Communications

There is a daily newspaper delivery in all of the villages, however only few households use the service. There are several TV stations and mobile phone networks available as well as fixed telephone lines. Each village also boasts a post office. 20.12.9 Development Needs and Community Problems During the survey conducted in 2010 the following community needs were identified.



Table 20.6

Priority Needs in the Study Area Gorayk

▪ Asphalt surfacing of roads ▪ Restoration and repair of the Recreation center ▪ Rehabilitation and reopening of the kindergarten ▪ Establishment of a bulk purchase stock ▪ Operation of any industrial plant in the community ▪ Employment ▪ Waste collection and removal system ▪ Subsidies, support to the farmers for purchasing seeds and fertilizers

Saravan ▪ Asphalt surfacing of roads ▪ Irrigation water supply improvement ▪ Gas supply ▪ Street lighting ▪ Rehabilitation of the sewerage system ▪ Establishment of a purchase point

Gndevaz ▪ Asphalt surfacing of roads ▪ Rehabilitation of the water pipes ▪ Organization of entertainment ▪ Street illumination

Source: MPC consultant survey, 2010

From the table it can be seen that the main focus is on infrastructure improvement. Asphalting of village roads and improvement of sewerage systems are a priority in all villages. Further to the table above employment and improved marketing facilities were repeatedly identified as priority needs.

Geoteam has been supporting community development activities in the three

villages since 2006 and more recently including Jermuk. Activities have included supporting the extension of a gas pipeline into Saravan, job creation through support for new enterprises (eg sewing of sample bags), supporting students that study Geology and Mining Engineering at the Universities of Armenia, refurbishment of a number of community facilities, and partnering with NGOs to extend development programs into the region (eg Kinderwall). 20.13 ESIA Recommendations and Scope of Works

A work plan covering all of the issues identified has been formally developed between Lydian/Geoteam and WAI and work is progressing against each of these topics. WAI has recommended that the scope of ESIA baseline and assessments to quantify and/or mitigate against environmental and social issues identified include:

▪ Groundwater and surface water hydrology, quality and sustainability - further hydro-geological study, modelling and geochemical assessment via the installation of monitoring wells needs to be undertaken. Site-wide water balance calculations need to be determined and the sustainability of sourcing examined and managed;



▪ Visual impact – The visual impact of the project needs to be assessed from

Jermuk, the public highway and village communities, with the results included in public consultations;


required;

▪ Land disturbance and take - a land use survey, biodiversity assessment and archaeological study are required to inform an impact assessment and assessment of any compensation required for grazing lands;

▪ Biodiversity - field surveys need to be undertaken on areas of major

disturbance. An assessment of flora harvested and collected by villagers is required. These studies will inform an overall impact assessment;

▪ Dust - dust dispersion modelling and impact assessment on edible plant

species used by the communities needs to be undertaken;

▪ Noise - Noise impact modelling, incorporating the operating specifics of mine fleet and machinery needs to be undertaken to inform the noise impact assessment on nearest sensitive receptors;

▪ Cyanide and hazardous materials - As the project will use cyanide in its

heap leach process, the usage, transport, handling and disposal of cyanide will need to be managed via formal plans and procedures in accordance with international best practice, particularly the International Cyanide Management Code (ICMC);



▪ Local labour - high unemployment and expectations require careful

management and a Skills Audit to assess future potential employment opportunities should be undertaken;


assessment (to include seasonal visitors) need to be undertaken;

▪ Local Industry - further appraisal of the main (agricultural and manufacturing) sectors is required, including individual contributions to domestic income. An assessment of potential diversification and increase in access to markets with advent of mining is considered beneficial;

▪ Public disclosure and consultation - to be undertaken in line with

international best practice, at all stages of the ESIA and mine development



processes (to include Jermuk) to manage expectations, allay concerns, establish NGO partnerships, inform community development and spending and influence mine design;

▪ Closure – considerations need to be addressed from conceptual stage via

the preparation and continual update of a Mine Closure and Rehabilitation Plan (to include heap leach rehabilitation).

WAI propose to monitor and review the process of environmental and social

baseline data collection by Lydian/Geoteam, and general progression of the ESIA works, and provide continuing technical support, as required, to ensure that both national and international requirements are met. The baseline programme is continuing and the next ESIA deliverable will be a baseline status report in Q3 2011.

WAI will also work closely with the BFS team to ensure that environmental and

social issues are taken into account with the design and siting of all major infrastructure and that appropriate alternatives are considered and/or mitigation and management measures are employed to ensure that impacts are addressed. It is therefore recommended that a time-bound, integrated ESIA and BFS schedule of works is prepared to facilitate this process.

WAI will need to ensure that the public consultation, which will be undertaken for

the purposes of the ESIA, will include available information from the BFS so that project stakeholders have the opportunity to input into the process.



21.0 CAPITAL AND OPERATING COSTS

Capital and operating costs are presented for both contract mining and owner operated fleet.

The following mining and capital cost text is extracted from CSA Report No

R268.2011 submitted to Lydian in July 2011. 21.1 Mine Capital Costs

Summaries of Life of Mine (LOM) capital costs for both options are shown in Table 21.1. Initial and working capital costs to achieve 10 Mtpa capacity are estimated at $121.6M for the owner operator mining option and $18.1M for the contractor mining option.

Table 21.1 Capital Costs Summary

Description Units LoM Totals Total Waste Mined (Dry Tonnes) Mt 136 Total PMM (Dry Tonnes) Mt 55.3 Total Rock Tonnes (Dry Tonnes) Mt 191.3 Mining Initial and Sustaining Capital Expenditure Facilities Capex (Site Preparation Works) US$M $8.7 Unit Cost per Tonne Rock Mined US$/t Mined $0.05 Unit Cost per Tonne of PMM Mined US$/t PMM $0.16 Mobilisation, Demobilisation, Site Establishment & Facilities Capex US$M $9.3 Unit Cost per Tonne Rock Mined US$/t Mined $0.05 Unit Cost per Tonne of PMM Mined US$/t PMM $0.17 Mining Plant Capex (All Major and Minor Mining Equipment) US$M $103.6 Unit Cost per Tonne Rock Mined US$/t Mined $0.54 Unit Cost per Tonne of PMM Mined US$/t PMM $1.87 Total Capital Expenditure US$M $121.6 Unit Cost per Tonne Rock Mined US$/t Mined $0.64 Unit Cost per Tonne of PMM Mined US$/t PMM $2.20 Contractor Mining Initial and Working Capital Expenditure Facilities Capex (Site Preparation Works) US$M $8.8 Unit Cost per Tonne Rock Mined US$/t Mined $0.05 Unit Cost per Tonne of PMM Mined US$/t PMM $0.16 Mobilisation, Demobilisation, Site Establishment & Facilities Capex US$M $9.3 Unit Cost per Tonne Rock Mined US$/t Mined $0.05 Unit Cost per Tonne of PMM Mined US$/t PMM $0.17 Mining Plant Capex (All Major and Minor Mining Equipment) US$M Unit Cost per Tonne Rock Mined US$/t Mined Unit Cost per Tonne of PMM Mined US$/t PMM Total Capital Expenditure US$M $18.1 Unit Cost per Tonne Rock Mined US$/t Mined $0.09 Unit Cost per Tonne of PMM Mined US$/t PMM $0.33



21.2 Mine Operating Costs

Summaries of LOM mining operating costs for both options are shown in table 21.2. Total LOM mining unit operating costs at Amulsar are estimated as $6.84/t PMM for the owner operator option and $8.86/t PMM for the contract mining option.

Table 21.2 Operating Costs Summary

Description Units LoM Totals Total Waste Mined (Dry Tonnes) Mt 133 Total PMM (Dry Tonnes) Mt 55.3 Total Rock Tonnes (Dry Tonnes) Mt 191.3 Owner Operator Mining Operating Costs Load and Haul Fleet US$M $197.23 Unit Cost per Tonne Rock Mined US$/t Mined $1.03 Unit Cost per Tonne of PMM Mined US$/t PMM $3.57 Ancillary Equipment US$M $45.48 Unit Cost per Tonne Rock Mined US$/t Mined $0.24 Unit Cost per Tonne of PMM Mined US$/t PMM $0.82 Overheads US$M $59.13 Unit Cost per Tonne Rock Mined US$/t Mined $0.31 Unit Cost per Tonne of PMM Mined US$/t PMM $1.07 Drill & Blast US$M $64.59 Unit Cost per Tonne Rock Mined US$/t Mined $0.34 Unit Cost per Tonne of PMM Mined US$/t PMM $1.17 Operating Costs - Works US$M $11.81 Unit Cost per Tonne Rock Mined US$/t Mined $0.06 Unit Cost per Tonne of PMM Mined US$/t PMM $0.21 Total Operating Expenditure US$M $378.2 Unit Cost per Tonne Rock Mined US$/t Mined $1.98 Unit Cost per Tonne of PMM Mined US$/t PMM $6.84 Contractor Mining Operating Costs Load and Haul Fleet US$M $278.89 Unit Cost per Tonne Rock Mined US$/t Mined $1.46 Unit Cost per Tonne of PMM Mined US$/t PMM $5.05 Ancillary Equipment US$M $60.65 Unit Cost per Tonne Rock Mined US$/t Mined $0.32 Unit Cost per Tonne of PMM Mined US$/t PMM $1.10 Overheads US$M $67.42 Unit Cost per Tonne Rock Mined US$/t Mined $0.35 Unit Cost per Tonne of PMM Mined US$/t PMM $1.22 Drill & Blast US$M $69.59 Unit Cost per Tonne Rock Mined US$/t Mined $0.36 Unit Cost per Tonne of PMM Mined US$/t PMM $1.26 Operating Costs - Works US$M $13.34 Unit Cost per Tonne Rock Mined US$/t Mined $0.07 Unit Cost per Tonne of PMM Mined US$/t PMM $0.24 Total Operating Expenditure US$M $489.9 Unit Cost per Tonne Rock Mined US$/t Mined $2.56 Unit Cost per Tonne of PMM Mined US$/t PMM $8.86



21.3 Process Capital Costs

A summary of the initial capital costs and sustaining capital costs is shown in the following Tables 21.3 and 21.4 respectively. Said tables include direct costs, indirect costs, and a contingency. A detailed description of the estimate is presented in Appendix 3.

Table 21.3 Summary Process Plant

Initial Capital Cost Description Total Cost, US$

DIRECT COSTS Area 10 - Primary Crushing 4,636,264Area 13 - Secondary Crushing 4,222,504Area 15 - Tertiary Crushing 9,475,435Area 17 - Product Storage Loadout & Overland Conveyor 23,038,905Area 20 Heap Leach 1,151,150ADR Plant 7,436,000Area 30 Solution Management 965,250Area 90 Auxiliary Equipment 50,000Process Piping 2,548,775Instrumentation 2,548,775Site Development 12,743,877Water & Electrical Power to Site 7,646,326

SUB-TOTAL DIRECT 76,463,262 INDIRECT COSTS EPCM 9,175,591Construction Indirect Costs include: 7,646,326 Construction Supervision Equipment Rental Field Office Expense Mobilization / Demobilization Consumables Spare Parts 1,783,112Initial Fill & Reagents 1,146,949Equipment Insurance & Freight Cost 5,352,428Camp Cost (200 people for 6 months) 1,280,950

SUB-TOTAL INDIRECT 26,385,357TOTAL DIRECT AND INDIRECT 102,848,619

Contingency – 30% 30,854,586TOTAL 133,703,205



Table 21.4 Summary Process Plant Sustaining Capital Cost

Description Total Cost DIRECT COSTS Area 10 - Primary Crushing 3,037,524Area 13 - Secondary Crushing 5,276,371Area 15 - Tertiary Crushing 5,174,638Area 17 - Product Storage Loadout & Overland Conveyor 9,932,780ADR Plant 2,000,000Process Piping 254,213Instrumentation 254,213Site Development 254,213

SUB-TOTAL DIRECT 26,183,953 INDIRECT COSTS EPCM 1,832,877Construction Indirect Costs incl: 2,618,395 Construction Supervision Equipment Rental Field office expense Mobilization / Demobilization Consumables Spare Parts 918,927Equipment Insurance & Freight Cost 1,832,877

SUB-TOTAL INDIRECT 7,203,076TOTAL DIRECT AND INDIRECT 33,387,029

Contingency - 30% 10,016,109TOTAL 43,403,137

21.3.1 Direct Costs The Direct capital costs were based on the following list of documents:

▪ Design Criteria ▪ Equipment list ▪ Western Mining Source Quote Data ▪ KD Engineering Equipment Database ▪ Engineering Drawings performed by KDE

- 10-F-01 Flowsheet Primary Crushing - 13-F-01 Flowsheet Secondary Crushing - 15-F-01 Flowsheet Tertiary Crushing - 17-F-01 Flowsheet Lime Addition - 19-F-01 Flowsheet Ore Stacking - 20-F-01 Flowsheet Heap Leach - 23-F-01 Flowsheet Carbon Adsorption - 25-F-01 Flowsheet Stripping and Refining - 27-F-01 Flowsheet Carbon Reactivation - 30-F-01 Flowsheet Utilities and Reagents



The direct costs exhibited in this estimate include, but are not limited to, labor, equipment and materials for the detailed construction activities set forth below: Equipment Costs An equipment list was developed and incorporated into the cost estimate. The estimate for equipment was developed from the following sources:

▪ Written or e-mailed budgetary estimates from vendors for major equipment. ▪ Budget unit costs were provided from a local contractor for building costs. ▪ Historical data and budget costs of similar projects for miscellaneous

equipment.

Moreover, the cost of “Installed Equipment” was estimated using a forty-three percent (43%) factor. Site Development Site development costs include excavations and back fills calculations using preliminary plans and historical data for similar plants. The initial phase cost was estimated using a twenty-five percent (25%) factor of the installed plant equipment cost. The factor was reduced to one (1%) for the second phase. The Project will require development at the following major locations:

▪ The mining areas ▪ The crushing plant area including administration and contract laboratory ▪ The mine services area, including the mine administration building, mine

workshops, refueling area, mine control areas and explosives yard ▪ The leach pad and storage ponds

Fresh, Raw and Process Water Fresh, raw and process water is discussed in Section 18 of this report. Electrical and Instrumentation Electrical and instrumentation are included using historical data for similar plants. Emergency powered generators have been included to continue to supply power to critical pieces of equipment (thickener mechanisms, agitators, etc.) in the event of a power outage. An estimate for a power line installation and substation from the local Electrical Utility is included in the factored estimate. Power Supply to Mine Power is not currently provided to the Project site although domestic usage power is available at neighboring main towns to the south and east. The supply of



power in this region is controlled by the Armenian Electrical Networks company (ANE) and will be stepped down at ANE owned substations and delivered to the mine and plant site. A small hydroelectric power plant is being developed on the Vorotan River to the NE of the Amulsar mine site which will have a total installed capacity of 1.8 to 2 MW. This is expected to be in operation in Q3 2011. The costs for Water (Fresh, Raw & Process) and Electrical Power (Electrical Instrumentation & power supply) to the site were estimated using a fifteen percent (5%) factor of the installed plant equipment cost. 21.3.2 Indirect Costs Certain indirect costs exhibited in this estimate include, but are not limited to, labor, equipment and materials for the detailed activities set forth below:

▪ EPCM for the initial phase were estimated using a twelve percent (12%) factor of the installed plant equipment cost. The factor was reduced to seven percent (7%) for the second phase and includes:

- Detailed Engineering - Procurement - Construction Management - Training

▪ Construction Indirect Costs for the initial phase were estimated using a ten

percent (10%) factor of the installed plant equipment cost and includes:

- Construction Supervision - Equipment Rental - Field Office Expenses - Mobilization/Demobilization - Consumables

▪ Spare parts costs were estimated using a five percent (5%) factor of the

installed plant equipment cost.

▪ Initial Fill & Reagents costs were estimated using a one and half percent (1.5%) factor of the installed plant equipment costs.

▪ Equipment Insurance and Freight costs were estimated using a seven

percent (7%) factor of the installed plant equipment costs. 21.3.3 Contingency and Accuracy

The KDE crushing and process plant portion of the cost estimate includes a 30 percent contingency for project unknowns and identified risks. Contingency is a



necessary part of the cost estimate and KDE utilized 30 percent based on the fact less than 5 percent of the engineering is completed to date. KDE believes the estimated contingency amount will be spent during the life of the project for identified risks and unknown items. While KDE has not performed a statistical analysis of the crushing plant and process plant accuracy of the capital cost estimate we believe based on previous experience with similar projects there is a high confidence that the accuracy of the process portion of the PEA capital cost estimate will end up between -10 percent and +35 percent of the KDE capital cost estimate. 21.3.4 Exclusions KDE has excluded the following items from the process plant estimate and they are included elsewhere:

▪ Permits, royalties and licenses ▪ Environmental testing and monitoring ▪ Metallurgical testing ▪ Escalation and Insurances ▪ Taxes, duty and import fees ▪ Solution and ponds ▪ Mining costs ▪ Reclamation and closure costs ▪ Geotechnical design and facility costs ▪ Camp Costs ▪ Allowance for design growth or specification changes

21.4 Process Operating Costs Annual and unit process operating cost estimates for Phase I (5 million tonnes per year) and Phase II (10 million tonnes per year) process operation are summarized in Table 21.5. Support Tables for the cost estimates are shown in Tables 21.6 to 21.10.



Table 21.5

Process Plant Operating Cost Estimate - Summary0MTPA

Cost

Centre

Phase I (5 Mtpa) Phase II (10 Mtpa)

Annual Cost (US$)

Unit Cost

(US$)/tonne Ore Treated

Cum Unit Cost


Annual

Cost (US$)

Unit Cost


Cum Unit Cost


Plant - Labor 5,718,161 1.14 1.14 6,154,183 0.62 0.62 Plant - Consumables 9,709,827 1.94 3.09 19,019,654 1.90 2.52 Power & Energy 2,008,449 0.40 3.49 2,839,674 0.28 2.80 Mechanical 3,467,415 0.69 4.18 4,776,612 0.48 3.28 Water 47,925 0.01 4.19 95,850 0.01 3.29

TOTAL USD 20,951,777 4.19 USD/t 32,885,973 3.29 USD/t

Unit Cost Gold Ounces Produced 163 USD/oz 128 USD/oz Net Gold Cash Revenue 1,037 USD/oz 1,072 USD/oz

The detailed power consumption estimate is based on the equipment noted in the equipment list and the installed power with estimates of the operating power draft and operating time. The process power consumption is summarized in Table 21.6.

Table 21.6Operating Cost Estimate

Heap Leach Power & Energy Section Phase I (5 Mtpa) Phase II (10 Mtpa)

Installed Power

kW

Power Demand

kW

Annual Cost (US$)

kWh/t Installed Power

kW

Power Demand

kW

Annual Cost (US$)

kWh/t

Area 10 - Primary Crushing 604 453 193,629 0.5532 994 745 315,239 0.4503 Area 13 - Secondary Crushing 961 805 357,263 1.0208 1,763 1,491 664,055 0.9486 Area 15 - Tertiary Crushing 2,049 1,757 785,855 2.2453 3,289 2,907 1,300,250 1.8575 Area 17 - Lime Addition 1,449 1,009 451,180 1.2891 1,345 1,009 451,180 0.6445 Area 19 - Ore Stacking 449 337 150,685 0.4306 748 547 244,545 0.3493 Area 20 - Heap Leach 646 266 155,470 0.4442 646 266 155,470 0.2221 Area 1 - Carbon Adsorption 6 4 466 0.0013 228 171 82,799 0.1183 Area 2 - Acid Wash 8 6 2,480 0.0071 8 6 2,480 0.0035 Area 3 - Carbon Strip 2 2 328 0.0009 2 2 328 0.0005 Area 4 - Strip Solution Handling 413 305 117,864 0.5082 413 305 177,864 0.2541 Area 5 - EW & Refining 171 128 50,206 0.1434 221 165 72,124 0.1030 Area 6 - Carbon Regeneration & Handling 148 111 63,961 0.1827 148 111 63,961 0.0914 Area 7 Reagent Mix / Storage 14 9 5,173 0.0148 14 9 5,173 0.0074 Area 30 Barren Solution Pumping 953 620 362,124 1.0345 953 620 362,124 0.5176 Area 90 - Auxiliary Equipment - - - - - - - - Total Heap Lech Power Cost (per annum) 2,756,695 3,897,591 Power Consumption 7.88 5.57 Cost per Tonne Ore 0.55 0.39

The labor cost estimate (Phase I and Phase II) for process operations are shown in Table 21.7 and Table 21.8 respectively. The labor rates and burden are based on the information provided by Lydian. General and Administrative labor estimates are provided in Table 21.9 and Table 21.10.



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

Position

Category

Number Persons Req'd.

Annual Salary (US$)

Annual Base Cost

(US$)

On Cost

Factor

Annual Oncost (US$)

Annual Cost

(US$) (US$/t) OPERATIONS Process Manager E 1 225,000 225,000 11,710 236,710 236,710 Process Superintendent N 1 60,000 60,000 3,460 63,460 63,460 General Foreman N 1 20,000 20,000 1,460 21,460 21,460 Reagent Operators N 6 7,000 42,000 4,858 46,858 281,146 General Laborers N 10 7,000 70,000 8,096 78,096 780,960 Sub-total 18 417,000 446,582 1,383,734 0.28 Shift Supervisors N 4 18,000 72,000 5,438 77,438 309,754 Crusher Operators N 10 9,000 90,000 9,096 99,096 990,960 Heap Leach Operators N 8 9,000 72,000 7,277 79,277 634,214 ADR Operators N 8 9,000 72,000 7,277 79,277 634,214 Gold Room Operators N 3 9,000 27,000 2,729 29,729 89,186 Sub total 33 333,000 364,817 2,658,329 0.53

Operations Total 51 750,000 811,399 4,042,063 0.81 Plant Metallurgist N 1 20,000 20,000 1,460 21,460 21,460 Metallurgical Technicians N 2 10,000 20,000 1,919 21,919 43,838 Data Entry Clerk N 1 5,000 5,000 710 5,710 5,710

Sub-total 4 45,000 49,088 71,008 0.01 LABORATORY Lab Manager N 1 50,000 50,000 2,960 52,960 52,960 Chief Chemist N 1 30,000 30,000 1,960 31,960 31,960 Chemists N 2 18,000 36,000 2,719 38,719 77,438 Assayers N 4 8,000 32,000 3,438 35,438 141,754 Sample prep. Operators N 4 8,000 32,000 3,438 35,438 141,754

Laboratory Total 12 180,000 194,515 445,865 0.09 MAINTENANCE Maintenance Superintendent N 1 50,000 50,000 2,960 52,960 52,960 Mechanical Foreman N 1 12,000 12,000 1,060 13,060 13,060 Electrical Foreman N 1 12,000 12,000 1,060 13,060 13,060 Instrument Technicians N 1 9,000 9,000 910 9,910 9,910 Fitters N 6 9,000 54,000 5,458 59,458 356,746 Welder/boilermakers N 6 9,000 54,000 5,458 59,458 356,746 Electricians N 6 9,000 54,000 5,458 59,458 356,746

Maintenance Total 21 245,000 267,361 1,159,225 0.23

GRAND TOTAL 88 1,220,000 1,322,364 5,718,161 1.14 TOTAL HEAP LEACH FIXED COST (per annum) 5,718,161 COST PER TONNE ORE 1.14



Table 21.8

Operating Cost Estimate - Plant Labor (Phase II)

Position

Category

Number Persons Req'd.

Annual Salary (USD)

Annual Base Cost

(US$)

On Cost

Factor

Annual Oncost (US$)

Annual Cost

(US$) (US$/t) OPERATIONS Process Manager E 1 225,000 225,000 11,710 236,710 236,710 Process Superintendent N 1 60,000 60,000 3,460 63,460 63,460 General Foreman N 1 20,000 20,000 1,460 21,460 21,460 Reagent operators N 6 7,000 42,000 4,858 46,858 281,146 General laborers N 10 7,000 70,000 8,096 78,096 780,960 Sub-total 18 417,000 446,582 1,383,734 0.28 Shift supervisors N 4 18,000 72,000 5,438 77,438 309,754 Crusher operators N 12 9,000 108,000 10,915 118,915 1,426,982 Heap Leach operators N 8 9,000 72,000 7,277 79,277 634,214 ADR operators N 8 9,000 72,000 7,277 79,277 634,214 Gold Room operators N 3 9,000 27,000 2,729 29,729 89,186 Sub total 35 351,000 384,636 3,094,351 0.62

Operations Total 53 768,000 831,218 4,478,086 0.90 Plant Metallurgist N 1 20,000 20,000 1,460 21,460 21,460 Metallurgical Technicians N 2 10,000 20,000 1,919 21,919 43,838 Data Entry Clerk N 1 5,000 5,000 710 5,710 5,710

Sub-total 4 45,000 49,088 71,008 0.01 LABORATORY Lab Manager N 1 50,000 50,000 2,960 52,960 52,960 Chief Chemist N 1 30,000 30,000 1,960 31,960 31,960 Chemists N 2 18,000 36,000 2,719 38,719 77,438 Assayers N 4 8,000 32,000 3,438 35,438 141,754 Sample prep. operators N 4 8,000 32,000 3,438 35,438 141,754

Laboratory Total 12 180,000 194,515 445,865 0.09 MAINTENANCE Maintenance Superintendent N 1 50,000 50,000 2,960 52,960 52,960 Mechanical Foreman N 1 12,000 12,000 1,060 13,060 13,060 Electrical Foreman N 1 12,000 12,000 1,060 13,060 13,060 Instrument Technicians N 1 9,000 9,000 910 9,910 9,910 Fitters N 6 9,000 54,000 5,458 59,458 356,746 Welder/boilermakers N 6 9,000 54,000 5,458 59,458 356,746 Electricians N 6 9,000 54,000 5,458 59,458 356,746

Maintenance Total 21 245,000 267,361 1,159,225 0.23

GRAND TOTAL 90 1,238,000 1,342,183 6,154,183 1.23 TOTAL HEAP LEACH FIXED COST (per annum) 6,154,183 COST PER TONNE ORE 1.23



Table 21.9

Operating Cost Estimate - Yerevan Office Administration Labor

Annual Salary (US$) Number On-Cost

Annual Cost

(US$) (US$/t) Labor General Manger 110,000 1 5,960 115,960 0.02 Office Manager 20,000 1 1,460 21,460 0.00 Administration Assistant 6,000 1 760 6,760 0.00 Senior Accountant 25,000 1 1,710 26,710 0.01 Accountant 15,000 2 1,210 32,419 0.01 Purchasing Officer 16,000 1 1,260 17,260 0.00 Logistics 7,000 2 810 15,619 0.00 Office Clerk 5,000 1 710 5,710 0.00 Translator 8,000 1 860 8,860 0.00 Full time GIS / Geographer 22,500 1 1,585 24,085 0.00 Database/IT 20,000 1 1,460 21,460 0.00 Cleaner 3,500 1 635 4,135 0.00 Public relation officer 15,000 1 1,210 16,210 0.00 Security Officers 15,000 2 1,210 32,419 0.01

Sub Total 17 349,063 0.07

Cost/day

(US$) Accommodation Yerevan Office Staff 15 16 87,600 0.02 Services Communications, consumables, 250,000 1 250,000 0.05 consultants, insurances, non-scheduled flights Toronto Representative Office 100,000 1 100,000 0.02 Sub Total 18 437,600 0.09 TOTAL ADMINISTRATION COST (per annum) 786,663 COST PER TONNE ORE 0.16



Table 21.10

Operating Cost Estimate - Site Administration Labor

Cost Center Category

Annual Salary (US$) Number

Salary On

Cost

Annual Cost (US$)

Site Labor Site Manager E 250,000 1 12,960 262,960 Environmental / OH&S Officer N 102,000 1 5,560 107,560 Environmental Group N 37,500 1 2,335 Site Nurse N 7,000 4 810 31,238 Camp Manager N 38,000 1 2,360 40,360 Cleaners N 5,000 2 710 11,419 Cooks N 12,000 3 1,060 39,179 Canteen Workers N 5,000 4 710 22,838 Site Security N 6,000 24 760 162,230

Sub Total 41 492,715

Cost/day

(US$) Accommodation Mine Staff 15 41 224,475 Services Communications, consumables, first aid, 150,000 1 150,000 light vehicles, insurances, misc. maintenance, non scheduled flights Site Office Overheads 250,000 1 250,000 TOTAL ADMINISTRATION COST (per annum) 1,117,190 COST PER TONNE ORE 0.22



Reagent cost estimates are shown in Table 21.11. The reagent consumption rates are based on similar projects.

Table 21.11

Operating Cost Estimate - Heap Leach ConsumablesConsumption Consumption

Section Usage Unit Cost Rate Usage Unit Cost Rate(t/a) (USD/t) kg/t (USD) (USD/t) (t/a) (USD/t) kg/t (USD) (USD/t)

1. Crushing & Agglomeration

Liners - primary 6.0 0.012 360,000 0.07 6.0 0.012 720,000 0.07Liners - secondary 6.0 0.012 360,000 0.07 6.0 0.012 720,000 0.07Liners - tertiary 6.0 0.012 360,000 0.07 6.0 0.012 720,000 0.07Liners - tertiary 6.0 0.012 360,000 0.07 6.0 0.012 720,000 0.07Loader 1,666,667 $0.30 500,000 0.10 3,333,333 $0.30 1,000,000 0.10Rubber lining

Sub Total 1,940,000 0.39 3,880,000 0.392. Heap Leach

HDPE Liners 0 0 0.00 0 0.00 0 0 0.00 0 0.00Poly pipe 0 0 0.00 0 0.00 0 0 0.00 0 0.00Plastic Pipe 0 0 0.00 0 0.00 0 0 0.00 0 0.00Drippers 0 0 0.00 0 0.00 0 0 0.00 0 0.00

Sub Total 0 0.00 0 0.003. ADR

Lime 4,000 200 0.80 800,000 0.16 8,000 200 0.80 1,600,000 0.16Sodium Cyanide 2,000 2,850 0.40 5,700,000 1.14 4,000 2,850 0.40 11,400,000 1.14Carbon 100 3,200 0.02 320,000 0.06 200 3,200 0.02 640,000 0.06Sodium Cyanide 50 2,850 0.01 142,500 0.03 100 2,850 0.01 285,000 0.03Sodium Hydroxide 100 350 0.02 35,000 0.01 200 350 0.02 70,000 0.01Hydrochloric Acid 450 300 0.09 135,000 0.03 900 300 0.09 270,000 0.03

Sub Total 7,132,500 1.43 14,265,000 1.434. Goldroom Per strip Per strip

Boiler Diesel (litres) 155 1,200 0.031 186,451 0.04 311 1,200 0.031 372,902 0.04Steel Wool 24 2,000 0.005 48,533 0.01 49 2,000 0.005 97,067 0.01Silca Sand 0.8 75 0.0002 59 0.00 1.6 75 0.000 119 0.00Borax 1.6 1,025 0.01 1,620 0.00 3.2 1,025 0.01 3,239 0.00Nitre 0.8 600 0.002 474 0.00 1.6 600 0.002 948 0.00Sodium Carbonate 0.8 240 0.001 190 0.00 1.6 240 0.001 379 0.00

Sub Total 237,327 0.05 474,654 0.055. Laboratory

Process Plant samples 10 15,000 150,000 0.03 10 15,000 150,000 0.02Grade Control samples 10 25,000 250,000 0.05 10 25,000 250,000 0.03

Sub Total 400,000 0.08 400,000 0.04

TOTAL HEAP LEACH CONSUMABLES COST (per annum) 9,709,827 19,019,654

COST PER TONNE ORE 1.94 1.90

CostAnnual

CostAnnual



Wear material cost estimates are provided in Table 21.12.

Table 21.12 Maintenance

Tonnage 5 Mtpa 10 Mtpa Source of Information Total Equipment Installed Cost, US$ 69,348,296 95,532,249 Capital Cost Estimate Maintenance Percentage, % 5.00 5.00 Other Projects Annual Maintenance Cost, US$ 3,467,415 4,776,612 Calculated Cost per Tonne, US$/t 0.69 0.48 Calculated The process water cost estimate, shown in Table 21.13, is based on the consumption at similar operation, and the delivered water price of $0.05 per tonne.

Table 21.13 Water

Tonnage 5 Mtpa 10 Mtpa Tonnes water per tonne ore 0.1917 0.1917 Cost, US$ per tonne water 0.05 0.05 Annual Maintenance Cost 47,925 95,850 Cost per tonne 0.01 0.01



22.0 ECONOMIC ANALYSIS

A pre-income tax economic analysis model was prepared. The model uses the production and cost estimates shown earlier in this report. Costs are in 2011 constant dollars. The economic analysis uses a gold sales price US$ 1,200 per ounce and a silver sales price of US$ 20.00 per ounce and plant estimated recoveries of 85 percent for gold for all PMM grades processed and 40 percent for silver based on a nominal average grade. Operating cost estimates and values for key design parameters that have been presented in previous sections of the PEA were used as required. The economic analysis was done on an all equity financing basis. Four different scenarios were prepared, they are;

1. Owner operating mining 2. Contract mining 3. Owner operating mining with Erato resource ounces 4. Contract mining with Erato resource ounces

22.1 Owner Operating Mining Case Table 22.1 shows the project's owner operating mining pre-income tax internal rate of return and the project's pre-income tax net present values at discount rates from 0 to 20 percent.

Table 22.1 Owner Operated Mining

Economic Analysis Summary Pre-Tax Internal Rate of Return (IRR), % 39.5 Pre-Tax Net Present Values US$ x 1000 @ 0 % discount rate 747,329 @ 5 % discount rate 493,594 @ 10 % discount rate 325,938 @ 15 % discount rate 213,021 @ 20 % discount rate 135,704

Table 22.2 summarizes the project's revenue, costs and pre-income tax cash flow and also shows the values in units of resource processed and saleable gold ounces.



Table 22.2

Owner Operated Mining Economic Analysis Summary - Before Tax Cash Flow and Unit Values

$US x 1000 $US/t

Resource $US/oz GoldMine Gate Value of All Resource Net of Transportation and Refining 1,666,157 30.15 1,194.75 Mining Operating Cost (378,238) (6.84) (271.22)Processing Cost (192,943) (3.49) (138.35)General & Administration (13,616) (0.25) (9.76)Cash Operating Cost (584,797) (10.58) (419.34)Cash Operating Cash Flow 1,081,360 19.57 775.41 Capital Cost including Pre-Production Development (378,752) (6.85) (271.59)Pre-Income Tax Cash Flow 702,608 12.71 503.82

Table 22.3 shows the annual economic analysis for owner Operated Mining.



Table 22.3Cash Flow Schedule (Owner Operating Case)

Year -2 Year -1 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10 Year 11 Total

Mine ProductionTOTAL OPEN PIT Resources Mined tonnes 4,164,020 5,025,234 7,419,931 10,026,901 10,001,776 9,636,299 8,993,610 55,267,771

Gold g/t 1.164 0.857 0.867 1.004 0.897 0.970 0.785 0.923 Silver g/t 3.630 3.630 3.630 3.630 3.630 3.630 3.630 3.630

Waste tonnes 5,000,016 10,000,006 17,308,974 27,188,923 27,267,096 26,594,351 22,609,185 135,968,549 Total Mined tonnes 9,164,035 15,025,240 24,728,905 37,215,823 37,268,871 36,230,650 31,602,795 191,236,320

Processing Plants ProductionTOTAL RESOURCE PROCESSED Resources Processed tonnes 3,750,000 5,000,000 7,500,000 10,000,000 10,000,000 10,000,000 9,017,771 55,267,771

Gold g/t 1.159 0.862 0.874 1.005 0.899 0.970 0.785 0.923 Silver g/t 3.630 3.630 3.630 3.630 3.630 3.630 3.630 3.630

TOTAL RECOVERY Gold Recovery % 85% 85% 85% 85% 85% 85% 85% 85%Silver Recovery % 40% 40% 40% 40% 40% 40% 40% 40%

TOTAL METAL RECOVERABLE Gold Recoverable ounces 118,780 117,818 179,066 274,605 245,731 265,103 193,463 1,394,566 Silver Recoverable ounces 175,061 233,414 350,122 466,829 466,829 466,829 420,976 2,580,059

Silver/Gold Ratio 1.5 2.0 2.0 1.7 1.9 1.8 2.2 1.9 Cumulative Recoverable Gold ounces 118,780 236,598 415,663 690,268 935,999 1,201,103 1,394,566 Cumulative Recoverable Silver ounces 175,061 408,475 758,597 1,225,426 1,692,255 2,159,083 2,580,059 Initial Year Percent of Recoverable Metal Recovery Factor % 75 75 75 75 75 75 100 Second Year Percent of Recoverable Metal Recovery Factor % 25 25 25 25 25 25 25 Gold Recovered ounces 89,085 118,058 163,754 250,720 252,950 260,260 259,739 1,394,566 Silver Recovered ounces 131,296 218,826 320,945 437,652 466,829 466,829 537,683 2,580,059



Table 22.3 continuedCash Flow Schedule (Owner Operating Case)


RevenueCOMMODITY PRICES Gold $/ounce 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00

Silver $/ounce 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00

GROSS SALES Gold $ 106,901,877 141,669,867 196,504,377 300,864,323 303,539,588 312,312,162 311,686,661 1,673,478,855 Silver $ 2,625,912 4,376,520 6,418,897 8,753,041 9,336,577 9,336,577 10,753,656 51,601,179

REFINING & TRANSPORTATION Gold $ (467,696) (619,806) (859,707) (1,316,281) (1,327,986) (1,366,366) (1,363,629) (7,321,470) $/ounce (5.25) (5.25) (5.25) (5.25) (5.25) (5.25) (5.25) (5.25)

Silver $ (97,256) (162,093) (237,737) (324,187) (345,799) (345,799) (398,284) (1,911,155) $/ounce (0.74) (0.74) (0.74) (0.74) (0.74) (0.74) (0.74) (0.74)

PAYABLES Gold % 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 Silver % 90.00 90.00 90.00 90.00 90.00 90.00 90.00 90.00

NET REVENUE Gold $ 106,434,181 141,050,061 195,644,671 299,548,042 302,211,603 310,945,796 310,323,031 1,666,157,385 $/ounce 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75

Silver $ 2,275,790 3,792,984 5,563,044 7,585,969 8,091,700 8,091,700 9,319,835 44,721,022 $/ounce 17.33 17.33 17.33 17.33 17.33 17.33 17.33 17.33

Total Net Revenue $ 108,709,972 144,843,045 201,207,714 307,134,010 310,303,302 319,037,496 319,642,866 1,710,878,407

Operating CostsMining Cost $ 21,159,550 28,935,863 45,227,177 65,899,474 70,918,741 74,024,449 72,072,433 378,237,688

$/tonne 5.64 5.79 6.03 6.59 7.09 7.40 7.99 6.84 Processing $ 17,249,999 21,000,000 26,325,000 32,900,000 32,900,000 32,900,000 29,668,468 192,943,467

$/tonne 4.60 4.20 3.51 3.29 3.29 3.29 3.29 3.49 General & Administration $ 1,912,500 1,950,000 1,950,000 2,000,000 2,000,000 2,000,000 1,803,554 13,616,054

$/tonne 0.51 0.39 0.26 0.20 0.20 0.20 0.20 0.25 Total Operating Cost $ - 40,322,049 51,885,863 73,502,177 100,799,474 105,818,741 108,924,449 103,544,456 584,797,209

Operating ProfitOperating Profit $ 68,387,923 92,957,183 127,705,537 206,334,536 204,484,561 210,113,047 216,098,411 - 1,126,081,198

$/gold ounce 767.67 787.38 779.86 822.97 808.40 807.32 831.98 807.48

Capital CostsMining Cost $ 9,068,273 54,520,535 2,829,335 29,678,135 48,670,030 3,789,335 3,465,335 1,899,335 (7,992,126) 145,928,186 Buy out Newmont Cost 15,600,000 15,600,000 Process Plant Direct Cost $ 38,231,631 38,231,631 26,183,953 102,647,215 Process Plant Indirect Cost & Contingency $ 28,619,971 28,619,971 17,219,184 74,459,127 Leach Pads $ 19,756,625 12,641,830 32,398,454 Closure and Reclamation $ 5,282,708 1,349,521 935,031 151,877 7,719,137 Total Capital Cost $ 66,851,602 95,676,500 70,120,535 2,829,335 85,723,102 48,670,030 3,789,335 3,465,335 1,899,335 -2,709,418 1,349,521 935,031 151,877 378,752,120

Pre-Income Tax Cash FlowPre-Income Tax Cash Flow $ (66,851,602) (95,676,500) (1,732,612) 90,127,848 41,982,435 157,664,506 200,695,226 206,647,712 214,199,076 2,709,418 (1,349,521) (935,031) (151,877) 747,329,078

$/gold ounce (19.45) 763.42 256.38 628.85 793.42 794.00 824.67 535.89 Cumulative Pre-income Tax Cash Flow $ (66,851,602) (162,528,102) (164,260,714) (74,132,866) (32,150,431) 125,514,075 326,209,301 532,857,013 747,056,089 749,765,507 748,415,986 747,480,955 747,329,078

(2) (1) 1 2 3 4 5 6 7 Payback, operating years 3.2 4.20 4

Pre-Income Net Present Values and Rate of ReturnNPV @ 0% discount rate $ 747,329,078 NPV @ 5% discount rate $ 493,593,525 NPV @ 10% discount rate $ 325,938,478 NPV @ 15% discount rate $ 213,021,295 NPV @ 20% discount rate $ 135,703,520

IRR 39.5%



Sensitivity Analysis The project's pre-income tax internal rate of return sensitivity relative to incremental changes in metal prices, recoveries, grades and costs are shown in Table 22.4 and Figure 22.1

Table 22.4Rate of Return Sensitivity

Percent Changes -20% -15% -10% -5% Base 5% 10% 15% 20%

Gold Price 24.8 28.8 32.6 36.1 39.5 42.8 46.0 49.0 51.9 Silver Price 39.2 39.3 39.3 39.4 39.5 39.6 39.7 39.8 39.9 Gold Recovery 24.9 28.9 32.6 36.1 39.5 42.8 45.9 49.0 51.9 Silver Recovery 39.2 39.3 39.4 39.4 39.5 39.6 39.7 39.8 39.9 Gold Grade 24.9 28.9 32.6 36.1 39.5 42.8 45.9 49.0 51.9 Silver Grade 39.5 39.5 39.5 39.5 39.5 39.5 39.6 39.6 39.6 Operating Cost 44.2 43.0 41.9 40.7 39.5 38.3 37.1 35.9 34.6 Capital Cost 49.9 47.0 44.3 41.8 39.5 37.4 35.4 33.5 31.8

Amulsar Gold Project Pre-IncomeTax Sensitivity IRR (Owner Operator)

0.0%

10.0%

20.0%

30.0%

40.0%

50.0%

60.0%

-20% -15% -10% -5% Base 5% 10% 15% 20%

Variance from Base Case

IRR

Pre

-Tax

Cal

cula

tion

Gold Price Silver Price Gold Recovery Silver Recovery Operating Cost Capital Cost Gold Grade Silver Grade

Figure 22.1 - Amulsar Gold Project Pre-Tax Sensitivity IRR (Owner Operator)

The project's pre-income tax net present value, using a five percent discount rate, sensitivity relative to incremental changes in metal prices, recoveries, grades and costs are shown in Table 22.5 and Figure 22.2.



Table 22.5

Owner Operated Mining NPV Sensitivity (US$ X 1000)

Percent Changes -20% -15% -10% -5%

Base at 5% 5% 10% 15% 20%

Gold Price 250,515 311,284 372,054 432,824 493,594 554,363 615,133 675,903 736,673Silver Price 486,877 488,556 490,235 491,914 493,594 495,273 496,952 498,631 500,311Gold Recovery 251,578 312,082 372,586 433,090 493,594 554,097 614,601 675,105 735,609Silver Recovery 487,125 488,742 490,359 491,976 493,594 495,211 496,828 498,445 500,062Gold Grade 251,578 312,082 372,586 433,090 493,594 554,097 614,601 675,105 735,609Silver Grade 493,076 493,205 493,335 493,464 493,594 493,723 493,853 493,982 494,112Operating Cost 578,794 557,494 536,194 514,894 493,594 472,293 450,993 429,693 408,393Capital Cost 558,158 542,017 525,876 509,735 493,594 477,452 461,311 445,170 429,029

Amulsar Gold Project Pre-Income Tax Sensitivity NPV@5% (Owner Operator)

0

100,000,000

200,000,000

300,000,000

400,000,000

500,000,000

600,000,000

700,000,000

800,000,000

-20% -15% -10% -5% Base 5% 10% 15% 20%


NPV

@ 5

% P

re-T

ax C

alcu

latio

n - U

S$


Figure 22.2 - Amulsar Gold Project Pre-Tax Sensitivity NPV@5% (Owner Operator)

As seen in Tables 22.4 and 22.5, the project's owner operating mining case pre-income tax rate of return is 39.5 percent and the project's owner operating mining case pre-income tax net present value at a 5 percent discount rate is US$ 493.6 million. A ten percent increase in the gold price increases the estimated rate of return to 46.0 percent and increases the project's net present value, at a 5 percent discount rate, to US$ 615.1 million.



22.2 Contract Mining Case Table 22.6 shows the project's contract mining pre-income tax internal rate of return and the project's pre-income tax net present values at discount rates from 0 to 20 percent.

Table 22.6 Contract Mining

Economic Analysis Summary Internal Rate of Return (IRR), % 45.4 Net Present Values US$ x 1000 @ 0 % discount rate 759,834 @ 5 % discount rate 514,528 @ 10 % discount rate 350,222 @ 15 % discount rate 237,970 @ 20 % discount rate 159,938


Table 22.7 Contract Mining

Economic Analysis Summary - Before Tax Cash Flow and Unit Values

$US x 1000 $US/t

Resource $US/oz GoldMine Gate Value of All Resource Net of Transportation and Refining 1,666,157 30.15 1,194.75 Mining Operating Cost (489,893) (8.86) (351.29)Processing Cost (192,943) (3.49) (138.35)General & Administration (13,616) (0.25) (9.76)Cash Operating Cost (696,453) (12.60) (499.40)Cash Operating Cash Flow 969,704 17.55 695.35 Capital Cost including Pre-Production Development (254,592) (4.61) (182.56)Pre-Income Tax Cash Flow 715,113 12.94 512.79

Table 22.8 shows the annual economic analysis.



Table 22.8Cash Flow Schedule (Contract Mining)


Mine ProductionTOTAL OPEN PIT Resources Mined tonnes 4,164,020 5,025,234 7,419,931 10,026,901 10,001,776 9,636,299 8,993,610 55,267,771

Gold g/t 1.164 0.857 0.867 1.004 0.897 0.970 0.785 0.923 Silver g/t 3.630 3.630 3.630 3.630 3.630 3.630 3.630 3.630

Waste tonnes 5,000,016 10,000,006 17,308,974 27,188,923 27,267,096 26,594,351 22,609,185 - 135,968,549 Total Mined tonnes 9,164,035 15,025,240 24,728,905 37,215,823 37,268,871 36,230,650 31,602,795 - 191,236,320

Processing Plants ProductionTOTAL RESOURCE PROCESSED Resources Processed tonnes 3,750,000 5,000,000 7,500,000 10,000,000 10,000,000 10,000,000 9,017,771 - 55,267,771

Gold g/t 1.159 0.862 0.874 1.005 0.899 0.970 0.785 - 0.923 Silver g/t 3.630 3.630 3.630 3.630 3.630 3.630 3.630 - 3.630

TOTAL RECOVERY Gold Recovery % 85% 85% 85% 85% 85% 85% 85% 85%Silver Recovery % 40% 40% 40% 40% 40% 40% 40% 40%

TOTAL METAL RECOVERABLE Gold Recoverable ounces 118,780 117,818 179,066 274,605 245,731 265,103 193,463 - 1,394,566 Silver Recoverable ounces 175,061 233,414 350,122 466,829 466,829 466,829 420,976 - 2,580,059

Silver/Gold Ratio 1.5 2.0 2.0 1.7 1.9 1.8 2.2 1.9 Cumulative Recoverable Gold ounces 118,780 236,598 415,663 690,268 935,999 1,201,103 1,394,566 Cumulative Recoverable Silver ounces 175,061 408,475 758,597 1,225,426 1,692,255 2,159,083 2,580,059 Initial Year Percent of Recoverable Metal Recovery Factor % 75 75 75 75 75 75 100 Second Year Percent of Recoverable Metal Recovery Factor % 25 25 25 25 25 25 25 Gold Recovered ounces 89,085 118,058 163,754 250,720 252,950 260,260 259,739 1,394,566 Silver Recovered ounces 131,296 218,826 320,945 437,652 466,829 466,829 537,683 2,580,059



Table 22.8 continuedCash Flow Schedule (Contract Mining)


RevenueCOMMODITY PRICES Gold $/ounce 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00

Silver $/ounce 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00

GROSS SALES Gold $ 106,901,877 141,669,867 196,504,377 300,864,323 303,539,588 312,312,162 311,686,661 1,673,478,855 Silver $ 2,625,912 4,376,520 6,418,897 8,753,041 9,336,577 9,336,577 10,753,656 51,601,179

REFINING & TRANSPORTATION Gold $ (467,696) (619,806) (859,707) (1,316,281) (1,327,986) (1,366,366) (1,363,629) (7,321,470) $/ounce (5.25) (5.25) (5.25) (5.25) (5.25) (5.25) (5.25) (5.25)

Silver $ (97,256) (162,093) (237,737) (324,187) (345,799) (345,799) (398,284) (1,911,155) $/ounce (0.74) (0.74) (0.74) (0.74) (0.74) (0.74) (0.74) (0.74)

PAYABLES Gold % 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 Silver % 90.00 90.00 90.00 90.00 90.00 90.00 90.00 90.00

NET REVENUE Gold $ 106,434,181 141,050,061 195,644,671 299,548,042 302,211,603 310,945,796 310,323,031 1,666,157,385 $/ounce 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75

Silver $ 2,275,790 3,792,984 5,563,044 7,585,969 8,091,700 8,091,700 9,319,835 44,721,022 $/ounce 17.33 17.33 17.33 17.33 17.33 17.33 17.33 17.33

Total Net Revenue $ 108,709,972 144,843,045 201,207,714 307,134,010 310,303,302 319,037,496 319,642,866 - 1,710,878,407

Operating CostsMining Cost $ 26,578,668 37,012,717 58,254,209 85,603,904 92,206,792 96,264,854 93,972,353 - 489,893,497

$/tonne 7.09 7.40 7.77 8.56 9.22 9.63 10.42 8.86 Processing $ 17,249,999 21,000,000 26,325,000 32,900,000 32,900,000 32,900,000 29,668,468 192,943,467

$/tonne 4.60 4.20 3.51 3.29 3.29 3.29 3.29 3.49 General & Administration $ 1,912,500 1,950,000 1,950,000 2,000,000 2,000,000 2,000,000 1,803,554 13,616,054

$/tonne 0.51 0.39 0.26 0.20 0.20 0.20 0.20 0.25 Total Operating Cost $ - 45,741,168 59,962,717 86,529,209 120,503,904 127,106,792 131,164,854 125,444,375 696,453,019

Operating ProfitOperating Profit $ 62,968,804 84,880,328 114,678,505 186,630,106 183,196,510 187,872,642 194,198,492 - 1,014,425,388

$/gold ounce 706.84 718.97 700.31 744.38 724.24 721.86 747.67 727.41

Capital CostsMining Cost $ 9,085,196 1,202,258 1,202,258 602,258 2,100,554 566,258 566,258 1,916,258 4,526,514 21,767,812 Buy out Newmont Cost 15,600,000 15,600,000 Process Plant Direct Cost $ 38,231,631 38,231,631 26,183,953 102,647,215 Process Plant Indirect Cost & Contingency $ 28,619,971 28,619,971 17,219,184 74,459,127 Leach Pads $ 19,756,625 12,641,830 32,398,454 Closure and Reclamation $ 5,282,708 1,349,521 935,031 151,877 7,719,137 Total Capital Cost $ 66,851,602 95,693,423 16,802,258 1,202,258 56,647,226 2,100,554 566,258 566,258 1,916,258 9,809,222 1,349,521 935,031 151,877 254,591,745

Pre-Income Tax Cash FlowPre-Income Tax Cash Flow $ (66,851,602) (95,693,423) 46,166,546 83,678,070 58,031,280 184,529,553 182,630,252 187,306,384 192,282,233 (9,809,222) (1,349,521) (935,031) (151,877) 759,833,643

$/gold ounce 518.23 708.79 354.38 736.00 722.00 719.69 740.29 544.85 Cumulative Pre-income Tax Cash Flow $ (66,851,602) (162,545,025) (116,378,479) (32,700,409) 25,330,871 209,860,423 392,490,676 579,797,060 772,079,293 762,270,072 760,920,551 759,985,520 759,833,643

(2) (1) 1 2 3 4 5 6 7 Payback, operating years 2.6 4 4


IRR 45.4%




Table 22.9Contract Mining

Rate of Return Sensitivity Percent Changes -20% -15% -10% -5% Base 5% 10% 15% 20%


Amulsar Gold Project Pre-IncomeTax Sensitivity IRR (Contract Mining)

0.0%

10.0%

20.0%

30.0%

40.0%

50.0%

60.0%

70.0%

-20% -15% -10% -5% Base 5% 10% 15% 20%


IRR

Pre-

Tax

Cal

cula

tion


Figure 22.3 - Amulsar Gold Project Pre-Tax Sensitivity IRR (Contract Mining)




Table 22.10

NPV Sensitivity (US$ X 1000) Percent

Changes -20% -15% -10% -5% Base at

5% 5% 10% 15% 20% Gold Price 271,449 332,219 392,989 453,759 514,528 575,298 636,068 696,838 757,607Silver Price 507,811 509,491 511,170 512,849 514,528 516,208 517,887 519,566 521,245Gold Recovery 272,513 333,017 393,520 454,024 514,528 575,032 635,536 696,040 756,544Silver Recovery 508,060 509,677 511,294 512,911 514,528 516,145 517,762 519,379 520,996Gold Grade 272,513 333,017 393,520 454,024 514,528 575,032 635,536 696,040 756,544Silver Grade 508,060 509,677 511,294 512,911 514,528 516,145 517,762 519,379 520,996Operating Cost 615,836 590,509 565,182 539,855 514,528 489,201 463,875 438,548 413,221Capital Cost 558,799 547,731 536,664 525,596 514,528 503,461 492,393 481,325 470,258

Amulsar Gold Project Pre-Income Tax Sensitivity NPV@5% (Contract Mining)

0

100,000,000

200,000,000

300,000,000

400,000,000

500,000,000

600,000,000

700,000,000

800,000,000

-20% -15% -10% -5% Base 5% 10% 15% 20%


NPV

@ 5

% P

re-T

ax C

alcu

latio

n - U

S$


Figure 22.4 - Amulsar Gold Project Pre-Tax Sensitivity NPV@5% (Contract Mining)

As seen in Tables 22.9 and 22.10, the project's contract mining case pre-income tax rate of return is 45.4 percent and the project's contract mining case pre-income tax net present value at a 5 percent discount rate is US$ 514.5 million. A ten percent increase in the gold price increases the estimated rate of return to 52.1 percent and increases the project's net present value, at a 5 percent discount rate, to US$ 636.1 million.



22.3 Owner Operating Mining with Erato Case Table 22.11 shows the project's owner operating mining with Erato before tax internal rate of return and the projects before tax net present values at discount rates from 0 to 20 percent.

Table 22.11 Owner Operated Mining with Erato

Economic Analysis Summary Internal Rate of Return (IRR), % 40.8 Net Present Values US$ x 1000 @ 0 % discount rate 956,362 @ 5 % discount rate 614,659 @ 10 % discount rate 397,665 @ 15 % discount rate 256,393 @ 20 % discount rate 162,415


Table 22.12 Owner Operated Mining with Erato

Economic Analysis Summary - Before Tax Cash Flow

$US x 1000 $US/t





Table 22.13Cash Flow Schedule (Owner Operating with Erato)

Year -2 Year -1 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10 Year 11 Year 12 Year 13 Total

Mine ProductionTOTAL OPEN PIT Resources Mined tonnes 4,164,020 5,025,234 7,419,931 10,026,901 10,001,776 9,636,299 8,993,610 9,600,000 9,600,000 74,467,771

Gold g/t 1.164 0.857 0.867 1.004 0.897 0.970 0.785 0.740 0.740 0.876 Silver g/t 3.630 3.630 3.630 3.630 3.630 3.630 3.630 3.630 3.630 3.630

Waste tonnes 5,000,016 10,000,006 17,308,974 27,188,923 27,267,096 26,594,351 22,609,185 38,400,000 38,400,000 212,768,549 Total Mined tonnes 9,164,035 15,025,240 24,728,905 37,215,823 37,268,871 36,230,650 31,602,795 48,000,000 48,000,000 287,236,320

Processing Plants ProductionTOTAL RESOURCE PROCESSED Resources Processed tonnes 3,750,000 5,000,000 7,500,000 10,000,000 10,000,000 10,000,000 9,017,771 9,600,000 9,600,000 74,467,771

Gold g/t 1.159 0.862 0.874 1.005 0.899 0.970 0.785 0.740 0.740 0.876 Silver g/t 3.630 3.630 3.630 3.630 3.630 3.630 3.630 3.630 3.630 3.630

TOTAL RECOVERY Gold Recovery % 85% 85% 85% 85% 85% 85% 85% 85% 85% 85%Silver Recovery % 40% 40% 40% 40% 40% 40% 40% 40% 40% 40%

TOTAL METAL RECOVERABLE Gold Recoverable ounces 118,780 117,818 179,066 274,605 245,731 265,103 193,463 194,139 194,139 1,782,844 Silver Recoverable ounces 175,061 233,414 350,122 466,829 466,829 466,829 420,976 448,156 448,156 3,476,370

Silver/Gold Ratio 1.5 2.0 2.0 1.7 1.9 1.8 2.2 2.3 2.3 1.9 Cumulative Recoverable Gold ounces 118,780 236,598 415,663 690,268 935,999 1,201,103 1,394,566 1,588,705 1,782,844 Cumulative Recoverable Silver ounces 175,061 408,475 758,597 1,225,426 1,692,255 2,159,083 2,580,059 3,028,215 3,476,370 Initial Year Percent of Recoverable Metal Recovery Factor % 75 75 75 75 75 75 75 75 100 Second Year Percent of Recoverable Metal Recovery Factor % 25 25 25 25 25 25 25 25 25 Gold Recovered ounces 89,085 118,058 163,754 250,720 252,950 260,260 211,373 193,970 242,674 1,782,844 Silver Recovered ounces 131,296 218,826 320,945 437,652 466,829 466,829 432,439 441,361 560,195 3,476,370



Table 22.13 continued

Cash Flow Schedule (Owner Operating with Erato)


RevenueCOMMODITY PRICES Gold $/ounce 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00

Silver $/ounce 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00

GROSS SALES Gold $ 106,901,877 141,669,867 196,504,377 300,864,323 303,539,588 312,312,162 253,647,730 232,764,091 291,208,600 2,139,412,616 Silver $ 2,625,912 4,376,520 6,418,897 8,753,041 9,336,577 9,336,577 8,648,778 8,827,213 11,203,892 69,527,406

REFINING & TRANSPORTATION Gold $ (467,696) (619,806) (859,707) (1,316,281) (1,327,986) (1,366,366) (1,109,709) (1,018,343) (1,274,038) (9,359,930) $/ounce (5.25) (5.25) (5.25) (5.25) (5.25) (5.25) (5.25) (5.25) (5.25) (5.25)

Silver $ (97,256) (162,093) (237,737) (324,187) (345,799) (345,799) (320,325) (326,934) (414,959) (2,575,089) $/ounce (0.74) (0.74) (0.74) (0.74) (0.74) (0.74) (0.74) (0.74) (0.74) (0.74)

PAYABLES Gold % 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 Silver % 90.00 90.00 90.00 90.00 90.00 90.00 90.00 90.00 90.00 90.00

NET REVENUE Gold $ 106,434,181 141,050,061 195,644,671 299,548,042 302,211,603 310,945,796 252,538,021 231,745,748 289,934,563 2,130,052,685 $/ounce 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75

Silver $ 2,275,790 3,792,984 5,563,044 7,585,969 8,091,700 8,091,700 7,495,607 7,650,251 9,710,040 60,257,085 $/ounce 17.33 17.33 17.33 17.33 17.33 17.33 17.33 17.33 17.33 17.33

Total Net Revenue $ 108,709,972 144,843,045 201,207,714 307,134,010 310,303,302 319,037,496 260,033,629 239,395,999 299,644,602 2,190,309,771

Operating CostsMining Cost $ 21,159,550 28,935,863 45,227,177 65,899,474 70,918,741 74,024,449 72,072,433 94,937,034 94,937,034 568,111,756

$/tonne 5.64 5.79 6.03 6.59 7.09 7.40 7.99 9.89 9.89 7.63 Processing $ 17,249,999 21,000,000 26,325,000 32,900,000 32,900,000 32,900,000 29,668,468 31,584,000 31,584,000 256,111,467

$/tonne 4.60 4.20 3.51 3.29 3.29 3.29 3.29 3.29 3.29 3.44 General & Administration $ 1,912,500 1,950,000 1,950,000 2,000,000 2,000,000 2,000,000 1,803,554 1,920,000 1,920,000 17,456,054

$/tonne 0.51 0.39 0.26 0.20 0.20 0.20 0.20 0.20 0.20 0.23 Total Operating Cost $ - 40,322,049 51,885,863 73,502,177 100,799,474 105,818,741 108,924,449 103,544,456 128,441,034 128,441,034 841,679,277

Operating ProfitOperating Profit $ 68,387,923 92,957,183 127,705,537 206,334,536 204,484,561 210,113,047 156,489,173 110,954,965 171,203,568 1,348,630,493

$/gold ounce 767.67 787.38 779.86 822.97 808.40 807.32 740.35 572.02 705.49 756.45

Capital CostsMining Cost $ 9,068,273 54,520,535 2,829,335 29,678,135 48,670,030 3,789,335 3,465,335 1,899,335 2,884,810 2,884,810 (7,992,126) 151,697,807 Buy out Newmont Cost 15,600,000 15,600,000 Process Plant Direct Cost $ 38,231,631 38,231,631 26,183,953 102,647,215 Process Plant Indirect Cost & Contingency $ 28,619,971 28,619,971 17,219,184 74,459,127 Leach Pads $ 19,756,625 12,641,830 7,746,379 40,144,833 Closure and Reclamation $ 5,282,708 1,349,521 935,031 151,877 7,719,137 Total Capital Cost $ 66,851,602 95,676,500 70,120,535 2,829,335 85,723,102 48,670,030 3,789,335 11,211,713 1,899,335 2,884,810 2,884,810 -2,709,418 1,349,521 935,031 151,877 392,268,119

Pre-Income Tax Cash FlowPre-Income Tax Cash Flow $ (66,851,602) (95,676,500) (1,732,612) 90,127,848 41,982,435 157,664,506 200,695,226 198,901,333 154,589,838 108,070,155 168,318,758 2,709,418 (1,349,521) (935,031) (151,877) 956,362,374

$/gold ounce (19.45) 763.42 256.38 628.85 793.42 764.24 731.36 557.15 693.60 536.43 Cumulative Pre-income Tax Cash Flow $ (66,851,602) (162,528,102) (164,260,714) (74,132,866) (32,150,431) 125,514,075 326,209,301 525,110,634 679,700,472 787,770,627 956,089,385 958,798,803 957,449,282 956,514,251 956,362,374

(2) (1) 1 2 3 4 5 6 7 8 9 Payback, operating years 3.2 4.20 4


IRR 40.8%




Table 22.14Owner Operated Mining with Erato

Rate of Return Sensitivity Percent Changes -20% -15% -10% -5% Base 5% 10% 15% 20%


Amulsar Gold Project Pre-Income Tax Sensitivity NPV@5% (Owner Operator with Erato)

0

100,000,000

200,000,000

300,000,000

400,000,000

500,000,000

600,000,000

700,000,000

800,000,000

900,000,000

1,000,000,000

-20% -15% -10% -5% Base 5% 10% 15% 20%


NPV

@ 5

% P

re-T

ax C

alcu

latio

n - U

S$


Figure 22.5 - Amulsar Gold Project Pre-Tax Sensitivity IRR (Owner Operating with Erato)




Table 22.15Owner Operated Mining with Erato

NPV Sensitivity (US$ X 1000) Percent

Changes -20% -15% -10% -5% Base at

5% 5% 10% 15% 20% Gold Price 316,430 390,987 465,544 540,101 614,659 689,216 763,773 838,330 912,887Silver Price 606,031 608,188 610,345 612,502 614,659 616,815 618,972 621,129 623,286Gold Recovery 317,735 391,966 466,197 540,428 614,659 688,890 763,121 837,352 911,583Silver Recovery 606,351 608,428 610,505 612,582 614,659 616,736 618,813 620,889 622,966Gold Grade 317,735 391,966 466,197 540,428 614,659 688,890 763,121 837,352 911,583Silver Grade 606,351 608,428 610,505 612,582 614,659 616,736 618,813 620,889 622,966Operating Cost 707,890 684,582 661,274 637,967 614,659 591,351 568,043 544,735 521,427Capital Cost 680,969 664,391 647,814 631,236 614,659 598,081 581,504 564,926 548,349

Amulsar Gold Project Pre-Income Tax Sensitivity NPV@5% (Owner Operator with Erato)

0

100,000,000

200,000,000

300,000,000

400,000,000

500,000,000

600,000,000

700,000,000

800,000,000

900,000,000

1,000,000,000

-20% -15% -10% -5% Base 5% 10% 15% 20%


NPV

@ 5

% P

re-T

ax C

alcu

latio

n - U

S$


Figure 22.6 - Amulsar Gold Project Pre-Tax Sensitivity NPV@5% (Owner Operating with Erato)

As seen in Tables 22.14 and 22.15, the project's owner operating mining with Erato case pre-income tax rate of return is 40.8 percent and the project's contract mining case pre-income tax net present value at a 5 percent discount rate is US$ 614.7 million. A ten percent increase in the gold price increases the estimated rate of return to 47.1 percent and increases the project's net present value, at a 5 percent discount rate, to US$ 763.8 million.



22.4 Contract Mining with Erato Case Table 22.16 shows the project's contract mining with Erato pre-income tax internal rate of return and the project's pre-income tax net present values at discount rates from 0 to 20 percent.

Table 22.16 Contract Mining with Erato

Economic Analysis Summary Before Tax Internal Rate of Return (IRR), % 46.0 Before Tax Net Present Values US$ x 1000 @ 0 % discount rate 895,236 @ 5 % discount rate 592,623 @ 10 % discount rate 396,225 @ 15 % discount rate 265,582 @ 20 % discount rate 176,790


Table 22.17 Contract Mining with Erato

Economic Analysis Summary - Before Tax Cash Flow and Unit Values

$US x 1000 $US/t





Table 22.18Cash Flow Schedule (Contract Mining with Erato)


Mine ProductionTOTAL OPEN PIT Resources Mined tonnes 4,164,020 5,025,234 7,419,931 10,026,901 10,001,776 9,636,299 8,993,610 7,750,000 7,750,000 70,767,771

Gold g/t 1.164 0.857 0.867 1.004 0.897 0.970 0.785 0.710 0.710 0.877 Silver g/t 3.630 3.630 3.630 3.630 3.630 3.630 3.630 3.630 3.630 3.630

Waste tonnes 5,000,016 10,000,006 17,308,974 27,188,923 27,267,096 26,594,351 22,609,185 24,800,000 24,800,000 185,568,549 Total Mined tonnes 9,164,035 15,025,240 24,728,905 37,215,823 37,268,871 36,230,650 31,602,795 32,550,000 32,550,000 256,336,320

Processing Plants ProductionTOTAL RESOURCE PROCESSED Resources Processed tonnes 3,750,000 5,000,000 7,500,000 10,000,000 10,000,000 10,000,000 9,017,771 7,750,000 7,750,000 70,767,771

Gold g/t 1.159 0.862 0.874 1.005 0.899 0.970 0.785 0.710 0.710 0.877 Silver g/t 3.630 3.630 3.630 3.630 3.630 3.630 3.630 3.630 3.630 3.630

TOTAL RECOVERY Gold Recovery % 85% 85% 85% 85% 85% 85% 85% 85% 85% 85%Silver Recovery % 40% 40% 40% 40% 40% 40% 40% 40% 40% 40%

TOTAL METAL RECOVERABLE Gold Recoverable ounces 118,780 117,818 179,066 274,605 245,731 265,103 193,463 150,373 150,373 1,695,312 Silver Recoverable ounces 175,061 233,414 350,122 466,829 466,829 466,829 420,976 361,792 361,792 3,303,644

Silver/Gold Ratio 1.5 2.0 2.0 1.7 1.9 1.8 2.2 2.4 2.4 1.9 Cumulative Recoverable Gold ounces 118,780 236,598 415,663 690,268 935,999 1,201,103 1,394,566 1,544,939 1,695,312 Cumulative Recoverable Silver ounces 175,061 408,475 758,597 1,225,426 1,692,255 2,159,083 2,580,059 2,941,851 3,303,644 Initial Year Percent of Recoverable Metal Recovery Factor % 75 75 75 75 75 75 75 75 100 Second Year Percent of Recoverable Metal Recovery Factor % 25 25 25 25 25 25 25 25 25 Gold Recovered ounces 89,085 118,058 163,754 250,720 252,950 260,260 211,373 161,146 187,966 1,695,312 Silver Recovered ounces 131,296 218,826 320,945 437,652 466,829 466,829 432,439 376,588 452,240 3,303,644



Table 22.18 continued

Cash Flow Schedule (Contract Mining with Erato)


RevenueCOMMODITY PRICES Gold $/ounce 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00 1,200.00

Silver $/ounce 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00

GROSS SALES Gold $ 106,901,877 141,669,867 196,504,377 300,864,323 303,539,588 312,312,162 253,647,730 193,374,684 225,559,589 2,034,374,198 Silver $ 2,625,912 4,376,520 6,418,897 8,753,041 9,336,577 9,336,577 8,648,778 7,531,763 9,044,809 66,072,873

REFINING & TRANSPORTATION Gold $ (467,696) (619,806) (859,707) (1,316,281) (1,327,986) (1,366,366) (1,109,709) (846,014) (986,823) (8,900,387) $/ounce (5.25) (5.25) (5.25) (5.25) (5.25) (5.25) (5.25) (5.25) (5.25) (5.25)

Silver $ (97,256) (162,093) (237,737) (324,187) (345,799) (345,799) (320,325) (278,954) (334,993) (2,447,143) $/ounce (0.74) (0.74) (0.74) (0.74) (0.74) (0.74) (0.74) (0.74) (0.74) (0.74)

PAYABLES Gold % 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 Silver % 90.00 90.00 90.00 90.00 90.00 90.00 90.00 90.00 90.00 90.00

NET REVENUE Gold $ 106,434,181 141,050,061 195,644,671 299,548,042 302,211,603 310,945,796 252,538,021 192,528,670 224,572,766 2,025,473,811 $/ounce 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75 1,194.75

Silver $ 2,275,790 3,792,984 5,563,044 7,585,969 8,091,700 8,091,700 7,495,607 6,527,528 7,838,834 57,263,156 $/ounce 17.33 17.33 17.33 17.33 17.33 17.33 17.33 17.33 17.33 17.33

Total Net Revenue $ 108,709,972 144,843,045 201,207,714 307,134,010 310,303,302 319,037,496 260,033,629 199,056,198 232,411,600 2,082,736,967

Operating CostsMining Cost $ 26,578,668 37,012,717 58,254,209 85,603,904 92,206,792 96,264,854 93,972,353 83,383,917 83,383,917 656,661,330

$/tonne 7.09 7.40 7.77 8.56 9.22 9.63 10.42 10.76 10.76 9.28 Processing $ 17,249,999 20,950,000 26,325,000 32,800,000 32,800,000 32,800,000 29,578,290 27,202,500 27,202,500 246,908,290

$/tonne 4.60 4.19 3.51 3.28 3.28 3.28 3.28 3.51 3.51 3.49 General & Administration $ 1,912,500 1,950,000 1,950,000 2,000,000 2,000,000 2,000,000 1,803,554 2,015,000 2,015,000 17,646,054

$/tonne 0.51 0.39 0.26 0.20 0.20 0.20 0.20 0.26 0.26 0.25 Total Operating Cost $ - 45,741,168 59,912,717 86,529,209 120,403,904 127,006,792 131,064,854 125,354,197 112,601,417 112,601,417 921,215,674

Operating ProfitOperating Profit $ 62,968,804 84,930,328 114,678,505 186,730,106 183,296,510 187,972,642 134,679,431 86,454,781 119,810,184 1,161,521,293

$/gold ounce 706.84 719.39 700.31 744.77 724.64 722.25 637.16 536.50 637.40 685.14

Capital CostsMining Cost $ 9,085,196 1,202,258 1,202,258 602,258 2,100,554 566,258 566,258 1,916,258 1,973,693 1,973,693 4,526,514 25,715,197 Buy out Newmont Cost 15,600,000 15,600,000 Process Plant Direct Cost $ 38,231,631 38,231,631 26,183,953 102,647,215 Process Plant Indirect Cost & Contingency $ 28,619,971 28,619,971 17,219,184 74,459,127 Leach Pads $ 19,756,625 12,641,830 7,746,379 40,144,833 Closure and Reclamation $ 5,282,708 1,349,521 935,031 151,877 7,719,137 Total Capital Cost $ 66,851,602 95,693,423 16,802,258 1,202,258 56,647,226 2,100,554 566,258 8,312,637 1,916,258 1,973,693 1,973,693 9,809,222 1,349,521 935,031 151,877 266,285,509

Pre-Income Tax Cash FlowPre-Income Tax Cash Flow $ (66,851,602) (95,693,423) 46,166,546 83,728,070 58,031,280 184,629,553 182,730,252 179,660,006 132,763,173 84,481,089 117,836,491 (9,809,222) (1,349,521) (935,031) (151,877) 895,235,784

$/gold ounce 518.23 709.21 354.38 736.40 722.40 690.31 628.10 524.25 626.90 528.07 Cumulative Pre-income Tax Cash Flow $ (66,851,602) (162,545,025) (116,378,479) (32,650,409) 25,380,871 210,010,423 392,740,676 572,400,681 705,163,855 789,644,944 907,481,435 897,672,213 896,322,692 895,387,661 895,235,784

(2) (1) 1 2 3 4 5 6 7 8 9 Payback, operating years 2.6 4 4


IRR 46.0%



Sensitivity Analysis The project's pre-income tax internal rate of return sensitivity relative to incremental changes in metal prices, recoveries, grades and costs are shown in Table 22.19 and Figure 22.7.

Table 22.19Contract mining with Erato Rate of Return Sensitivity

Percent Changes -20% -15% -10% -5% Base 5% 10% 15% 20%


Amulsar Gold Project Pre-IncomeTax Sensitivity IRR (Contract Mining with Erato)

0.0%

10.0%

20.0%

30.0%

40.0%

50.0%

60.0%

70.0%

-20% -15% -10% -5% Base 5% 10% 15% 20%


IRR

Pre-

Tax

Cal

cula

tion


Figure 22.7 - Amulsar Gold Project Pre-Tax Sensitivity IRR (Contract Mining with Erato)




Table 22.20

Contract mining with Erato NPV Sensitivity (US$ X 1000)

Percent Changes -20% -15% -10% -5%

Base at 5% 5% 10% 15% 20%

Gold Price 306,907 378,336 449,765 521,194 592,623 664,052 735,481 806,910 878,339Silver Price 584,366 586,430 588,494 590,559 592,623 594,687 596,751 598,816 600,880Gold Recovery 308,157 379,274 450,390 521,507 592,623 663,739 734,856 805,972 877,089Silver Recovery 584,672 586,660 588,647 590,635 592,623 594,611 596,598 598,586 600,574Gold Grade 273,914 351,986 431,128 511,340 592,623 674,976 758,398 842,891 928,454Silver Grade 584,672 586,660 588,647 590,635 592,623 594,611 596,598 598,586 600,574Operating Cost 700,872 673,810 646,747 619,685 592,623 565,561 538,499 511,436 484,374Capital Cost 638,278 626,864 615,450 604,037 592,623 581,209 569,796 558,382 546,968

Amulsar Gold Project Pre-Income Tax Sensitivity NPV@5% (Contract Mining with Erato)

0

100,000,000

200,000,000

300,000,000

400,000,000

500,000,000

600,000,000

700,000,000

800,000,000

900,000,000

1,000,000,000

-20% -15% -10% -5% Base 5% 10% 15% 20%


NPV

@ 5

% P

re-T

ax C

alcu

latio

n - U

S$


Figure 22.8 - Amulsar Gold Project Pre-Tax Sensitivity NPV@5% (Contract Mining with Erato)

As seen in Tables 22.19 and 22.20, the project's contract mining with Erato case pre-income tax rate of return is 46.0 percent and the project's contract mining case pre-income tax net present value at a 5 percent discount rate is US$ 592.6 million. A ten percent increase in the gold price increases the estimated rate of return to 52.6 percent and increases the project's net present value, at a 5 percent discount rate, to US$ 735.5 million.



23.0 ADJACENT PROPERTIES The following adjacent properties information is taken from the NI 43-101 report dated 19th May 2011 mentioned in the Introduction of this report. Lydian’s discovery of the gold-bearing epithermal system at Amulsar remains the only significant new mineral discovery in Armenia since the country’s independence from the Soviet Union. 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 have been adversely impacted by the metal price slump of late 2008-09. 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 exploration companies active in Armenia include Global Gold, Caldera Resources, Anglo African Minerals and Vallex who are the largest Armenian mining company. There are no competitors exploring for gold in the immediate area around Amulsar, although a small locally owned silver mining licence is located on the northwest of the Amulsar licence area.



24.0 OTHER RELEVANT DATA AND INFORMATION

The information presented herein is considered to be sufficient for a preliminary economic analysis. It is anticipated that additional details regarding mining methods, metallurgical testwork and flowsheet development, permitting, heap leaching, and infrastructure items will be expanded upon in an upcoming BFS.



25.0 INTERPRETATION AND CONCLUSION

As a result of considerable effort by Lydian employees and geological, mining, geotechnical, mineral processing, and environmental consultants sufficient technical information is available to produce this Preliminary Economic Assessment. This engineering study qualifies as a PEA level document according to Canadian National Instrument 43-101. The economic models utilized in this report indicated a range of before tax rate of return from 39.5 to 46.0 percent at an assumed gold price of US$ 1,200 per ounce. If the gold price used in the economic evaluation, which excludes inferred resources, is increased to current levels (US$ 1,500 per ounce in July 2011) then the before tax rate of return ranges from 54.7 to 61.8 percent. 25.1 Mining

Based on the optimisation work that has been completed, using the input parameters provided by Lydian, the Amulsar gold deposit has the potential to be a financially viable operation.

The conversion factors from resources to PMM are favourable for both the Indicated resources only option and the Indicated and Inferred resources option. As the optimisation pit shells reach both the lateral and lower extents of the current resource, this indicates that exploration should be extended both laterally and at depth.

The initial optimisation work indicates that a portion of the resources which are currently classified as Inferred have the potential to be economically mineable. Further drilling should therefore be carried out on this portion of the Inferred resources to bring them to the Indicated Category.

The chosen Amulsar Whittle optimisation shell for the Indicated resource only option contains 31.0 million tonnes of PMM (allowing for 5% losses and 5% dilution) at an average head grade of 0.98 g/t and 75.9 million tonnes of waste, resulting in an overall strip ratio of 2.45:1. The selected pit shell provides a “specified discounted cashflow” of US$263.75 million with two push backs. These figures exclude capital expenditure.

The chosen Amulsar Whittle optimisation shell for the Indicated and Inferred resource option contains 53.6 million tonnes of PMM (allowing for 5% losses and 5% dilution) at an average head grade of 0.94 g/t and 138.0 million tonnes of waste resulting in an overall strip ratio of 2.57:1. The selected optimum pit shell provides a “specified discounted cashflow” of US$367.08 million with two push backs. These figures exclude capital expenditure.

The level of confidence in the Scoping Study completed to date would not support the presentation of a Reserves Statement for the Amulsar Project.

After applying reasonable factors for losses, mining recovery and dilution, the following estimates of potentially mineable mineralization (“PMM”) can be proposed:



Conventional Open Pit Mining (Tigranes area and Artavasdes area)

PMM based on Indicated resources 36.5 Mt @ 0.92 g/t Au

PMM based on Inferred Resources 18.8 Mt @ 0.93 g/t Au

In addition to the figures quoted above, there is the potential for extensions to the Amulsar PMM to the north in the Erato area. Sensitivity analysis undertaken during the Optimisation studies has indicated this additional mineralisation would have a positive effect on the project financials.

The studies to date have not indicated any major difficulties or costs in establishing the necessary infrastructure facilities to support the operation.

Enquiries in Armenia have indicated that there are currently no contract mining companies in Armenia with the capacity or equipment to satisfy the Amulsar production schedule. Enquiries have also shown that the purchase of haul trucks greater than 180 tonne capacity in Armenia will be challenging. As a result, this study has only utilised haul trucks smaller than 180 tonne capacity.

An optimisation utilising the final PEA financial and physical parameters reported results in line with the earlier optimisations completed in the former stages of the study.

25.2 Metallurgical Testing and Mineral Processing

Results of past test programs revealed that heap leach gold extraction was influenced by particle size distribution. Test data and heap leach projections indicated that acceptable gold extraction rates and levels required crushing the gold ore to a size distribution of 80 percent passing 12 millimeters. Based on this information, an equipment assemblage was developed to crush and leach an average of 30,000 tonnes per day of ore at the size distribution mentioned above.

The Amulsar ore body is low grade. Preliminary studies completed to date indicate that it is unlikely that gold could be economically extracted if fine grinding was required as part of a beneficiation scheme. The differential gold extraction between heap leaching and fine grinding plus agitated leaching would likely not offset the capital and operating cost increases imposed by grinding. However, additional work is required to eliminate agitation leaching as a beneficiation technique.

Additional work will be devoted to infrastructure components including power acquisition, road upgrading, raw water procurement, and the leach pad and pond location. The major project risk is due to the fact that there are no existing heap leach gold mines currently operating in the country. Therefore there are also limited manpower resources available to operate the facilities. On a positive note, the Amulsar gold ore does not contain any deleterious elements and with appropriate training it is anticipated that operations will reach acceptable efficiencies.



25.3 Environmental and Social

A number of environmental and social issues were identified through the scoping

exercise and public consultations. The social and environmental issues identified were not prohibitive but will form a focal point for further baseline studies in order to better characterise them and to provide information on how they might be mitigated. Risks will also be directly addressed via assessments, consultation and the preparation and implementation of relevant management plans.

The findings of the Scoping Study and the terms of reference for continuing work

were presented, in May 2011, to the neighbouring communities as part of the public consultation and disclosure process. Key issues and observations were noted and appropriate actions have been incorporated in the proposed ESIA scope of programme of works. Lydian has commissioned several discrete studies in response to the findings of the Scoping Study and public consultation undertaken and has pledged support and resources to the actions identified.



26.0 RECOMMENDATIONS

The following opportunities were proposed by the client and their consultants to further reduce the overall capital or operating costs:

Mining

Formulate a Project Development Schedule, in line with corporate objectives,

that defines a Staged Study approach with associated data collection phases. The scope of work for each study phase should be clearly defined.

It is recommended the key technical data collection for the next phase would

include:

▪ Additional drilling to better define the structure and anomalies in the Amulsar deposit and expand and reclassify the resource base

▪ Geotechnical drilling, test work and analysis

▪ Hydrogeology test work and modelling

▪ Drilling to collect a representative sample for metallurgical testwork.

Mineral Processing

Additional metallurgical testing will be required to fully define the different

possible leaching techniques to economically process the Amulsar ores. An extensive column leach test program should be conducted at METCON Research or similar laboratory.

Crushing

Further studies to improve the economics include the following:

▪ Contract Mining verses Owner operated mining ▪ Further review of phased or ramp up of the production rates ▪ In-pit crushing ▪ Single primary gyratory crusher verses two jaw crushers ▪ Further review the topography to minimize earthwork, foundation and

conveying costs

Leach Pad and Metal Recovery

▪ Finalize leach pad location and construction schedule ▪ Identify required process water resources ▪ Finalize process plant location



Infrastructure

Review access road costs, with equipment that will be onsite there is a good chance that the road construction could be reduced by combining project with overall construction including the leach pad and ponds. Power

Further review to define the closest source from the local Armenian grid power.

Review the best options to get power to site and begin negotiations with the Armenian power utility company to further economize the design options. Environmental and Social

A work plan covering all of the issues identified has been formally developed

between Lydian/Geoteam and WAI and work is progressing against each of these topics. WAI has recommended that the scope of ESIA baseline and assessments to quantify and/or mitigate against environmental and social issues identified include:

▪ Groundwater and surface water hydrology, quality and sustainability -

further hydro-geological study, modelling and geochemical assessment via the installation of monitoring wells needs to be undertaken. Site-wide water balance calculations need to be determined and the sustainability of sourcing examined and managed;

▪ Visual impact – The visual impact of the project needs to be assessed from

Jermuk, the public highway and village communities, with the results included in public consultations;


required; ▪ Land disturbance and take - a land use survey, biodiversity assessment

and archaeological study are required to inform an impact assessment and assessment of any compensation required for grazing lands;

▪ Biodiversity - field surveys need to be undertaken on areas of major

disturbance. An assessment of flora harvested and collected by villagers is required. These studies will inform an overall impact assessment;

▪ Dust - dust dispersion modelling and impact assessment on edible plant

species used by the communities needs to be undertaken; ▪ Noise - Noise impact modelling, incorporating the operating specifics of mine

fleet and machinery needs to be undertaken to inform the noise impact assessment on nearest sensitive receptors;



▪ Cyanide and hazardous materials - As the project will use cyanide in its heap leach process, the usage, transport, handling and disposal of cyanide will need to be managed via formal plans and procedures in accordance with international best practice, particularly the International Cyanide Management Code (ICMC);



▪ Local labour - high unemployment and expectations require careful

management and a Skills Audit to assess future potential employment opportunities should be undertaken;


assessment (to include seasonal visitors) need to be undertaken; ▪ Local Industry - further appraisal of the main (agricultural and

manufacturing) sectors is required, including individual contributions to domestic income. An assessment of potential diversification and increase in access to markets with advent of mining is considered beneficial;

▪ Public disclosure and consultation - to be undertaken in line with

international best practice, at all stages of the ESIA and mine development processes (to include Jermuk) to manage expectations, allay concerns, establish NGO partnerships, inform community development and spending and influence mine design;

▪ Closure - considerations need to be addressed from conceptual stage via the

preparation and continual update of a Mine Closure and Rehabilitation Plan (to include heap leach rehabilitation).

WAI propose to monitor and review the process of environmental and social

baseline data collection by Lydian/Geoteam, and general progression of the ESIA works, and provide continuing technical support, as required, to ensure that both national and international requirements are met. The baseline programme is continuing and the next ESIA deliverable will be a baseline status report in Q3 2011.

WAI will also work closely with the BFS team to ensure that environmental and

social issues are taken into account with the design and siting of all major infrastructure and that appropriate alternatives are considered and/or mitigation and management measures are employed to ensure that impacts are addressed. It is therefore recommended that a time-bound, integrated ESIA and BFS schedule of works is prepared to facilitate this process.



WAI will need to ensure that the public consultation, which will be undertaken for the purposes of the ESIA, will include available information from the BFS so that project stakeholders have the opportunity to input into the process.

Project Planning

▪ Commence work on a project operating plant and project execution plan ▪ Collect local and international unit construction/fabrication and supplies/labor

costs. ▪ Investigate mine equipment leasing and project financing plan

Next Stage Budget and Schedule

The authors of this report are of the opinion that the character of the Lydian’s

Amulsar Gold Project is of sufficient merit to commence with BFS beginning in August 2011.

It is recommended that the project be advanced to the BFS stage. Costs for this

level of study are estimated as follows:

US$Mine Design 220,000Mineral Processing & Infrastructure 366,254Leach Pad, Ponds, Pit Slopes & Hydrology 570,186Hydro-geological 189,099Seismic 77,285

Total 1,422,824

It is estimated that a final Lydian BFS will require 6 to 8 months to complete.



27.0 REFERENCES CSA Global Pty Ltd, "Mineral Resource Estimate Lydian International Limited Amulsar Gold Project - 43-101 Technical Report", May 19, 2011 CSA Global Pty Ltd, "Scoping Study Report Lydian International Ltd Amulsar Gold Project Armenia", August 4, 2011 Golder Associates Inc., “Conceptual Design and Scoping-Level Cost Estimates – Heap Leach Facility, Amulsar Gold Project, Armenia”, July 19, 2011 SGS Lakefield Research Limited, “An Investigation into Gold Recovery from the Amulsar Project Samples, Project 11956-001 – Final Report”, September 30, 2008 SGS Minerals Services UK Ltd, “Report on Cyanide Leach Testing of Three Different Gold Ore Composite Samples from the Amulsar Deposit, Armenia, Project 10866-124”, March 26, 2010 Wardell Armstrong LLP, “Metallurgical Testing of Samples from the Amulsar Deposit”, January 13, 2011



28.0 APPENDICES



Appendix 1

Design Criteria

REV BY DATE KDE DATE DESCRIPTION PAGES

NO APPR

P1 JJ 6/7/11 BCS 6/7/11 Preliminary 17

DATE:

LYDIAN INTERNATIONAL LTD. APPROVAL

SIGNATURE:

DOCUMENT NO. KDE Q 439-02-010

DESIGN CRITERIA

Amulsar - Phase II Preliminary Economic Assessment

LYDIAN INTERNATIONAL LTD.

K D Engineering

KDE FORM No. A131a-7/12/99

Client: LYDIAN INTERNATIONAL LTD.

Project: Amulsar - Phase II Preliminary Economic Assessment

Project No.: 439-02 Date: Jun-07-2011

Nominal Design Nominal Design Source Code

1 SCOPE

C

2 RELEVANT CODES AND STANDARDS

I

Code Source

A Assumptions For Study

B Calculated

C Client Information

I Industry Standard Practice

K KDE

O Information Provided by Others

P Published Information / Criteria

T Engineering, Testwork or Reports

V Vendor Data

3 SITE DATA

Location

C

Site Elevation, Meters Above Sea Level 2,988 C

Weather Data

Ambient Air Temperature

Maximum: °C 34.2 C

Minimum: °C -27.6 C

Precipitation Data

Precipitation, mm ( 24 Hr / 100 year event) 95 O

Precipitation, mm (Annual Average) 663 C

Max 211 C

Maximum Snow Coverage, mm 63 C

Snow load, kg per square meter ?? A

Evaporation, mm (Annual Average) 881 C

Wind Direction, Prevailing TBD C

Design Wind Load, kilometers per hour 150 A

Seismic Zone O

DESIGN CRITERIA


Rev: P1

The following criteria have been developed to support the Preliminary Economic Assessment (PEA) study for LYDIAN INTERNATIONAL Ltd. at the Amulsar property in

Armenia. Precious metals will be recovered from low grade material by heap leaching. To prepare the ore for leaching, a crushing plant is planned. In Phase I, the ore

will be processed through the crushing plant at 5 million tonnes per year; Phase II will expand the facility to process 10 million tonnes per year. Crushed ore will be

deposited on the heap using a conveyor - stacker system and will be leached with a dilute cyanide solution in the conventional manner. Precious metals will be

recovered from the leach solution by adsorption on activated carbon. The precious metals will be periodically desorbed from the activated carbon and the stripped

carbon will be reactivated. Precious metals will be recovered from the strip solution by electrowinning. The product from electrowinning will be dried and smelted to

produce dore bullion which will be shipped from the site to refiners.

Phase II (10 Million Tonnes per year)Phase I (5 Million Tonnes per year)

European codes, standards and regulations will be used. Applicable codes, standards and regulations will be referenced in each Technical Specification that is

applicable for the particular piece of equipment or system that is being designed. Specific design standards will be referenced, as required, in each Technical

Specification.

Amulsar is located 170 km from Yerevan, Armenia and 6 km from Gorhayk village.

Site is located in a relatively high seismic risk area

PGArock between 0.4g and 0.5g for a seismic event

with a 10% probability of exceedance in 50 years (475-

yr event)

KDE FORM NO.: E132-7/12/99 Page 1 of 17 KD Engineering





DESIGN CRITERIA


Rev: P1


4 PROCESS DESCRIPTION

K

5 REFERENCE DOCUMENTS

5.1 MINE SCHEDULE

Not Available

5.2 FLOWSHEETS

10-F-01 Flowsheet, Primary Crushing K

13-F-01 Flowsheet, Secondary Crushing K

15-F-01 Flowsheet, Tertiary Crushing K

17-F-01 Flowsheet, Lime Addition K

19-F-01 Flowsheet, Ore Stacking K

20-F-01 Flowsheet, Heap Leach K

23-F-01 Flowsheet, Carbon Adsorption K

25-F-01 Flowsheet, Stripping and Refining K

27-F-01 Flowsheet, Carbon Reactivation K

30-F-01 Flowsheet, Utilities and Reagents K

5.3 METALLURGICAL REPORTS

C

6 PROCESS DESIGN CRITERIA

6.1 GENERAL

6.1.1 Fuel Source

6.1.2 Electric Power

Operating Voltage O

Operating Frequency O

Operating Phase O

Control Voltage O

Control Frequency O

Control Phase O

6.1.3 ROM Total, Tonnes C

Daily Ore Feed Rate, Tonnes per Day (Nominal) 14,085 28,169 B

Daily Ore Feed Rate, Tonnes per Day (Design) 16,197 32,394 B

6.1.4 Mining Days per Year 355 355 355 355 A

6.1.5 Annual Ore Feed Rate, Tonnes (Nominal) 5,000,000 10,000,000 C

6.1.6 Annual Ore Feed Rate, Tonnes (Design) 5,750,000 11,500,000 B

6.1.7 Nominal Mine Life, years 24.0 12.0 B

6.1.8 Average Ore Grade, grams per tonne

Gold, Resource Total C

Silver Resource Total C

6.1.9 Average ROM Gold Recovery, Percent T

6.1.10 Average ROM Silver Recovery, Percent A

6.1.11 Average Annual Production Rate, Ounces

Gold 139,856 160,834 279,711 321,668 B

220

50 Hz

3

Diesel, Gasoline

6000/400

50 Hz

1.00

2.00

87

43

120,000,000

Increasing throughput from 5 to 10 million tonnes per year will entail installation of duplicate equipment for a majority of the processing unit operations.

Run-of-mine ore will be hauled by truck to the crushing plant area by rear-dump trucks and dumped into a crusher feed hopper. An open stockpile is provided adjacent

to the crusher so trucks can dump if the crusher is not available. A single crushing line will process the design tonnage at the indicated average daily throughput noted

below. The line consists of a dump hopper, apron feeder, stationary grizzly, rock breaker, jaw crusher and associated transfer equipment. The primary crusher reduces

the size of run-of-mine ore from a maximum of 700 mm to approximately 80% passing 175 mm. Crushed ore drops onto a belt conveyor and transported to the coarse

ore storage bin. Apron feeders and a transport conveyor feed the coarse to to a double deck scalping grizzly to remove undersize material. Vibrating screen oversize is

fed to the secondary cone crusher. Secondary and tertiary crushed product are conveyed to the fine ore screen feed bin. Fine screen oversize at plus 14 millimeter is

fed to the tertiary crushers. Fine screen undersize is the crushing circuit product and is transported by agglomeration by belt conveyor. Lime, required for pH control, and Portland cement, required for agglomeration, will be added to the transfer convey

The crushed ore will be stacked on the pad using conveyors and a slewing and luffing stacker. A bulldozer will be used to rip and crossrip the heap surface before

placement of a new lift of ore.

Pregnant solution from the leaching process will be pumped from the pregnant solution pond to a carbon Adsorption-Desorption-Recovery (ADR) plant to recover the

gold from solution. Adsorption will take place in one train of 5 carbon adsorption columns in series. Solution will flow from each adsorption stage to the next stage by

gravity. Carbon will be advanced countercurrent to the solution flow by a carbon advance pump. Barren solution discharging from adsorption will flow to a barren pond

from which it will be pumped to the ore heap for further leaching. High strength cyanide solution will be injected into the barren solution to maintain the cyanide

concentration at the desired level.

Precious metals will be recovered from the strip solution by electrowinning. When the electrowinning cell is appropriately loaded with gold and silver, the stainless steel

wool cathodes will be washed with high pressure water inside the cells to flush the precious metal sludge from the cathodes. The precious metals sludge slurry then flow

to a holding tank and to a pressure filter. The stripped cathodes will be returned to the electrowinning cells. The filter cake will be treated in a retort to dry the cake. The

dried product from the retort will be smelted in an induction furnace and poured into 500 oz to 1000 oz Dore bars. The Dore bars will be shipped to commercial facilities

for refining.

The stripped carbon will be transferred by pressure eduction from the elution column over a screen and into the reactivation furnace feed tank. Carbon reactivation will

be accomplished in an electric rotary kiln-type furnace.

3

An installation for mixing and adding reagents to the process and elution circuits and a system for adding new carbon are included in the process design.

Loaded carbon from the adsorption circuit will be transferred to the acid wash vessel. The acid will be removed from the carbon by washing with fresh water. After acid

washing, the carbon will be pumped to the elution vessel. The pressurized Zadra process will be used for gold elution. Hot strip solution will be circulated through the

carbon elution vessel to desorb the precious metals from the carbon.

SGS - Report on cyanide leach testing of one drill hole sample from the Amulsar deposit, Armenia dated September, 2008

SGS - Report on cyanide leach testing of three different gold ore composite samples from the Amulsar deposit, Armenia dated March, 2010

Wardell-Armstrong - Report on cyanide leach testing of two of the SGS composite samples, dated January, 2011






DESIGN CRITERIA


Rev: P1


Silver 138,248 158,985 276,496 317,971 B

6.2 ORE CHARACTERISTICS

6.2.1 Specific Gravity T

Average T

6.2.2 Run of Mine Ore Size, mm (Estimated)

Percent Passing

700 mm A

430 mm A

47 mm A

6.2.3 Run of Mine Ore Moisture

Normal, Percent A

6.2.4 Heap Retained Moisture, Percent (Dry Basis)

Crushed Oxide A

6.2.5 Heap Saturation Moisture, Percent (Dry Basis)

Crushed Oxide A

6.2.6 Draw Down Angle, Degrees A

6.2.7 Bulk Density Tonnes per Cubic Meter

For Volume Calculations A

For Structural Calculations A

6.2.8 Bond Impact Crushing Work Index, kWh/Tonne A

6.2.9 Bond Abrasion Index, g A

6.3 OPERATING TIMES

6.3.1 Crushing Schedule

Days per Year C

Shifts per Day C

Hours per Shift C

Availability A

6.3.2 Ore Stacking Schedule

Days per Year C

Shifts per Day C

Hours per Shift C

6.3.3 Leaching and Adsorption

Days per Year C

Shifts per Day C

Hours per Shift C

Availability I

6.3.4 Carbon Striping and Electrowinning

Days per Week C

Hours per Day C

Availability I

6.3.5 Carbon Reactivation

Days per Week C

Hours per Day C

Availability I

6.3.6 Refining

Days per Week (1 Day Melting) C

Hours per Day C

6.4 PRIMARY CRUSHING

6.4.1 Number of Primary Crushers Operating (Total) 1 1 2 2 K

Type Jaw Jaw Jaw Jaw K

Close Side Setting, mm 130 155 130 155 K

Pocket Discharge Method Conveyor Conveyor Conveyor Conveyor I

Crusher Design, Dry Tonnes per Hour (EACH) 782 900 782 900 B

6.4.2 Rock breaker Yes Yes Yes Yes K

6.4.3 Truck Size, tonnes 90 90 90 90 C

6.4.4 Dump Pocket Size, tonnes 180 180 180 180 A

Feed Method C

C

Number of Truck Dump Sites per Crusher 2 2 2 2 C

Dumps per hour required, number 9 10 9 10 B

Crusher Dump Cycle, minutes 13.8 12.0 13.8 12.0 B

Station Dump Cycle, minutes 13.8 12.0 13.8 12.0 B

6.4.5 Crusher Operating Time, hrs. / Day at Design Throughput B

98%

8

98%

100

80

Direct Dump from Mine Truck or by Loader

18.0

The primary crushing line consists of dump hopper, apron feeder, grizzly screen, rock breaker, crusher and associated dust collection and transfer equipment. Run of Mine (ROM) ore is

dumped into the dump hopper. The apron feeder screen transfers the ore at a controlled rate to the grizzly screen. The screen oversize feeds the jaw crusher. A rock breaker is available to

service the crusher or screen. The crusher reduces the size of run-of-mine ore from maximum 700 mm to approximately 80% passing 175 mm. Crushed ore drops onto a belt conveyor that

transports the crushed ore to a crushed ore storage bin. A bin is used instead of a coarse ore pile due to the extreme wind and potential dust loss.

Dust is controlled in the dump pocket with water sprays and bin vents service at the contained transfer points.

98%

20

2.41

7

5

8

Run-of-mine ore is transported to the crushing plant area by rear-dump trucks and dumped into a crusher feed hopper. An open stockpile is provided adjacent to the crusher so trucks can

dump if the crusher is not available. One crushing line will be required to process 5 million tonnes per year, while two parallel crushing lines are required to process 10 million tonnes per year.

8

24

24

8

11.7

3

1.90

45

1.50

7

18.0

0.15

355

3

8

75%

355

3

365

3






DESIGN CRITERIA


Rev: P1


6.4.6 Primary Crusher Feed

100 Percent Passing Size, mm A

50 Percent Passing Size, mm A

Throughput Rate, Tonnes / Hr (Design) 782 900 782 900 B

6.4.7 Apron Feeder

Number of Feeders, Total 1 1 2 2 V

Number of Feeders, Per Crusher 1 1 1 1 V

Discharge Rate, Tonnes per hour 782 900 782 900 V

6.4.8 Grizzly Screen

Number of Screens, Total 1 1 2 2 V

Number of Screens, Per Crusher 1 1 1 1 V

Discharge Rate, Tonnes per hour 782 900 782 900 V

Grizzly Screen Openings, mm

Slot Size 100 150 100 150 V

Slot Length / Opening Ratio 4 to 1 4 to 1 4 to 1 4 to 1 A

6.4.9 Primary Crusher Discharge

80 Percent Passing Size, mm 135 165 135 165 A

100 Percent Passing Size, mm 285 330 285 330 A

Throughput Rate, Tonnes / Hr 782 900 782 900 A

6.5 SECONDARY CRUSHING

6.5.1 Secondary Feed Bin

Number of Draw Points 2 2 4 4 K

Total Secondary Feed Rate, Tonnes / Hr 782 900 1,565 1,800 V

Live Capacity, design

Hours 8.0 8.0 8.0 8.0 K

Tonnes 6,260 7,199 12,520 14,397 B

Cubic meters 4,173 4,799 8,346 9,598 B

Total Capacity, design

Volume Utilization Factor, Percent 75 75 75 75 A

Tonnes 8,346 9,598 16,693 19,197 B

Cubic meters 5,564 6,399 11,128 12,798 B

Bin Dimensions (Conceptual)

Height, Meters 24.1 25.3 24.1 25.3 B

Length, Meters 21.5 22.5 42.9 45.0 B

Width, Meters 10.7 11.2 10.7 11.2 B

Material Size, mm

P80 135 165 135 165 B

P100 (slabby rock) 285 330 285 330 B

6.5.2 Secondary Feeders

Type Apron Apron Apron Apron A

Number, total 2 2 4 4 K

Number per Secondary Line 2 2 2 2 K

Feeder Capacity, tonnes per hour each 900 1035 900 1035 B

6.5.3 Secondary Feed Screen

Type Vibrating Vibrating Vibrating Vibrating I



Capacity, tonnes per hour each 900 1035 900 1035 B

Deck opening sizes, mm

Top 100 x 100 75 x 75 100 x 100 75 x 75 K

Bottom 28 x 28 22.5 x 22.5 28 x 28 22.5 x 22.5 K

6.5.4 Secondary Crusher:



Type MP 800 Standard Cone MP 800 Standard Cone K

Bowl Type K

Feed Method I

Feed Rate, tonnes per hour 900 1,035 900 1,035 B

6.6 FINE ORE SCREENING

6.6.1 Fine Ore Screen Feed Bin

New Feed, Tonnes per Hour 782 900 1,565 1,800 V

Crusher Circulating Load % 67.2 83.6 67.2 83.6 B

Circulating Load, Tonnes per Hour 526 752 1,052 1,504 V

Total Bin Feed, Tonnes per Hour 1,308 1,652 2,617 3,304 V



Hours 0.5 0.5 0.5 0.5 A

Tonnes 654 826 1,308 1,652 B

Cubic meters 436 551 872 1,101 B



Tonnes 1,308 1,652 2,617 3,304 B

Cubic meters 872 1,101 1,745 2,202 B


Height, Meters 13.0 14.1 13.0 14.1 B

Length, Meters 11.6 12.5 23.1 25.0 A

Width, Meters 5.8 6.3 5.8 6.3 A

210

700

Two parallel secondary crushing trains are provided for 2 x 5 million tonnes per year. Each train will pull ore from the same coarse ore bin, so that some ore can be crushed through all

secondary and tertiary crushers even if one of the jaw crushing lines is down. The bin discharge is fed to a double deck scalping screen to remove fines from the secondary crusher feed.

Crushed product and the screen undersize are conveyed to the fine ore screen feed bin.

Screen Oversize

Secondary screen undersize and secondary and tertiary crusher discharge are fed to the tertiary screen feed bin by a tripper conveyor. The bin has discharge conveyors that feed the tertiary

screens. Screen oversize is conveyed by a tripper conveyor to the tertiary crusher feed bins. Screen undersize is conveyed by a tripper conveyor to the fine ore load out bin.






DESIGN CRITERIA


Rev: P1


Material Size, mm

P80 33 25 33 25 B

P100 (slabby rock) 80 52 80 52 B

6.6.2 Fine Ore Screen Feeders

Type Belt Belt Belt Belt A


Number per Screen Line 1 1 1 1 K

Feeder Capacity, tonnes per hour each 654 826 654 826 B

6.6.3 Fine Ore Screens

Type Vibrating Vibrating Vibrating Vibrating I

Number, total 2 2 4 4 V

Number per Screen Line 1 1 1 1 V

Capacity, tonnes per hour each 654 826 654 826 V

Deck opening sizes, mm

Top 32 x 32 32 x 32 V

Bottom 16 x 16 16 x 16 V

6.7 TERTIARY CRUSHING

6.7.1 Tertiary Crusher Feed Bin

Total Bin Feed, Tonnes per Hour 752 1,504 V

Number of Draw Points 2 4 K


Hours 0.5 0.5 A

Tonnes 376 752 B

Cubic meters 251 501 B


Volume Utilization Factor, Percent 50 50 A

Tonnes 752 1,504 B

Cubic meters 501 1,003 B


Height, Meters 10.8 10.8 B

Length, Meters 9.6 19.2 A

Width, Meters 4.8 4.8 A

Material Size, mm

P80 32 32 B

P100 (slabby rock) 52 52 B

Reclaim Method Belt Conveyor Belt Conveyor Belt Conveyor I

Number of Reclaim Conveyors Two Total One for each

Crusher

Two Total One for each

Crusher

Four Total One for each

Crusher

I

6.7.2 Tertiary Feeders

Type Belt Belt A

Number, total 2 4 K

Number per Crusher Line 1 1 K

Feeder Capacity, tonnes per hour each 432 432 B

6.7.3 Tertiary Crushers

Number 2 4 V

Type HP 800 Short Head Cone HP 800 Short Head Cone V

Bowl Type Medium Medium V

Feed Method Belt from Bin Belt from Bin V

Feed Rate, tonnes per hour per crusher 376 376 B

6.8 CRUSHED ORE SURGE BIN

6.8.1 Crushing Plant Final Product

100 Percent Passing Size, mm 17 17 17 17 V

80 Percent Passing Size, mm 12 12 12 12 V

Total Bin Feed, Tonnes per Hour 782 900 1,565 1,800 V

6.8.2 Crushed Product Bin



Hours 2 2 2 2 A

Tonnes 1,565 1,800 3,130 3,599 B

Cubic meters 1,043 1,200 2,087 2,400 B



Tonnes 3,130 3,599 6,260 7,199 B

Cubic meters 2,087 2,400 4,173 4,799 B


Height, Meters 17.4 18.2 17.4 18.2 B

Length, Meters 15.5 16.2 31.0 32.4 A

Width, Meters 7.7 8.1 7.7 8.1 A

6.8.3 Ore Reclaim Feed Method Belt Feeders Belt Feeders Belt Feeders Belt Feeders A

6.8.4 Crushed Ore Reclaim

Number of Reclaim Points 2 2 4 4 K

Reclaim Rate per Feeder, tonnes per hour 1,500 2,500 1,500 2,500 A

Truck Size, Tonnes 90 90 90 90 C

Truck Fill Time, minutes 3.6 2.2 3.6 2.2 B

Trucks per Operating Hour 8.7 10.0 17.4 20.0 B

Bin Dump Cycle, minutes 6.9 6.0 3.5 3.0 B

Station Dump Cycle, minutes 13.8 12.0 13.8 12.0 B

6.9 ORE STACKING

6.9.1 Stacking Rate, tonnes per day 14,085 16,197 28,169 32,394 B

Ore is fed from the tertiary feed bin to the tertiary crushers. The feed bin is required to allow the crushers to operate in a "choke fed" condition. A conveyor feeds material into each crusher as

needed to control the level in the crusher. Crusher discharge is recycled to the fine ore screen feed bin.






DESIGN CRITERIA


Rev: P1


6.9.2 Bulk Density of ROM Ore, tonnes per cubic meter 1.5 tonne per m3 (dry

basis)

1.5 tonne per m3 (dry

basis)


basis)


basis)

I

6.9.3 Stacking Method Truck Dump from Prepared

Access Corridors; Dozer

Spreading of Ore; Dozer

Ripping before Leaching

Truck Dump from Prepared












K

6.9.4 Truck Size 90 Tonnes 90 Tonnes 90 Tonnes 90 Tonnes C

6.9.5 Lime Addition Method Lime Addition Silo with

Rotary Valve to Trucks

Lime Addition Silo with






C

6.9.6 Lime Addition Rate, kg per tonne 0.5 0.5 0.5 0.5 T

6.9.7 Dosing System Sized to Dose One Truck in Sized to Dose One Truck in Sized to Dose One Truck in Sized to Dose One Truck in A

Less Than 30 Seconds Less Than 30 Seconds Less Than 30 Seconds Less Than 30 Seconds

6.9.8 Lime

Consumption, Kg per Tonne Ore

Oxide 0.5 0.5 0.5 0.5 T

Daily Consumption, Tonnes 100% basis 7.0 8.1 14.1 16.2 B

Delivered Concentration, Pct CaO 90 90 90 90 A

Size Pebble (-25 mm / + 6 mm) Pebble (-25 mm / + 6 mm) Pebble (-25 mm / + 6 mm) Pebble (-25 mm / + 6 mm) A

Shipping Container Bulk Truck Bulk Truck Bulk Truck Bulk Truck C

Lime Addition to Crushed Ore, Tonnes per Hour

Oxide 0.5 0.5 0.5 0.5 T

Storage Capacity, Days 7.0 7.0 7.0 7.0 A

Storage Capacity, Tonnes 55 63 110 126 B

Delivery

Truck Capacity, tonnes 20 20 20 20 A

Average Trucks per Day 0.4 0.4 0.8 0.9 B

Mine Truck Size, Tonnes of Ore 90 90 90 90 C

Lime Addition KG per Truck 50 50 50 50 B

Trucks per Day 156 180 313 360 B

Truck Load out Hours per Day 18 18 18 18 A

Trucks per Hour 9 10 17 20 B

Lime Addition Rate, kg / hr 391 450 782 900 B

Lime Bulk Density, kg / cubic meter 800 800 800 800 A

Lime Addition Rate, m3 / hr 0.49 0.56 0.98 1.12 B

Lime Addition Rate, ft3 / hr 17.3 19.8 34.5 39.7 B

6.10 CRUSHED ORE AGGLOMERATION


Tonnes per hour 782 900 1,565 1,800 B

Product Size, P80, microns 12,000 12,000 B

Percent Solids 96.5 96.5 96.5 96.5 B

6.10.2 Cement Addition, kg per tonne 5 5 5 5 T

Cement Addition, Tonnes per hour 3.9 4.5 7.8 9.0 B

Agglomeration Moisture, Percent Solids 95.0 93.0 95.0 93.0 A

Agglomeration Solution, Tonnes per hour 13.0 35.4 26.0 70.9 B

6.11 CRUSHED ORE CONVEYING


80 Percent Passing Size, micron 12,000 12,000 B

Leach Pad Feed, Tonnes per Hour 786 904 1,573 1,809 B

Crushed ore at minus 12 mm is fed to a drum agglomerator where it is mixed with cement and leach solution. Cement addition is varied with the ore tonnage and ore placement in the heap

based on input from the operations staff. Cement is trucked to the site and unloaded in an on site storage silo. The cement is metered from the silo to the agglomeration feed conveyor by a

rotary valve.






DESIGN CRITERIA


Rev: P1


6.12 ORE STACKING

6.12.1 Stacking Rate, tonnes per day 14,085 16,197 28,169 32,394 B

6.12.2 Bulk Density of ROM Ore, tonnes per cubic meter 1.5 A

6.12.3 Stacking Method Leach Pad Grasshopper

Conveyors and leach Pad

Radial Stacking Conveyor;

Dozer Ripping before

Leaching

K

6.13 LEACH PADS

6.13.1 TypeOne-time leaching of a

single ore heap liftO

6.13.2 Construction Phasing

Phase 1

Ore Storage, tonnes O

Lined Area TBD O

Phase 2

Ore Storage, tonnes Remainder O

Lined Area TBD O

6.13.3 Liner System

Composite liner system

consisting of 1.5 mm (60

mil) smooth LLDPE

geomembrane underlain by

0.3-m minimum compacted

thickness of low-

permeability cohesive soil

layer

O

6.13.4 Hydraulic Head 0.6 m maximum on pad

liner

6.13.5 Drain Pipes

Perforated corrugated PE

secondary drain pipes

placed diagonally across

the pad to primary pipes.

Perforated corrugated PE

primary drain pipes placed

along the pad center and

pad toe to transfer pipe.

Solid HDPE transfer pipe

from the primary pipes to

the process ponds.

Drainage from transfer pipe

to either of the pregnant or

barren ponds is by valve

control

O

6.13.6 Capacity

Drain pipe capacity is the

solution flow plus infiltration

flow from the 100-yr/24-hr

storm

O

6.13.7 Cover Drain Fill

Drain pipes on pad

embedded within 0.6-m

minimum loose lift

thickness liner cover drain

fill comprised of free-

draining, hard and durable

granular material

O

6.14 HEAP CONSTRUCTION

6.14.1 Method of Construction Stacker C

6.14.2 Nominal Heap Lift Height, meters 8 8 8 8 O

6.14.3 Number of Lifts 13 13 O

6.14.4 Maximum Heap Height, meters 104 104 B

6.14.5 Ore Natural Angle of Repose 37 37 A

6.14.6 Downstream Face Heap Side Slope

Lift stacked at

approximately 1.3H:1V

loose lift natural angle-of-

repose, 5-m minimum

setback from perimeter

berm crest inside edge to

heap toe as needed for

solution application pipeline

and access corridors

O

6.14.7 Required Heap Volume, m3

80,000,000 80,000,000

Heap Width On Top of Heap , m 593 B

Heap Length On top of heap, m 1793 B

Total Heap Width, m 869 B

Total Heap Length, m 2069 B

Heap Volume

Calculated Total Heap Volume, m3

147,466,162 B

Calculated Total solids Heap Area, m2

1,798,035 B

6.2 HEAP LEACHING






DESIGN CRITERIA


Rev: P1


6.15.1 Number of Leach Cycles 1 1 1 1 C

6.15.2 Primary Extraction Time, days 30 30 30 30 A

Secondary Extraction Time, days 80 80 80 80 A

Total Extraction Time, days 110 110 110 110 B

6.15.3 Ore Feed Rate, Tonnes per Day 14,085 16,197 28,169 32,394 B

Tonnes Under Leach 1,549,296 1,781,690 3,098,592 3,563,380 B

Heap Height, meters 8 8 8 8 O

Active Leach Area, Square Meters 129,108 148,474 258,216 296,948 B

Solution Application Rate, (l/hr) / m2

8 10 8 10 A

Average Barren Solution Flow Rate, m3 / hr 1,033 1,485 2,066 2,969 B

Solution Application Method Senniger Sprays Senniger Sprays Senniger Sprays Senniger Sprays O

6.15.4 Solution Pumping Availability, percent 98 98 98 98 A

6.15.5 Barren Solution Pumping Rate, m3 / hr 1,033 1,485 2,066 2,969 B

6.15.6 Antiscalant Dosage 6 ppm 6 ppm 6 ppm 6 ppm I

6.15.7 Solution Filters

Type Self-cleaning Self-cleaning Self-cleaning Self-cleaning I

Screen Aperture 0.25 mm (60 mesh) 0.25 mm (60 mesh) 0.25 mm (60 mesh) 0.25 mm (60 mesh) I

The following criteria describe the heap operation. Barren solution from the ADR plant will be pumped to the heap for distribution. Leach solution will be sprayed on the heap to promote

evaporation.






DESIGN CRITERIA


Rev: P1


6.15.8 General Water Balance

ROM Ore

Solids, Tonnes per Day 14,085 16,197 28,169 32,394 B

Average Solids, Tonnes per Daily Hour 587 675 1,174 1,350 B

Average Liquid, Tonnes per Daily Hour 18.2 20.9 36.3 41.7 B

Average Pulp, Tonnes per Daily Hour 605.0 695.8 1210.0 1391.5 B

Weight Percent Solids 97.00 97.00 97.00 97.00 A

Percent Moisture (Wet Weight Basis) 3.00 3.00 3.00 3.00 B

Percent Moisture (Dry Weight Basis) 3.09 3.09 3.09 3.09 B

Agglomerated Ore


Average Solids, Tonnes per Daily Hour 587 675 1174 1350 B



Weight Percent Solids 95.00 93.00 95.00 93.00 A


Percent Moisture (Dry Weight Basis) 5.26 7.53 5.26 7.53 B

Agglomeration Water Addition, Tonnes per Hour 12.74 29.92 25.47 59.85 B

Leached Ore


Average Solids, Tonnes per Daily Hour 586.9 674.9 1173.7 1349.8 B



Weight Percent Solids 89.53 89.53 89.53 89.53 B


Percent Moisture (Dry Weight Basis) 11.70 11.70 11.70 11.70 T

Leach Water Addition, Tonnes per Hour 37.77 28.16 75.55 56.33 B

6.15.9 Solution Evaporation Estimate

Percent of Solution Application Rate 6.00 6.00 6.00 6.00 A

Solution Evaporation Rate, (l/hr) / m2 0.48 0.60 0.48 0.60 B

Average, m3 / hr 62.0 89.1 123.9 178.2 B

6.15.10 Average Leach Makeup Water Estimate, Tonnes / hr 112.5 147.2 225.0 294.3 B

6.16 SOLUTION STORAGE (WITH GOLDER)

6.16.1 Number of Solution Ponds

Pregnant Solution 1 O

Intermediate Solution 1 O

Excess Solution 1 O

6.16.2 Number of Solution Tanks

PLS Surge Tank 1 K

Barren Solution 1 K

6.16.3 Basis for Sizing of PLS Surge Tank B

6.16.4 Basis for Sizing of Barren Tank B

6.16.5 Barren Tank Size, cubic meters 1,035 1,035 B

6.16.6 Nominal Tank Size, meters

Length 15.2 15.2 B

Width 15.2 15.2 B

Height 5.0 5.0 B

Free Board 0.5 0.5 A

6.16.7 Pregnant Solution Management

Pregnant Pond Number 1 O

Basis for Sizing of Pregnant Pond, Hours Retention 24-hour operational storage A

6.16.8 Intermediate Solution Management

Intermediate Pond Number 1 O

Basis for Sizing of Pregnant Pond, Hours Retention 6-hour operational storage A

6.16.9 Event Solution Management

Event Pond Number 1 O

Basis for Sizing of Event Pond, Hours Retention 100-yr/24-hr storm flow

from the leach pad and

collection pond areas, Plus

0.6 m freeboard

O

60 minutes of average PLS flown to the tank

to the tank to the tank

60 minutes of average 30 minutes of average

barren solution flowbarren solution flow






DESIGN CRITERIA


Rev: P1


6.17 CARBON ADSORPTION

6.17.1 Mechanical Availability, percent 98 98 A

6.17.2 Total Solution Processing

Method Carbon Column Adsorption K

Total Pregnant Solution Flow Rate, m3 / hr 991 1,337 1,981 2,675 B

Specific Flowrate, m3 / hr / m

2 60 81 60 81 B

Number of Trains 1 1 2 2 K

Number of Columns/Train 5 5 5 5 K

Type Upflow Cascade Upflow Cascade Upflow Cascade Upflow Cascade K

Nominal Column Dimensions, meters

Diameter 4.57 4.572 4.57 4.57 B

Height 6.40 6.40 6.40 6.40 B

Tonnes Carbon per Train 2 A

Design Metal Loading kg (Au+Ag)/t 5.0 6.0 5.0 6.0 A

Tenor of Pregnant Solution (nominal), ppm Au 0.52 0.52 0.52 0.52 B

Tenor of Pregnant Solution (nominal), ppm Ag 0.51 0.51 0.51 0.51 B

Average Gold to Carbon, kg Gold per Day 12.25 16.54 24.51 33.08 B

Average Silver to Carbon, kg Silver per Day 12.11 16.35 24.23 32.70 B

kg Gold + Silver per Day to Carbon 24.37 32.89 48.73 65.79 B

Loaded Carbon, Tonnes per Day 4.87 5.48 9.75 10.96 B

6.17.3 Carbon Advance Method Recessed Impellor Pump Recessed Impellor Pump I

6.17.4 Carbon-Solution Advance Flowrate, m3 / hr 40 40 B

6.17.5 Carbon type Coconut Shell Coconut Shell I

6.17.6 New Carbon Size, mesh 6 by 12 Mesh 6 by 12 Mesh K

6.17.7 Circuit Carbon Size, mesh 6 by 20 Mesh 6 by 20 Mesh K

6.17.8 Carbon bulk density, kg / m3 0.47 0.47 K

6.18 CARBON DESORPTION

6.18.1

kg Gold per Day to Carbon 12.25 16.54 24.51 33.08 B

kg Silver per Day to Carbon 12.11 16.35 24.23 32.70 B

kg Gold + Silver per Day to Carbon 24.37 32.89 48.73 65.79 B

Carbon Loading, kg per Tonne Carbon 5.00 6.00 5.00 6.00 A

Loaded Carbon, Tonnes per Day 4.87 5.48 9.75 10.96 B

6.18.2 Carbon Desorption Method Pressure Zadra Pressure Zadra K

6.18.3 Elution Temperature 135oC 150

oC 135

oC 150

oC K

6.18.4 Design Pressure 690 kilopascal (100 psi) 690 kilopascal (100 psi) K

6.18.5 Solution Chemistry

Weight percent NaCN 0.2 - 0.5 0.2 - 0.5 K

Weight percent NaOH 1.0 - 2.0 1.0 - 2.0 C

6.18.6 Solution Specific Gravity 1.02 1.02 1.02 1.02 A

6.18.7 Specific Heat of Solutions, Cal/(kg oC) 1.13 1.13 1.13 1.13 A

6.18.8 Solution Heater

Type Diesel or Propane fired

boiler

Diesel or Propane fired

boiler

A

6.18.9 Heat Exchangers

Type Plate and Frame Plate and Frame K

Number 3 3 K

6.18.10 Elution Column Carbon Capacity, tonnes (Calculated based on 1.5 cycle per days)3.2 3.7 6.5 7.3 K

The following criteria describe the precious metal recovery plant. Pregnant solution is pumped to the carbon columns. One train of five stages in series are installed. The train will be

stepped down to allow gravity flow of solution from stage to stage. Carbon will be advanced counter currently through the columns using a recessed impeller pump mounted external

to the carbon columns. Carbon pump suction and discharge can be controlled by automatic valves so one pump serves all CIC tanks in the train. Loaded carbon from the number 1

column will be pumped to a dewatering screen which will discharge the carbon into the loaded carbon storage bin.

Barren solution from the number 5 column will flow through a wire sampler and over a vibrating safety screen. The solution from the safety screen will flow to the barren solution tank,

where the cyanide strength will be increased to the concentration required for leaching by adding 20 percent cyanide solution. Barren solution will be pumped through flow meters and

returned to the leach solution distribution system through a pipeline

Total Loaded Carbon

Carbon loaded with precious metals is processed in the Carbon Desorption circuit. As outlined below a pressure Zadra system is planned.






DESIGN CRITERIA


Rev: P1


Elution Column Carbon Capacity, tonnes (Design) 4.0 4.0 8.0 8.0 K

6.18.11 Number of Columns 1 2 K

6.18.12 Height: Diameter Ratio 4 4 K

6.18.13 Desorption Flowrate, Bed Volumes per Hour 2 2 K

6.18.14 Elution column capacity, m3 8.5 8.5 B

Elution column inside diameter, m 1.4 1.4 B

Elution column nominal height, m 5.6 5.6 B

Elution column internal freeboard, m 0.5 0.5 A

Elution column internal total height, m 6.1 6.1 B

Internal volume, m3 9.3 9.3 B

6.18.15 Vessel Bed Volumes, m3 8.5 8.5 B

Desorption Flow rate, cubic meters per hour 17.0 17.0 B

Hours per Strip 16.0 16.0 B

6.18.16 Stripping Schedule

Number of Elution Cycles per Week 8.5 9.6 8.5 9.6 B

Number of Elution Cycles per Day 1.2 1.4 1.2 1.4 B

6.18.17 Solution Heater Exchange System

Inlet Design, oC 65 65 65 65 K

Outlet, oC 135 150 135 150 K

6.18.18 Carbon Advance Rate, tonnes per day 4.87 5.48 9.75 10.96 B

6.18.19 Vessel Material Carbon Steel Carbon Steel A

6.19 ELECTROWINNING

6.19.1 Number of cells, operating 2 2 4 4 K

6.19.2 Number of cells, Installed 2 4 K

6.19.3 Electrowinning Temperature 85oC 85

oC K

6.19.4 Gold Recovery ± 70% per Pass Through

Cells

± 70% per Pass Through

Cells

A

6.19.5 Mechanical Availability 98% + 98% + A

Pregnant solution from the elution vessel will flow through heat exchangers for heat recovery. Solution will then flow through electrowinning cells to the barren electrolyte tank. Two

electrowinning cells are provided so that metal can be recovered when one cell is being cleaned. Solution from the barren tank is pumped through the heat exchangers and into the elution

column to close the loop.






DESIGN CRITERIA


Rev: P1


6.19.6 Metal to be recovered from CIC carbon (maximum), kg per cycle

Gold 16.34 24.81 32.68 49.63 B

Silver 16.15 24.53 32.30 49.06 B

Other 4.04 6.13 8.08 12.26 A

Total 36.53 55.47 73.05 110.95 B

Current efficiency, %

Gold 14% 14% 14% 14% A

Silver 40% 40% 40% 40% A

Other 80% 80% 80% 80% A

Plating time, minutes

Gold 511 511 511 511 B

Silver 511 511 511 511 B

Other 511 511 511 511 B

Plating current, Amps

Gold 1,866 2,834 3,732 5,669 B

Silver 1,179 1,791 2,358 3,581 B

Other 250 380 500 760 B

Total 3,295 5,005 6,591 10,010 B

6.19.7 Eluate flow, m3

/ hr 15.6 15.6 31.1 31.1 B

6.19.8 Linear flow velocity, meters per hour 10.8 10.8 10.8 10.8 B

6.19.9 Cathode Cross sectional area, square meters 0.72 0.72 0.72 0.72 K

6.19.10 Superficial current density, A / m2 of cathode 125.00 125.00 125.00 125.00 K

6.19.11 Cathodes required per cell 28 28 B

6.19.12 Number of rectifiers 1 4 K

6.19.13 Rectifier voltage 6 6 K

6.19.14 Cathodes Stainless Steel Wool Stainless Steel Wool Stainless Steel Wool Stainless Steel Wool K

Number 28 28 B

6.19.15 Anodes Stainless Steel Stainless Steel Stainless Steel Stainless Steel K

Number 29 29 B

6.19.16 Overall Recovery 95+ 95+ 95+ 95+ A

6.19.17 Rectifier Amperage, Maximum 2,502 2,502 B

6.20 ACID WASH

6.20.1 Acid Type Hydrochloric Acid Hydrochloric Acid K

6.20.2 Acid Wash Solution Strength, weight percent HCl 3 3 K

6.20.3 Number of Acid Wash Columns 1 1 K

6.20.4 Acid Wash Vessel Material FRP FRP K

6.20.5 Acid Delivery Containers 1 m3 Tote Bins 1 m

3 Tote Bins K

6.20.6 Schedule Every Elution Cycle Every Elution Cycle K

6.20.7 Carbon Capacity, tonnes 4.0 4.0 K

6.20.8 Acid Solution Flow Rate, Bed Volumes per Hour 2 2 K

6.20.9 Acid Wash Solution Flowrate, m3 / hr 15.6 31.1 K

6.20.10 Wash Time 4 to 6 Hours 4 to 6 Hours A

6.20.11 Wash Temperature Ambient Ambient I

6.20.12 Mechanical Availability 98% + 98% + A

6.20.13 Neutralization Yes Yes A

Caustic to pH 7 Caustic to pH 7 K

6.21 CARBON REACTIVATION

6.21.1 Carbon Reactivation Kiln Type Horizontal Rotary Indirect Horizontal Rotary Indirect K

6.21.2 Feed Hopper Capacity, Tonnes 3 3 K

6.21.3 Number 1 1 I

6.21.4 Type Horizontal Rotary Horizontal Rotary I

6.21.5 Throughput rate, kg per hour

kg per hour 500 500 K

Operating Time, Percent 47.5 95 A

tonnes per day 5.7 11.4 B

Percent of Carbon Reactivated 104% 104% I

6.21.6 Retention Time, Minutes 10 at temperature 10 at temperature K

To maintain acceptable carbon activity, the carbon must be acid washed. Hydrochloric acid is planned for this service. The carbon can be acid washed before or after the elution process.

Before carbon is returned to the CIC process, it is thermally reactivated. One reactivation kiln is provided.






DESIGN CRITERIA


Rev: P1


6.21.7 Heating Method Indirect propane or diesel Indirect propane or diesel A

6.21.8 Carbon, Average oC 600 600 K

6.21.9 Carbon, Design oC 850 850 K

6.21.10 Quench Tank Capacity, Tonne 3 3 K

6.21.11 Carbon Properties

Size, Process 6 by 20 Mesh 6 by 20 Mesh K

Size, New 6 by 12 Mesh 6 by 12 Mesh K

Void Fraction, Percent

(Wet Settled Bed) 40 40 A

Specific Gravity of Dry Carbon 2.23 2.23 A

Bulk Density of Settled Carbon, Kilograms per cubic Meter

Attritted Circuit Carbon 600 600 A

Virgin Carbon 450 450 A






DESIGN CRITERIA


Rev: P1


6.21.12 Kiln Dewatering Screen

Aperture 16 Mesh 16 Mesh K

Screen Type Vibrating Vibrating K

6.21.13 Carbon Sizing Screen

Aperture 16 Mesh 16 Mesh K

Screen Type Vibrating Vibrating K

6.22 GOLD SMELTING

6.22.1 Sludge Filter

Wet Precipitate, kg / day 42.9 57.9 85.8 115.9 B

Sludge Moisture % 15 15 15 15 E

Dry Precipitate, kg / day 36.5 49.2 73.0 98.5 B

Dore kg / day 24.4 32.9 48.7 65.8 B

Electrowon Metal Grade (Estimated Data), %

Gold 33.6 33.6 33.6 33.6 B

Silver 33.2 33.2 33.2 33.2 E

Other 33.2 33.2 33.2 33.2 E

Total 100.0 100.0 100.0 100.0 B

Silver to Gold Ratio 1.0 1.0 1.0 1.0 B

Electrowon Material Mass (Design Basis), kg dry / day

Gold 12.25 16.54 24.51 33.08 B

Silver 12.11 16.35 24.23 32.70 B

Other 12.11 16.35 24.23 32.70 E

Design Total 36.5 49.2 73.0 98.5 B

Design Factors 1 1.15 1 1.15 A

Recovery Variation 1 1.1 1 1.1 A

Max Total 36.5 62.3 73.0 124.6 B

Sludge Density Estimates, t/m3

Typical Gold 2.7 2.7 E

Design Gold 2.0 2.0 E

Typical Silver 1.5 1.5 E

Design Silver 1.1 1.1 E

Filter Compression Rate 2.0 2.0 E

Expected Volume, m3 / day 0.047 0.096 B

Expected Volume, ft3 / day 1.67 3.39 B

Filter Selection (Design Basis)

Volumetric Size, m3 0.28 0.28 E

Filter Type Plate/Frame Plate/Frame K

Filter Specifications

Number of Units 1 1 K

Cycles per week 1.2 2.4 B

6.22.2 Retort

Precious Metal Produced per Week, kg 341 461 B

Retort Furnace K

Cake Precipitate Volume, m3 / day 0.047 0.096 B

Broken Cake Expansion, % 50 50 E

Cake Drying Volume, m3 / day 0.071 0.144 B

Cake Drying Volume, ft3 / day 2.5 5.1 B

Oven Volume, m3 0.25 0.25 E

Number of Units 1 1 K

Retort Cycle, Days (at 60% fill) 2.1 1.0 B

Cycle Time, hours 16 B

Max Temperature, degrees C 550 K

Pressure, torr 150 K

Operation Time, Hours 22 K

Condenser Water Bath K

Propane or diesel

The following section provides general criteria for the precious metal refining area. Since the sludge handling and refining are batch operations, additional capacity can provided by running

additional batches as required.






DESIGN CRITERIA


Rev: P1


6.22.3 Smelting Furnace

Type C

Crucible Total Volume, Liters 500 500 E

Dried Precipitate Volume, m3 / day 0.047 0.096 B

Dry Precipitate, kg / day 73.0 98.5 B

Dore, kg / day 48.7 65.8 B

Flux Addition kg per kg Precipitate 4.5 3.8 E

Flux Mass, kg / day 219.9 252.9 B

Flux Bulk Density, t / m3 1.6 1.6 E

Flux Volume, m3 / day 0.137 0.158 B

Total Charge Volume, m3 / day 0.185 0.254 B

Calculated Smelts per Day 0.46 0.64 B

Calculated Smelts per Week 3.23 4.45 B

Crucible Fill Fraction 80 80 E

Crucible Size, m3 0.5 0.5 E

Number Required 1 K

Smelting Temperature, oC 1250 K

Meltdown Time, Hours 1.5 to 2.0 K

Precious Metal Produced per Week, kg 341 461 B

6.22.4 Flux Consumption

Borax, kg per Tonne Ore 0.01 A

Sodium Nitrate, kg per Tonne Ore 0.002 A

Silica, kg per Tonne Ore 0.0002 A

Sodium Carbonate, kg per Tonne Ore 0.001 A

6.23 REAGENTS

6.23.1 Activated Carbon

Type Coconut I

Consumption, kg per Tonne Ore 0.02 A

Container Size, kg 50 A

Annual Consumption, tonnes per Year 230.0 B

Consumption, bags per year 4600 B

Shipping Container Bags A

6.23.2 Acid

Type Hydrochloric A

Consumption, kg per Tonne Ore 0.09 I

Consumption, kg per Day 2,915 B

Consumption, Totes per Day 14.6 B

Delivered Concentration, Pct by Wt 98 I

Acid Delivery Containers 200 L Drum A

Usage Concentration, Pct by Wt 3 - 5 I

Storage Capacity, Months 0.5 A

Storage Capacity, kg 43,732 B

Storage Capacity, Drum 219 B

6.23.3 Sodium Hydroxide

Consumption, kg per Tonne Ore 0.02 I

Container Size, kg 25 A

Consumption, kg per Day 648 B

Consumption, bags per day 26 B

Delivered Concentration, Pct by Wt 90 (prills) A

Storage Concentration, Pct by Wt 25 I

Mix Tank (Tank and Distribution Heat Traced and Insulated)

Diameter, meters 1.5 K

Height, meters 1.8 K

Volume, cubic meters 2.7 B

Sodium Hydroxide Capacity, kg 844 B

Propane or diesel

A reagent storage and mixing facility is required for activated carbon management as well as hydrochloric acid, sodium hydroxide, sodium cyanide, descalant, and lime

addition. This installation is integral to the processing plant and will have container handling equipment, mixing and slaking systems, and reagent metering pumps.






DESIGN CRITERIA


Rev: P1


Dry Storage Capacity, Months 0.5 A

Dry Storage Capacity, kg 9,718 B

Dry Storage Capacity, Bags 389 B

6.23.4 Sodium Cyanide

Consumption, kg per Tonne Ore 0.40 T

Container Size, kg 80 kg drums A

Consumption, kg per Day 12,958 B

Consumption, Drums per day 162.0 B

Delivered Concentration, Pct by Wt 98 (briquette) A

Storage Concentration, Pct by Wt 20 A

Dry Storage Capacity, Months 0.5 A

Dry Storage Capacity, kg 194,366 B

Dry Storage Capacity, Boxes 2,430 B

Mix Tank




Sodium Cyanide Capacity, kg 2,706 B

Storage Tank




Sodium Cyanide Capacity, kg 4,060 B

6.23.5 Antiscalant

Container 200 liter drums C

Average Addition Rate 6 ppm I

Consumption, kg per Day 428 B

Consumption, drums per day 2 B

Shipping Container drums A

Storage Capacity, Months 1 A

Storage Capacity, kg 12,828 B

Storage Capacity, Drums 64 B

6.23.6 Lime

Consumption, kg per Tonne Ore

Oxide 0.5 T

Daily Consumption, Tonnes 100% basis 27.0 B

Delivered Concentration, Pct CaO 60 A

Shipping Container Bulk A

Storage Capacity, Days 7 A

Storage Capacity, Tonnes 189 B

Silo Capacity, Tonnes 200 K

Delivery

Truck Capacity, tonnes 45 A

Average Trucks per Day 1 B

6.23.7 Cement

Average Consumption, kg/t 5 5 5 5 T

Daily Consumption, Tonnes 100% basis 270.0 B

Delivered Concentration, Pct CaO 60 A

Shipping Container Bulk A

Storage Capacity, Days 1 A

Storage Capacity, Tonnes 270 B

Silo Capacity, Tonnes 300 K

Delivery

Truck Capacity, tonnes 45 A

Average Trucks per Day 10 B






DESIGN CRITERIA


Rev: P1


6.24 UTILITIES

6.24.1 Raw Water

Fire Water Flow, cubic meters per hour 227.1 A

Fire Fighting Time, hours 1 A

Fire Reserve, cubic meters 227.1 B

Raw Water Consumption, cubic meters per hour 294.3 B

Raw Water Live Retention Time, hours 8 A

Raw Water Surge Volume, cubic meters 2355 B

Total Volume, Fire Reserve + Raw Water, cubic meters 2582 B

Fire Reserve, percent 9 B

Approximate Tank Diameter, meters 14.9 B

Approximate Tank Height, meters 15.4 B

Tank Base Elevation, meters TBD

Plant Floor Elevation, meters TBD

Raw Water Head (Nominal), meters #VALUE!

6.24.2 Potable Water

Potable Water Consumption, cubic meters per hour 11.355 B

Potable Water Live Retention Time, hours 2 A

Potable Water Surge Volume, cubic meters 23 B

Approximate Tank Diameter, meters 6.1 B

Approximate Tank Height, meters 6.6 B

Tank Base Elevation, meters TBD

Plant Floor Elevation, meters TBD

Potable Water Head (Nominal), meters #VALUE!

6.24.3 Process Water

Loop Flow rate, cubic meters per hour 22.2

Loop Discharge Pressure, kPa 414

6.24.4 Air

Crusher Plant

Discharge Volume, Nm3 / Hr 60 A

Discharge Pressure, kPa 700

Process Plant

Discharge Volume, Nm3 / Hr 60 A

Discharge Pressure, kPa 700

Air is required for the crushing and ADR plant operation. Two separate systems, one in each area are planned.

The raw water distribution system provides raw water for process requirements such as process water makeup, reagent mixing, and gland water. The top portion of the mine head tank will

be used to provide raw water surge for the system. The bottom portion of the mine head tank will provide the fire water storage for the system.

The potable water distribution system provides well water for safety showers and other uses as required. A small head tank located near the raw water head tank will be used to provide

surge and head for the system.

The process water loop includes pumps and piping to provide process water in the plant area. The process water will be pumped to all ADR process areas and may contain solids and other

chemical contaminates. The water is used to facilitate carbon movement.




Appendix 2

Equipment List

KD Engineering

Tucson, Arizona

P1 JJ 6/7/2011 BCS 6/7/2011 Preliminary 4

LYDIAN INTERNATIONAL LTD. APPROVAL

DATE:

DESCRIPTIONDATE

SIGNATURE:

BY DATEKDE

APPR

DOCUMENT NO: Q439-03-008

PAGES

LYDIAN INTERNATIONAL LTD.

EQUIPMENT LIST

Amulsar - Phase II Preliminary Economic Assessment

REV

NO

Q439-03-008 Description: Amulsar - Phase II Preliminary Economic Assessment

P1 Date:

Area 10 - Primary Crushing

10-AF-108 Apron Feeder 1 VFD - 2.3m x 10m (90" x 33') 14.7 11.0

10-HO-100 Dump hopper 1 200 tonne -

10-RB-105 Rock Breaker 1 75.1 56.0

10-SN-110 Grizzly Feeder 1 HRVGF 64" x 24' w/VL14 Mechanism MM13011 40.2 30.0

10-CR-115 Primary Crusher 1 C160 Jaw Crusher 335.1 250.0

10-CV-120 Primary Crusher Discharge Conveyor 1 1219mm x 15m (48" x 49') 30.2 22.5

10-MA-125 Tramp Iron Magnet 1 Suspension type 54" with rectifier 6.7 5.0

10-CP-130 Air Compressor 1 Model MM37 29.5 22.0

10-PV-130 Air Receiver 1 55.0 41.0

10-CV-135 Primary Crusher Transfer Conveyor 1 1372mm x 312m (54" x 1024') 100.5 75.0

10-DC-140 Dust Collection 1 30,000 CFM; DFO 4-64 99.9 74.5

10-AD-150 Air Dryer 1 Model TS1A-50; with filter IRGP275 2.9 2.2

10-CN-160 Crane Hoist 1 37 ton electric monorail 20.1 15.0

10-AF-208 Apron Feeder 1 VFD - 2.3m x 10m (90" x 33') 14.7 11.0

10-HO-200 Dump hopper 1 200 tonne -

10-RB-205 Rock Breaker 1 75.1 56.0

10-SN-210 Grizzly Feeder 1 HRVGF 64" x 24' w/VL14 Mechanism MM13011 40.2 30.0

10-CR-215 Primary Crusher 1 C160 Jaw Crusher 335.1 250.0

10-CV-220 Primary Crusher Discharge Conveyor 1 1219mm x 15m (48" x 49') 30.2 22.5

10-MA-225 Tramp Iron Magnet 1 Suspension type 54" with rectifier 6.7 5.0


SUBTOTAL Area 10 - Primary Crushing 1,332.0 993.7

Area 13 - Secondary Crushing

13-BN-105 Coarse Ore Storage Bin 1 24mdia x 31mheight -

13-AF-110 Apron Feeder 1 VFD, 1829mm x 7m (72" x 23') 10.1 7.5


13-CV-120 Coarse Ore Transfer Conveyor 1 1219mm x 105m (48" x 345') 201.1 150.0

13-SC-125 Secondary Vibrating Screen 1 RiplFlo XH 2400 x 6100 DD Screen 49.6 37.0

13-CR-130 Secondary Crusher 1 MP800 Cone Crusher 804.3 600.0

13-CV-135 Crusher Transfer Conveyor 1 1524mm x 235m (60" x 771') 33.5 25.0


13-WT-150 Belt Weightometer 1 0.1 0.1



13-PV-160 Air Receiver 1 1.3 1.0




13-CV-220 Coarse Ore Transfer Conveyor 1 1219mm x 105m (48" x 345') 201.1 150.0

13-SC-225 Secondary Vibrating Screen 1 RiplFlo XH 2400 x 6100 DD Screen 49.6 37.0

13-CR-230 Secondary Crusher 1 MP800 Cone Crusher 804.3 600.0

SUBTOTAL Area 13 - Secondary Crushing 2,363 1,763

Area 15 - Tertiary Crushing

15-BN-100 Fine Ore Screen Feed Bin 1 10mh x 7.3mw x 29.2ml; 2131m3 -

15-BF-105 Tertiary Screening Belt Feeder 1 VFD; 1677mm x 8m (66" x 26') 10.1 7.5


15-SC-110 Tertiary Vibrating Screen 1 MF 3600 x 7300 SD Screen 73.7 55.0


15-CV-125 Tertiary Screening Oversize Transfer Conveyor 1 1372mm x 220m (54" x 722') 201.1 150.0

15-BN-130 Tertiary Crusher Feed Bin 1 9.9mh x 5.4mw x 21.6ml; 1150m3 -

15-BF-135 Tertiary Crusher Belt Feeder 1 VFD; 1219mm x 11m (54" x 36') 10.1 7.5


15-CR-140 Tertiary Crusher 1 HP800 Cone Crusher 737.3 550.0




15-CV-150 Screen Tripper Conveyor 1 1524mm x 60m (60" x 197') 201.1 150.0

15-CV-127 Crusher Tripper Conveyor 1 1219mm x 60m (48" x 197') 201.1 150.0

15-CV-170 Fine Ore Collection Conveyor 1 1219mm x 60m (48" x 197') 201.1 150.0

15-CV-175 Fine Ore Transfer Conveyor 1 1219mm x 85m (48" x 279') 201.1 150.0


15-PV-180 Air Receiver 1 1.3 1.0










SUBTOTAL Area 15 - Tertiary Crushing 4,409.1 3,289.2

Area 17 - Lime Addtion

17-DC-105 Lime Bin Vent 1 10.1 7.5

17-BN-110 Pebble Lime Storage Silo 1 200 tonnes -

17-BX-115 Bin Activator 1 5.0 3.7

17-SJ-120 Screw Conveyor Feeder 1 5.0 3.7

17-CV-175 Crushed Ore Tripper Conveyor 1 VFD; 1372mm x 54m (54" x 177') 67.0 50.0

17-BN-180 Crushed Ore Surge Bin 1 9.9mh x 10mw x 40ml; 3920m3 -

17-BF-181 Belt Feeder 1 VFD; 1524mm x 8m (60" x 26') 10.1 7.5




17-CV-195 Overland Conveyor 1 1,067 mm x 4.5 km 1,675.6 1,250.0

SUBTOTAL Area 17 - Lime Addtion 1,802.8 1,344.9

Area 19 - Ore Stacking

19-CV-101 Leach Pad Grasshopper Conveyor 1 42" x 100' 25.0 18.7








19-CV-131 Leach Pad Radial Stacking Conveyor 1 402.1 300.0






KDKDKDKD EngineeringEngineeringEngineeringEngineering

HP kW

June 7, 2011

Spare Qty. Detail

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EQUIPMENT LIST

Code DescriptionOperating

Qty.

Phase II

Phase II

Phase II

Phase II


P1 Date:


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June 7, 2011

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EQUIPMENT LIST


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19-CV-124 Leach Pad Grasshopper Conveyor 42" x 100' 25.0 18.7

SUBTOTAL Area 19 - Ore Stacking 1,002.1 747.6

Area 20 Heap Leach

PD - 2001 Pregnant Solution Pond 1 See Golder Design

PD - 2002 Event Pond 1 See Golder Design

PP - 2001 Pregnant Solution Pump 1

Vertical Turbine type, Goulds Model VIT-FF, Size

14 RJHC, 2 stages, 110 LPS (1750gpm) TDH 30 M

(100 ft)

120.6 90.0




(100 ft)

120.6 90.0




(100 ft)

120.6 90.0

PP - 2004 Event Pond Pump 1Self priming type, Gorman Rupp, Model 04A20-

B4x4, 31.5 LPS (500gpm), TDH 33.5 (110 ft)50.9 38.0

PP - 2005 Event Pond Pump 1Self priming type, Gorman Rupp, Model 04A20-

B4x4, 31.5 LPS (500gpm), TDH 33.5 (110 ft)50.9 38.0

PD - 2003 Intermediate Solution Pond 1 See Golder Design -

PP - 2006 Pregnant Solution Pump 1 134.0 100.0



SUBTOTAL Area 20 Heap Leach 866.0 646.0

Area 1 Carbon Adsorption

CC 101 Carbon Column No 1 1

Cascade Type, Four Tonne Carbon Capacity / 1033

cubic meter per hour solution flow, CS, Tanks

w/304SS Diaphragm Includes 0.25 meter

Magmeter

-




w/304SS Diaphragm

-




w/304SS Diaphragm

-




w/304SS Diaphragm

-




w/304SS Diaphragm

-

CP 101 Control Panel CIC Columns 1 CIC Valve Control Panel -

PC 101 Carbon Transfer Pump 1

Hidrostal Type, 31.8 cubic meters per hour, 4.7

meters TDH, 35% Solids Carbon, NaCN solution,

Graphite packing

2.9 2.2

PS 101 Carbon Adsorption Sump Pump 156.8 cubic meters per hour, 10 meters TDH, 5 %

Solids Carbon, NaCN Solution5.0 3.7

SA 101 Wire Sampler 1 Pregnant solution sampler -

SA 102 Wire Sampler 1 Barren solution sampler -

SC 101 Trash Screen 1 Trash screen, 4 Mesh -

SC 102 Carbon Safety Screen 1 Carbon safety screen, 35 Mesh screen -

SS 101 Safety Shower and Eye Wash Station 1 CIC Area Safety shower and eye wash station -

CC 111 Carbon Column No 1 B 1



w/304SS Diaphragm

-




w/304SS Diaphragm

-




w/304SS Diaphragm

-




w/304SS Diaphragm

-




w/304SS Diaphragm

-

CP 111 Control Panel CIC Columns 1 CIC Valve Control Panel -

PC 111 Carbon Transfer Pump 1



Graphite packing

2.9 2.2

SA 111 Wire Sampler 1 Pregnant solution sampler -

SA 112 Wire Sampler 1 Barren solution sampler -

SC 111 Trash Screen 1 Trash screen, 4 Mesh -

SC 112 Carbon Safety Screen 1 Carbon safety screen, 35 Mesh screen -

PP-3004 Barren Solution Pump 1 VFD; 300 HP 294.9 220.0

SUBTOTAL Area 1 Carbon Adsorption 305.8 228.1

Area 2 Acid Wash

FA 201 Acid Vent Fan 1 170 cubic meter per hour at 51 mm water column 0.5 0.4

PC 201 Acid Wash Carbon Pump 1


meters TDH, 35% Solids Carbon, HCl solution,

Graphite packing

2.9 2.2

PP 201 Acid Circulation Pump 18.5 cubic meters per hour, 8 meters TDH,

Hydrochloric Acid Solution1.0 0.8

PP 202 Concentrated Hydrochloric Acid Transfer Pump 14.6 cubic meters per hour, 6.1 meters TDH,


PP 203 Neutralization Pump 132 cubic meters per hour, 8 meters TDH,


PS 201 Acid Area Sump Pump 132 cubic meters per hour, 8 meters TDH,


SS 201 Safety Shower and Eye Wash 1 Safety shower and eye wash -

TK 201 Acid Wash Vessel 1 FRP, 4.0 Ton -

TK 202 Dilute Acid Tank 1 FRP, 1.8 meter diameter by 2.0 meter high -

TK 203 Neutralization Tank FRP, 1.8 meter diameter by 2.6 meter high -

SUBTOTAL Area 2 Acid Wash 10.1 7.5

Phase II


P1 Date:


HP kW

June 7, 2011

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Document No:

Revision:

EQUIPMENT LIST


Qty.

Area 3 Carbon Strip

PC 301 Stripped Carbon Pump 1



Graphite packing

2.9 2.2

TK 301 Strip Vessel 1 Carbon Strip Column, 4 tonne capacity -

TK 303 Blowoff Column 1 -

TK 304 Blowoff Column 1 -

SS 301 Safety Shower and Eye Wash 1 Strip Area Safety Shower -

TK 302 Strip Vessel 1 Carbon Strip Column, 4 tonne capacity -

SUBTOTAL Area 3 Carbon Strip 2.9 2.2

Area 4 Strip Solution Handling

FL 401 Carbon Bucket Trap 1 35 mesh -

FL 402 Carbon Bucket Trap 1 35 mesh -

HE 401 Electric Immersion Heater Skid 1Skid Mounted Vendor Package Unit includes

control unit, 400kW536.2 400.0

HX 401 Recovery Heat Exchanger 1 Plate and Frame, Stainless Steel plates -

HX 402 Trim Heat Exchanger 1 Shell and Tube, carbon steel -

HX 403 Trim Heat Exchanger 1 Shell and Tube, carbon steel -

PP 401 Strip Solution Pump 18.5 cubic meters per hour, 53.4 meters TDH, NaCN

/ NaOH solution at 80 degrees Celsius7.4 5.5

PP 402 Strip Solution Pump 18.5 cubic meters per hour, 53.4 meters TDH, NaCN

/ NaOH solution at 80 degrees Celsius7.4 5.5

PS 401 Strip Area Sump Pump 116 cubic meters per hour, 12 meters TDH, 5 %

Solids Carbon, NaCN / NaOH Solution2.0 1.5

TK 401 Barren Strip Solution Tank 1 2.7 meters diameter by 3.6 meters high -

SUBTOTAL Area 4 Strip Solution Handling 552.9 412.5

Area 5 Electrowinning and Refining

EC 501 Electrowinning Cell 1Model 75EC20 complete with 21 anodes, 20 SS

cathodes and buss bar-



FA 501 E / W Exhaust Fan 12550 cubic meters per hour, 152 mm water

column, cyanide vapors5.0 3.7

FA 502 Collector Exhaust Fan 12550 cubic meters per hour, 203 mm water

column, fine particulate at 90 degrees Celsius10.1 7.5

FL 501 Filter Press 1Electrowinning sludge filter, plate and frame, 85

Liter Pneumatic / Hydraulic Closure-

FL 502 Refinery Sump Filter 1 -

FL 503 Furnace Dust Collector 1 142 square meter, 90 degrees Celsius -

MX 501 Flux Mixer 1 95 liter batch size 0.5 0.4

MF 501 Bullion Induction Furnace 1Induction furnace, Induction, 75 kW, 0.7 cubic

foot, with Hydraulics & Chiller100.5 75.0

MR 501 Retort 1Retort, Electric, 2 cubic foot, with Hg Trap and

Chiller33.5 25.0

PP 501 Filter Feed Pump 1Electrowinning sludge pump, 20 cubic meters per

hour at 70 meters TDH-

PS 501 Refinery Sump Pump 1Vertical rubber covered, 16 cubic meters per hour

at 12 meters TDH2.0 1.5

RC 501 Rectifier 1 DC Rectifier, SCR, 0-1500 A, 0-9 V 33.5 25.0


SA 501 Wire Sampler 1 Pregnant strip solution sampler -

SA 502 Wire Sampler 1 Barren strip solution sampler -

SP 501 High Pressure Sprayer 1 1.1 cubic meter per hour at 2000 meters TDH 10.1 7.5

SS 501 Safety Shower and Eye Wash 1 Safety shower and eye wash station -

EQ 501 Miscellaneous Equipment 1Vault door GSA Class 5, Electronic Lock, Includes

Day Gate & Wall Flange-







SUBTOTAL Area 5 Electrowinning and Refining 295.7 220.6

Area 6 Carbon Regeneration and Handling

AG 601 Carbon Fines Tank Agitator 1 2.0 1.5

AG 602 Carbon Attrition Agitator 1 2.0 1.5

FA 601 Carbon Regeneration Draft Fan 1 2.9 2.2

FE 601 Kiln Screw Feeder 1 Kiln feed screw conveyor SS Screw 0.8 0.6

FL 601 Carbon Fines Filter 1 Plate and Frame, 425 liter -

KN 601 Carbon Reactivation Kiln 1Regeneration kiln package, Horizontal, Electric 120

kW174.3 130.0

PC 601 Activated Carbon Pump 1



Graphite packing

2.9 2.2

PP 601 Carbon Fines Pump 1Carbon Fines Pump, 9.1 cubic meters per hour,

56.4 meters TDH, NaCN / NaOH solution7.4 5.5

SC 601 Kiln Dewatering Screen 1 1220 mm diameter, 16 mesh screen 2.9 2.2

SC 602 Carbon Sizing Screen 1 1220 mm diameter, 16 mesh screen 2.9 2.2

TK 601 Kiln Feed Bin 1 2 tonne -

TK 602 Carbon Quench Tank 1Quench tank, Carbon Steel, 2 meters diameter by

1.6 meters vertical side-

TK 603 Activated Carbon Storage Tank 1Carbon Storage Tank, 2 meters diameter by 2

meters vertical side-

TK 604 Carbon Fines Tank 1Carbon Fines Storage Tank, 1.8 meters diameter

by 1.8 meters high-

TK 605 Carbon Attrition Tank 1Carbon Attrition Tank, 1.8 meters diameter by 1.8

meters high-

SUBTOTAL Area 6 Carbon Regeneration and Handling 198.2 147.9

Area 7 Reagent Mix / Storage

AG 701 NaCN Mix Agitator 1 2.0 1.5

BV 701 Bin Vent 1 2.0 1.5

BV 702 Bin Vent 1 2.0 1.5

CN 701 Crane 1 1 Tonne Hoist 1.0 0.8

PP 701 Caustic Mix / Transfer Pump 1Caustic mixing/circulating pumps 4.5 m3/hr, 7.6 m

tdh, 25% NaOH1.0 0.8

PP 702 NaCN Transfer Mix Pump 1Cyanide Transfer pump, 17 m3/hr, 10.7 m tdh,

30% NaCN1.5 1.1

PP 703 NaCN Distribution Pump 1Cyanide Distribution pump, 9.1 m3/hr, 30.5 m tdh,

30% NaCN2.9 2.2

PP 704 NaCN Distribution Pump 1Cyanide Distribution pump, 9.1 m3/hr, 30.5 m tdh,

30% NaCN2.9 2.2

PP 705 Antiscalant Metering Pump 1 0 to 100 milliliters per minute, 10 meters TDH 0.7 0.5

PP 706 Antiscalant Metering Pump 1 0 to 100 milliliters per minute, 10 meters TDH 0.7 0.5

Phase II

Phase II


P1 Date:


HP kW

June 7, 2011

Spare Qty. Detail

Document No:

Revision:

EQUIPMENT LIST


Qty.

PS 701 Reagent Area Sump Pump 1Vertical rubber covered, 16 cubic meters per hour

at 12 meters TDH2.0 1.5

SS 701 Safety Shower and Eye Wash 1 Safety shower and eye wash station -

TK 701 Caustic Mix / Storage Tank 1Carbon steel, 1.5 meter diameter by 1.8 meter

high-

TK 702 NaCN Mix Tank 1Carbon Steel, 2.5 meter diameter x 2.8 meter high,

w/SuperSack Bag Breaker & Hoist-

TK 703 NaCN Storage Tank 1 Carbon Steel, 2.9 meter diameter x 3.2 meter high -

SUBTOTAL Area 7 Reagent Mix / Storage 18.8 14.0

Area 30 Barren Solution Pumping

CP-3001 Air Compressor 1 Air Compressor 10.1 7.5

DR-3001 Air Dryer 1 Air Dryer 1.3 1.0

FL-3001 Barren Solution Liquid-solids separation system 1

Filter, OREMAX torando filter 10 inch, Self

Cleaning, three filter units, W/ inlet and outlet

headers, 14 inch flanges, rated 200 psi, NEMA 4

controller station complete with valves.

-

PD - 3001 Containment Pond 1Barren Solution Pump 854 m3 per hour design

total flow-

PP - 3001 Barren Solution Pump 1

Vertical Turbine in can type, Goulds Model VIC-L,

Size 14 RJHC, 8 stages, 120 LPS (1900gpm) initial

TDH 73 M (240 ft) final TDH 122 M (200 ft)

300.3 224.0





300.3 224.0





300.3 224.0





300.3 224.0

PP - 3013 Process Solution Pump 1

Horizontal centrifugal, Goulds Model 3196, Size

2x3 - 8MTX, 12.6 LPS (200 gpm), TDH 36.6 M (120

ft)

15.0 11.2

PP - 3014 Process Solution Pump 1

Horizontal centrifugal, Goulds Model 3196, Size

2x3 - 8MTX, 12.6 LPS (200 gpm), TDH 36.6 M (120

ft)

15.0 11.2

PP - 3006 Containment Pump 1Self priming type, Gorman Rupp, Model 04A20-

B4x4, 31.5 LPS (500gpm), TDH 33.5 (110 ft)24.9 18.6

PP - 3007 Barren Solution Sump Pump 1 10.1 7.5

TK - 3006 Barren Solution Surge Tank 1 Insulated, -

VS - 3001 Air Receiver 1 -

SUBTOTAL Area 30 Barren Solution Pumping 1,277.5 953.0

Area 90 Auxiliary Equipment

GN - 9001 Emergency Generator 1

Emergency Diesel Generator, 1460KW, 380volt,

50Hz,1500rpm, 0.8pf, with radiator cooling and

main circuit breaker, installed in 40ft ISO

container, with approx 1000gal fuel tank

-

GN - 9002 Emergency Generator 1

Emergency Diesel Generator, 1460KW, 380volt,

50Hz,1500rpm, 0.8pf, with radiator cooling and

main circuit breaker, installed in 40ft ISO

container, with approx 1000gal fuel tank

-

SUBTOTAL Area 90 Auxiliary Equipment - -

TOTAL 14,436.9 10,769.9



Appendix 3

Capital Cost Detail

Project No: 439-02

Amulsar Phase I Revision: P2Date: June 7, 2011

Description Equipment Costs Method / Factor Installed Cost

DIRECT COSTS

Area 10 - Primary Crushing 3,242,143 Eqpmt. Cost x 1.43 4,636,264

Area 13 - Secondary Crushing 2,952,800 Eqpmt. Cost x 1.43 4,222,504

Area 15 - Tertiary Crushing 6,626,178 Eqpmt. Cost x 1.43 9,475,435

Area 17 - Product Storage Loadout & Overland Conveyor 16,386,367 Eqpmt. Cost x 1.43 23,432,505

Area 20 Heap Leach 805,000 Eqpmt. Cost x 1.43 1,151,150

ADR Plant 5,200,000 Eqpmt. Cost x 1.43 7,436,000

Area 30 Solution Management 675,000 Eqpmt. Cost x 1.43 965,250

Area 90 Auxiliary Equipment 50,000 50,000

Equipment Cost 35,937,488 Installed Equipment Cost 51,369,108

Process Piping (5% of Plant Equip. Installed cost) 2,568,455

Instrumentation (5% of Plant Equip. Installed cost) 2,568,455

Site Development (25% of Plant Equip. Installed cost) 12,842,277

Water & Electrical Power to Site (15% of Plant Equip. Installed cost) 7,705,366

TOTAL DIRECT COSTS 77,053,662

PLANT INDIRECT COSTSEPCM (12% of Direct Cost) 9,246,439 Construction Indirect Costs incl: (10% of Direct Cost) 7,705,366 Construction Supervision Equipment Rental Field office expense Mobilization / Demobilization Consumables Spare Parts (5% of Equipment Cost) 1,796,874 Initial Fill & Reagents (1.5% of Direct Cost) 1,155,805 Equipment Insurance & Freight Cost (7% of Direct Cost) 5,393,756

TOTAL INDIRECT COSTS 25,298,241

TOTAL DIRECT AND INDIRECT 102,351,903

CONTINGENCY 30% of Direct and Indirect 30,705,571

TOTAL 133,057,474

SUMMARY FACTORED COST ESTIMATE WORKSHEET

KD EngineeringOPERATING SPARE DESCRIPTION HP UNIT COST COST reference comments calculations

Area 10 ‐ Primary Crushing10‐AF‐108 Apron Feeder 1 350,000 350,000 VFD ‐ 2.3m x 10m (90" x 33')10‐HO‐100 Dump hopper 1 450,000 450,000 200 tonne10‐RB‐105 Rock Breaker 1 201,800 201,80010‐SN‐110 Grizzly Feeder 1 159,618 159,618 HRVGF 64" x 24' w/VL14 Mechanism MM1301110‐CR‐115 Primary Crusher 1 800,000 800,000 C160 Jaw Crusher10‐CV‐120 Primary Crusher Discharge Conveyor 1 49,000 49,000 1219mm x 15m (48" x 49')10‐MA‐125 Tramp Iron Magnet 1 16,175 16,175 Suspension type 54" with rectifier10‐CP‐130 Air Compressor 1 25,000 25,000 Model MM3710‐PV‐130 Air Receiver 1 3,000 3,00010‐CV‐135 Primary Crusher Transfer Conveyor 1 1,024,000 1,024,000 1372mm x 312m (54" x 1024')10‐DC‐140 Dust Collection 1 61,000 61,000 30,000 CFM; DFO 4‐6410‐AD‐150 Air Dryer 1 5,000 5,000 Model TS1A‐50; with filter IRGP27510‐CN‐160 Crane Hoist 1 97,550 97,550 37 ton electric monorailTOTAL Area 10 ‐ Primary Crushing 3,242,143

Area 13 ‐ Secondary Crushing13‐BN‐105 Coarse Ore Storage Bin 1 458,500 458,500 24mdia x 31mheight13‐AF‐110 Apron Feeder 1 175,000 175,000 VFD, 1829mm x 7m (72" x 23')13‐AF‐111 Apron Feeder 1 175,000 175,000 VFD, 1829mm x 7m (72" x 23')13‐CV‐120 Coarse Ore Transfer Conveyor 1 175,000 175,000 VFD, 1829mm x 7m (72" x 23')13‐SC‐125 Secondary Vibrating Screen 1 175,000 175,000 VFD, 1829mm x 7m (72" x 23')13‐CR‐130 Secondary Crusher 1 431,250 431,250 1219mm x 105m (48" x 345')13‐CV‐135 Crusher Transfer Conveyor 1 1,156,500 1,156,500 1524mm x 235m (60" x 771')13‐DC‐140 Dust Collection 1 61,000 61,000 30,000 CFM; DFO 4‐6413‐WT‐150 Belt Weightometer 1 15,000 15,00013‐CN‐155 Crane Hoist 1 97,550 97,550 37 ton electric monorail13‐CP‐160 Air Compressor 1 25,000 25,000 Model MM3713‐PV‐160 Air Receiver 1 3,000 3,00013‐AD‐165 Air Dryer 1 5,000 5,000 Model TS1A‐50; with filter IRGP275TOTAL Area 13 ‐ Secondary Crushing 2,952,800

Area 15 ‐ Tertiary Crushing15‐BN‐100 Fine Ore Screen Feed Bin 1 143,500 143,500 10mh x 7.3mw x 29.2ml; 2131m3 (562,950 gal)15‐BF‐105 Tertiary Screening Belt Feeder 1 130,000 130,000 VFD; 1677mm x 8m (66" x 26')15‐BF‐106 Tertiary Screening Belt Feeder 1 130,000 130,000 VFD; 1677mm x 8m (66" x 26')15‐SC‐110 Tertiary Vibrating Screen 1 101,800 101,800 MCS. ‐ 3.0 x 6.1m ‐ Single Deck15‐SC‐111 Tertiary Vibrating Screen 1 101,800 101,800 MCS. ‐ 3.0 x 6.1m ‐ Single Deck15‐CV‐125 Tertiary Screening Oversize Transfer Conveyor 1 902,500 902,500 $1250 x lf15‐BN‐130 Tertiary Crusher Feed Bin 1 67,500 67,500 ASSUMED 55% OF fine ORE BIN15‐BF‐135 Tertiary Crusher Belt Feeder 1 180,000 180,000 $5000/lf KDE databse15‐BF‐136 Tertiary Crusher Belt Feeder 1 180,000 180,000 $5000/lf KDE databse15‐CR‐140 Tertiary Crusher 1 1,397,514 1,397,514 KDE Database ‐ Tertiary Crusher ‐ Metso HP800 Cone Crusher15‐CR‐141 Tertiary Crusher 1 1,397,514 1,397,514 KDE Database ‐ Tertiary Crusher ‐ Metso HP800 Cone Crusher15‐DC‐155 Dust Collection 1 61,000 61,00015‐CN‐165 Crane Hoist 1 43,550 43,550 MCS 10 TON OH TROLLEY 41150@30'+ 120X20' ADD15‐CV‐150 Screen Tripper Conveyor 1 689,500 689,500 $3500/LF KDE databse15‐CV‐127 Crusher Tripper Conveyor 1 591,000 591,000 $3000/LF KDE databse15‐CV‐170 Fine Ore Collection Conveyor 1 197,000 197,000 $1000 x lf KDE databse15‐CV‐175 Fine Ore Transfer Conveyor 1 279,000 279,000 $1000 x lf KDE databse15‐CP‐180 Air Compressor 1 25,000 25,00015‐PV‐180 Air Receiver 1 3,000 3,00015‐AD‐185 Air Dryer 1 5,000 5,000TOTAL Area 15 ‐ Tertiary Crushing 6,626,178

Area 17 ‐ Product Storage Loadout & Overland Conveyor17‐DC‐105 Lime Bin Vent 1 50,000 50,000 ASSUMED17‐BN‐110 Pebble Lime Storage Silo 1 60,000 60,000 380 ton17‐BX‐115 Bin Activator 1 30,000 30,000 ASSUMED17‐SJ‐120 Screw Conveyor Feeder 1 25,000 25,000 KDE 17‐CV‐170 Overland Conveyor 1 9,289,510 9,289,510 4.5Km @ 900$/lf17‐CV‐175 Crushed Ore Tripper Conveyor 1 963,357 963,357 350M @ 1,200$/lf17‐CV‐176 Portable Conveyors 24 200,000 4,800,000 Budget Price17‐CV‐177 Radial Relescopping Conveyor 1 875,000 875,000 Budget Price17‐BN‐180 Crushed Ore Surge Bin 1 213,500 213,50017‐BF‐181 Belt Feeder 1 20,000 20,000 $5000/lf KDE databse17‐BF‐182 Belt Feeder 1 20,000 20,000 $5000/lf KDE databse17‐BF‐183 Belt Feeder 1 20,000 20,000 $5000/lf KDE databse17‐BF‐184 Belt Feeder 1 20,000 20,000 $5000/lf KDE databseTOTAL Area 17 ‐ Product Storage Loadout & Overland Conveyor 16,386,367

Area 20 Heap LeachPD ‐ 2001 Pregnant Solution Pond 1 60,000 60,000PD ‐ 2002 Event Pond 1 125,000 125,000 kde 429‐02 CCE WB ‐ Pregnant solution pump

EQUIPMENT LIST ‐ OPTION G: 2 x 5 MTPA, 12mm

PP ‐ 2001 Pregnant Solution Pump 1 60,000 60,000 Ref A‐ 6.3Mx7.1M ‐ Pregnant solution TankPP ‐ 2002 Pregnant Solution Pump 1 60,000 60,000 kde 429‐02 CCE WB ‐ Barren solution pumpPP ‐ 2003 Pregnant Solution Pump 1 1 60,000 120,000 kde 429‐02 CCE WB ‐45x45x15 ‐ Barren Solution TankPP ‐ 2004 Event Pond Pump 1 1 40,000 80,000 kde 429‐02 CCE WB ‐PP ‐ 2005 Event Pond Pump 1 1 40,000 80,000 kde 429‐02 CCE WB ‐ Take offPD ‐ 2003 Intermediate Solution Pond 1 100,000 100,000 kde 429‐02 CCE WB ‐PP ‐ 2006 Pregnant Solution Pump 1 40,000 40,000 kde 429‐02 CCE WB ‐PP ‐ 2007 Pregnant Solution Pump 1 40,000 40,000 kde 429‐02 CCE WB ‐ MonorailPP ‐ 2008 Pregnant Solution Pump 1 40,000 40,000 kde 429‐02 CCE WB ‐‐ 6'Dia x 6'High c/w 20kG bag brkrTOTAL Area 20 Heap Leach 805,000

ADR Plant 1 5,200,000 5,200,000CC 101 Carbon Column No 1 inc 0CC 102 Carbon Column No 2 inc 0CC 103 Carbon Column No 3 inc 0CC 104 Carbon Column No 4 inc 0CC 105 Carbon Column No 5 inc 0CP 101 Control Panel CIC Columns inc 0PC 101 Carbon Transfer Pump inc 0PS 101 Carbon Adsorption Sump Pump inc 0SA 101 Wire Sampler inc 0SA 102 Wire Sampler inc 0SC 101 Trash Screen inc 0SC 102 Carbon Safety Screen inc 0SS 101 Safety Shower and Eye Wash Station inc 0TOTAL ADR Plant 5,200,000

Area 2 Acid WashFA 201 Acid Vent Fan inc 0PC 201 Acid Wash Carbon Pump inc 0PP 201 Acid Circulation Pump inc 0PP 202 Concentrated Hydrochloric Acid Transfer Pump inc 0PP 203 Neutralization Pump inc 0PS 201 Acid Area Sump Pump inc 0SS 201 Safety Shower and Eye Wash inc 0TK 201 Acid Wash Vessel inc 0TK 202 Dilute Acid Tank inc 0TK 203 Neutralization Tank inc 0TOTAL Area 2 Acid Wash 0

Area 3 Carbon StripPC 301 Stripped Carbon Pump inc 0TK 301 Strip Vessel inc 0TK 303 Blowoff Column inc 0TK 304 Blowoff Column inc 0SS 301 Safety Shower and Eye Wash inc 0TOTAL Area 3 Carbon Strip 0

Area 4 Strip Solution HandlingFL 401 Carbon Bucket Trap inc 0FL 402 Carbon Bucket Trap inc 0HE 401 Electric Immersion Heater Skid inc 0HX 401 Recovery Heat Exchanger inc 0HX 402 Trim Heat Exchanger inc 0HX 403 Trim Heat Exchanger inc 0PP 401 Strip Solution Pump inc 0PP 402 Strip Solution Pump inc 0PS 401 Strip Area Sump Pump inc 0TK 401 Barren Strip Solution Tank inc 0TOTAL Area 4 Strip Solution Handling 0

Area 5 Electrowinning and RefiningEC 501 Electrowinning Cell inc 0EC 502 Electrowinning Cell inc 0FA 501 E / W Exhaust Fan inc 0FA 502 Collector Exhaust Fan inc 0FL 501 Filter Press inc 0FL 502 Refinery Sump Filter inc 0FL 503 Furnace Dust Collector inc 0MX 501 Flux Mixer inc 0MF 501 Bullion Induction Furnace inc 0MR 501 Retort inc 0PP 501 Filter Feed Pump inc 0PS 501 Refinery Sump Pump inc 0RC 501 Rectifier inc 0RC 502 Rectifier inc 0SA 501 Wire Sampler inc 0

SA 502 Wire Sampler inc 0SP 501 High Pressure Sprayer inc 0SS 501 Safety Shower and Eye Wash inc 0EQ 501 Miscellaneous Equipment inc 0TOTAL Area 5 Electrowinning and Refining 0

Area 6 Carbon Regeneration and HandlingAG 601 Carbon Fines Tank Agitator inc 0AG 602 Carbon Attrition Agitator inc 0FA 601 Carbon Regeneration Draft Fan inc 0FE 601 Kiln Screw Feeder inc 0FL 601 Carbon Fines Filter inc 0KN 601 Carbon Reactivation Kiln inc 0PC 601 Activated Carbon Pump inc 0PP 601 Carbon Fines Pump inc 0SC 601 Kiln Dewatering Screen inc 0SC 602 Carbon Sizing Screen inc 0TK 601 Kiln Feed Bin inc 0TK 602 Carbon Quench Tank inc 0TK 603 Activated Carbon Storage Tank inc 0TK 604 Carbon Fines Tank inc 0TK 605 Carbon Attrition Tank inc 0TOTAL Area 6 Carbon Regeneration and Handling 0

Area 7 Reagent Mix / StorageAG 701 NaCN Mix Agitator inc 0BV 701 Bin Vent inc 0BV 702 Bin Vent inc 0CN 701 Crane inc 0PP 701 Caustic Mix / Transfer Pump inc 0PP 702 NaCN Transfer Mix Pump inc 0PP 703 NaCN Distribution Pump inc 0PP 704 NaCN Distribution Pump inc 0PP 705 Antiscalant Metering Pump inc 0PP 706 Antiscalant Metering Pump inc 0PS 701 Reagent Area Sump Pump inc 0SS 701 Safety Shower and Eye Wash inc 0 TK 701 Caustic Mix / Storage Tank inc 0 TK 702 NaCN Mix Tank inc 0 TK 703 NaCN Storage Tank inc 0TOTAL Area 7 Reagent Mix / Storage 0

Area 30 Solution ManagementCP‐3001 Air Compressor 1 30,000 30,000DR‐3001 Air Dryer 1 15,000 15,000FL‐3001 Barren Solution Liquid‐solids separation system 1 30,000 30,000PD ‐ 3001 Containment Pond 1 50,000 50,000PP ‐ 3001 Barren Solution Pump 1 80,000 80,000PP ‐ 3002 Barren Solution Pump 1 80,000 80,000PP ‐ 3003 Barren Solution Pump 1 80,000 80,000PP ‐ 3004 Barren Solution Pump 1 80,000 80,000PP ‐ 3013 Process Solution Pump 1 60,000 60,000PP ‐ 3014 Process Solution Pump 1 60,000 60,000PP ‐ 3006 Containment Pump 1 40,000 40,000PP ‐ 3007 Barren Solution Sump Pump 1 7,500 7,500TK ‐ 3006 Barren Solution Surge Tank 1 57,500 57,500VS ‐ 3001 Air Receiver 1 5,000 5,000TOTAL Area 30 Solution Management 675,000

Area 90 Auxiliary EquipmentGN ‐ 9001 Emergency Generator 1 25,000 25,000GN ‐ 9002 Emergency Generator 1 25,000 25,000TOTAL Area 90 Auxiliary Equipment 50,000

Project No: 439-02

Amulsar Phase I Revision: P2Date: June 7, 2011


DIRECT COSTS

Area 10 - Primary Crushing 3,242,143 Eqpmt. Cost x 1.43 4,636,264

Area 13 - Secondary Crushing 2,952,800 Eqpmt. Cost x 1.43 4,222,504

Area 15 - Tertiary Crushing 6,626,178 Eqpmt. Cost x 1.43 9,475,435

Area 17 - Product Storage Loadout & Overland Conveyor 16,386,367 Eqpmt. Cost x 1.43 23,432,505

Area 20 Heap Leach 805,000 Eqpmt. Cost x 1.43 1,151,150

ADR Plant 5,200,000 Eqpmt. Cost x 1.43 7,436,000

Area 30 Solution Management 675,000 Eqpmt. Cost x 1.43 965,250

Area 90 Auxiliary Equipment 50,000 50,000


Process Piping (5% of Plant Equip. Installed cost) 2,568,455

Instrumentation (5% of Plant Equip. Installed cost) 2,568,455

Site Development (25% of Plant Equip. Installed cost) 12,842,277

Water & Electrical Power to Site (15% of Plant Equip. Installed cost) 7,705,366


PLANT INDIRECT COSTSEPCM (12% of Direct Cost) 9,246,439 Construction Indirect Costs incl: (10% of Direct Cost) 7,705,366 Construction Supervision Equipment Rental Field office expense Mobilization / Demobilization Consumables Spare Parts (5% of Equipment Cost) 1,796,874 Initial Fill & Reagents (1.5% of Direct Cost) 1,155,805 Equipment Insurance & Freight Cost (7% of Direct Cost) 5,393,756




TOTAL 133,057,474



Area 10 ‐ Primary Crushing10‐AF‐108 Apron Feeder 1 350,000 350,000 VFD ‐ 2.3m x 10m (90" x 33')10‐HO‐100 Dump hopper 1 450,000 450,000 200 tonne10‐RB‐105 Rock Breaker 1 201,800 201,80010‐SN‐110 Grizzly Feeder 1 159,618 159,618 HRVGF 64" x 24' w/VL14 Mechanism MM1301110‐CR‐115 Primary Crusher 1 800,000 800,000 C160 Jaw Crusher10‐CV‐120 Primary Crusher Discharge Conveyor 1 49,000 49,000 1219mm x 15m (48" x 49')10‐MA‐125 Tramp Iron Magnet 1 16,175 16,175 Suspension type 54" with rectifier10‐CP‐130 Air Compressor 1 25,000 25,000 Model MM3710‐PV‐130 Air Receiver 1 3,000 3,00010‐CV‐135 Primary Crusher Transfer Conveyor 1 1,024,000 1,024,000 1372mm x 312m (54" x 1024')10‐DC‐140 Dust Collection 1 61,000 61,000 30,000 CFM; DFO 4‐6410‐AD‐150 Air Dryer 1 5,000 5,000 Model TS1A‐50; with filter IRGP27510‐CN‐160 Crane Hoist 1 97,550 97,550 37 ton electric monorailTOTAL Area 10 ‐ Primary Crushing 3,242,143

Area 13 ‐ Secondary Crushing13‐BN‐105 Coarse Ore Storage Bin 1 458,500 458,500 24mdia x 31mheight13‐AF‐110 Apron Feeder 1 175,000 175,000 VFD, 1829mm x 7m (72" x 23')13‐AF‐111 Apron Feeder 1 175,000 175,000 VFD, 1829mm x 7m (72" x 23')13‐CV‐120 Coarse Ore Transfer Conveyor 1 175,000 175,000 VFD, 1829mm x 7m (72" x 23')13‐SC‐125 Secondary Vibrating Screen 1 175,000 175,000 VFD, 1829mm x 7m (72" x 23')13‐CR‐130 Secondary Crusher 1 431,250 431,250 1219mm x 105m (48" x 345')13‐CV‐135 Crusher Transfer Conveyor 1 1,156,500 1,156,500 1524mm x 235m (60" x 771')13‐DC‐140 Dust Collection 1 61,000 61,000 30,000 CFM; DFO 4‐6413‐WT‐150 Belt Weightometer 1 15,000 15,00013‐CN‐155 Crane Hoist 1 97,550 97,550 37 ton electric monorail13‐CP‐160 Air Compressor 1 25,000 25,000 Model MM3713‐PV‐160 Air Receiver 1 3,000 3,00013‐AD‐165 Air Dryer 1 5,000 5,000 Model TS1A‐50; with filter IRGP275TOTAL Area 13 ‐ Secondary Crushing 2,952,800

Area 15 ‐ Tertiary Crushing15‐BN‐100 Fine Ore Screen Feed Bin 1 143,500 143,500 10mh x 7.3mw x 29.2ml; 2131m3 (562,950 gal)15‐BF‐105 Tertiary Screening Belt Feeder 1 130,000 130,000 VFD; 1677mm x 8m (66" x 26')15‐BF‐106 Tertiary Screening Belt Feeder 1 130,000 130,000 VFD; 1677mm x 8m (66" x 26')15‐SC‐110 Tertiary Vibrating Screen 1 101,800 101,800 MCS. ‐ 3.0 x 6.1m ‐ Single Deck15‐SC‐111 Tertiary Vibrating Screen 1 101,800 101,800 MCS. ‐ 3.0 x 6.1m ‐ Single Deck15‐CV‐125 Tertiary Screening Oversize Transfer Conveyor 1 902,500 902,500 $1250 x lf15‐BN‐130 Tertiary Crusher Feed Bin 1 67,500 67,500 ASSUMED 55% OF fine ORE BIN15‐BF‐135 Tertiary Crusher Belt Feeder 1 180,000 180,000 $5000/lf KDE databse15‐BF‐136 Tertiary Crusher Belt Feeder 1 180,000 180,000 $5000/lf KDE databse15‐CR‐140 Tertiary Crusher 1 1,397,514 1,397,514 KDE Database ‐ Tertiary Crusher ‐ Metso HP800 Cone Crusher15‐CR‐141 Tertiary Crusher 1 1,397,514 1,397,514 KDE Database ‐ Tertiary Crusher ‐ Metso HP800 Cone Crusher15‐DC‐155 Dust Collection 1 61,000 61,00015‐CN‐165 Crane Hoist 1 43,550 43,550 MCS 10 TON OH TROLLEY 41150@30'+ 120X20' ADD15‐CV‐150 Screen Tripper Conveyor 1 689,500 689,500 $3500/LF KDE databse15‐CV‐127 Crusher Tripper Conveyor 1 591,000 591,000 $3000/LF KDE databse15‐CV‐170 Fine Ore Collection Conveyor 1 197,000 197,000 $1000 x lf KDE databse15‐CV‐175 Fine Ore Transfer Conveyor 1 279,000 279,000 $1000 x lf KDE databse15‐CP‐180 Air Compressor 1 25,000 25,00015‐PV‐180 Air Receiver 1 3,000 3,00015‐AD‐185 Air Dryer 1 5,000 5,000TOTAL Area 15 ‐ Tertiary Crushing 6,626,178

Area 17 ‐ Product Storage Loadout & Overland Conveyor17‐DC‐105 Lime Bin Vent 1 50,000 50,000 ASSUMED17‐BN‐110 Pebble Lime Storage Silo 1 60,000 60,000 380 ton17‐BX‐115 Bin Activator 1 30,000 30,000 ASSUMED17‐SJ‐120 Screw Conveyor Feeder 1 25,000 25,000 KDE 17‐CV‐170 Overland Conveyor 1 9,289,510 9,289,510 4.5Km @ 900$/lf17‐CV‐175 Crushed Ore Tripper Conveyor 1 963,357 963,357 350M @ 1,200$/lf17‐CV‐176 Portable Conveyors 24 200,000 4,800,000 Budget Price17‐CV‐177 Radial Relescopping Conveyor 1 875,000 875,000 Budget Price17‐BN‐180 Crushed Ore Surge Bin 1 213,500 213,50017‐BF‐181 Belt Feeder 1 20,000 20,000 $5000/lf KDE databse17‐BF‐182 Belt Feeder 1 20,000 20,000 $5000/lf KDE databse17‐BF‐183 Belt Feeder 1 20,000 20,000 $5000/lf KDE databse17‐BF‐184 Belt Feeder 1 20,000 20,000 $5000/lf KDE databseTOTAL Area 17 ‐ Product Storage Loadout & Overland Conveyor 16,386,367

Area 20 Heap LeachPD ‐ 2001 Pregnant Solution Pond 1 60,000 60,000PD ‐ 2002 Event Pond 1 125,000 125,000 kde 429‐02 CCE WB ‐ Pregnant solution pump


PP ‐ 2001 Pregnant Solution Pump 1 60,000 60,000 Ref A‐ 6.3Mx7.1M ‐ Pregnant solution TankPP ‐ 2002 Pregnant Solution Pump 1 60,000 60,000 kde 429‐02 CCE WB ‐ Barren solution pumpPP ‐ 2003 Pregnant Solution Pump 1 1 60,000 120,000 kde 429‐02 CCE WB ‐45x45x15 ‐ Barren Solution TankPP ‐ 2004 Event Pond Pump 1 1 40,000 80,000 kde 429‐02 CCE WB ‐PP ‐ 2005 Event Pond Pump 1 1 40,000 80,000 kde 429‐02 CCE WB ‐ Take offPD ‐ 2003 Intermediate Solution Pond 1 100,000 100,000 kde 429‐02 CCE WB ‐PP ‐ 2006 Pregnant Solution Pump 1 40,000 40,000 kde 429‐02 CCE WB ‐PP ‐ 2007 Pregnant Solution Pump 1 40,000 40,000 kde 429‐02 CCE WB ‐ MonorailPP ‐ 2008 Pregnant Solution Pump 1 40,000 40,000 kde 429‐02 CCE WB ‐‐ 6'Dia x 6'High c/w 20kG bag brkrTOTAL Area 20 Heap Leach 805,000

ADR Plant 1 5,200,000 5,200,000CC 101 Carbon Column No 1 inc 0CC 102 Carbon Column No 2 inc 0CC 103 Carbon Column No 3 inc 0CC 104 Carbon Column No 4 inc 0CC 105 Carbon Column No 5 inc 0CP 101 Control Panel CIC Columns inc 0PC 101 Carbon Transfer Pump inc 0PS 101 Carbon Adsorption Sump Pump inc 0SA 101 Wire Sampler inc 0SA 102 Wire Sampler inc 0SC 101 Trash Screen inc 0SC 102 Carbon Safety Screen inc 0SS 101 Safety Shower and Eye Wash Station inc 0TOTAL ADR Plant 5,200,000

Area 2 Acid WashFA 201 Acid Vent Fan inc 0PC 201 Acid Wash Carbon Pump inc 0PP 201 Acid Circulation Pump inc 0PP 202 Concentrated Hydrochloric Acid Transfer Pump inc 0PP 203 Neutralization Pump inc 0PS 201 Acid Area Sump Pump inc 0SS 201 Safety Shower and Eye Wash inc 0TK 201 Acid Wash Vessel inc 0TK 202 Dilute Acid Tank inc 0TK 203 Neutralization Tank inc 0TOTAL Area 2 Acid Wash 0

Area 3 Carbon StripPC 301 Stripped Carbon Pump inc 0TK 301 Strip Vessel inc 0TK 303 Blowoff Column inc 0TK 304 Blowoff Column inc 0SS 301 Safety Shower and Eye Wash inc 0TOTAL Area 3 Carbon Strip 0

Area 4 Strip Solution HandlingFL 401 Carbon Bucket Trap inc 0FL 402 Carbon Bucket Trap inc 0HE 401 Electric Immersion Heater Skid inc 0HX 401 Recovery Heat Exchanger inc 0HX 402 Trim Heat Exchanger inc 0HX 403 Trim Heat Exchanger inc 0PP 401 Strip Solution Pump inc 0PP 402 Strip Solution Pump inc 0PS 401 Strip Area Sump Pump inc 0TK 401 Barren Strip Solution Tank inc 0TOTAL Area 4 Strip Solution Handling 0

Area 5 Electrowinning and RefiningEC 501 Electrowinning Cell inc 0EC 502 Electrowinning Cell inc 0FA 501 E / W Exhaust Fan inc 0FA 502 Collector Exhaust Fan inc 0FL 501 Filter Press inc 0FL 502 Refinery Sump Filter inc 0FL 503 Furnace Dust Collector inc 0MX 501 Flux Mixer inc 0MF 501 Bullion Induction Furnace inc 0MR 501 Retort inc 0PP 501 Filter Feed Pump inc 0PS 501 Refinery Sump Pump inc 0RC 501 Rectifier inc 0RC 502 Rectifier inc 0SA 501 Wire Sampler inc 0

SA 502 Wire Sampler inc 0SP 501 High Pressure Sprayer inc 0SS 501 Safety Shower and Eye Wash inc 0EQ 501 Miscellaneous Equipment inc 0TOTAL Area 5 Electrowinning and Refining 0

Area 6 Carbon Regeneration and HandlingAG 601 Carbon Fines Tank Agitator inc 0AG 602 Carbon Attrition Agitator inc 0FA 601 Carbon Regeneration Draft Fan inc 0FE 601 Kiln Screw Feeder inc 0FL 601 Carbon Fines Filter inc 0KN 601 Carbon Reactivation Kiln inc 0PC 601 Activated Carbon Pump inc 0PP 601 Carbon Fines Pump inc 0SC 601 Kiln Dewatering Screen inc 0SC 602 Carbon Sizing Screen inc 0TK 601 Kiln Feed Bin inc 0TK 602 Carbon Quench Tank inc 0TK 603 Activated Carbon Storage Tank inc 0TK 604 Carbon Fines Tank inc 0TK 605 Carbon Attrition Tank inc 0TOTAL Area 6 Carbon Regeneration and Handling 0

Area 7 Reagent Mix / StorageAG 701 NaCN Mix Agitator inc 0BV 701 Bin Vent inc 0BV 702 Bin Vent inc 0CN 701 Crane inc 0PP 701 Caustic Mix / Transfer Pump inc 0PP 702 NaCN Transfer Mix Pump inc 0PP 703 NaCN Distribution Pump inc 0PP 704 NaCN Distribution Pump inc 0PP 705 Antiscalant Metering Pump inc 0PP 706 Antiscalant Metering Pump inc 0PS 701 Reagent Area Sump Pump inc 0SS 701 Safety Shower and Eye Wash inc 0 TK 701 Caustic Mix / Storage Tank inc 0 TK 702 NaCN Mix Tank inc 0 TK 703 NaCN Storage Tank inc 0TOTAL Area 7 Reagent Mix / Storage 0

Area 30 Solution ManagementCP‐3001 Air Compressor 1 30,000 30,000DR‐3001 Air Dryer 1 15,000 15,000FL‐3001 Barren Solution Liquid‐solids separation system 1 30,000 30,000PD ‐ 3001 Containment Pond 1 50,000 50,000PP ‐ 3001 Barren Solution Pump 1 80,000 80,000PP ‐ 3002 Barren Solution Pump 1 80,000 80,000PP ‐ 3003 Barren Solution Pump 1 80,000 80,000PP ‐ 3004 Barren Solution Pump 1 80,000 80,000PP ‐ 3013 Process Solution Pump 1 60,000 60,000PP ‐ 3014 Process Solution Pump 1 60,000 60,000PP ‐ 3006 Containment Pump 1 40,000 40,000PP ‐ 3007 Barren Solution Sump Pump 1 7,500 7,500TK ‐ 3006 Barren Solution Surge Tank 1 57,500 57,500VS ‐ 3001 Air Receiver 1 5,000 5,000TOTAL Area 30 Solution Management 675,000

Area 90 Auxiliary EquipmentGN ‐ 9001 Emergency Generator 1 25,000 25,000GN ‐ 9002 Emergency Generator 1 25,000 25,000TOTAL Area 90 Auxiliary Equipment 50,000

Project No: 439-03

Amulsar Phase II (Future Costs) Revision: P2Date: June 2, 2011


DIRECT COSTS

Process Plant :

Area 10 - Primary Crushing Phase II 2,124,143 Eqpmt. Cost x 1.43 3,037,524

Area 13 - Secondary Crushing Phase II 3,689,770 Eqpmt. Cost x 1.43 5,276,371

Area 15 - Tertiary Crushing Phase II 3,618,628 Eqpmt. Cost x 1.43 5,174,638

Area 17 - Product Storage Loadout Phase II 6,946,000 Eqpmt. Cost x 1.43 9,932,780

ADR Plant Phase II 2,000,000 2,000,000


Process Piping (1% of Plant Equip. Installed cost) 254,213

Instrumentation (1% of Plant Equip. Installed cost) 254,213

Site Development (1% of Plant Equip. Installed cost) 254,213


PLANT INDIRECT COSTSEPCM (7% of Direct Cost) 1,832,877 Construction Indirect Costs incl: (10% of Direct Cost) 2,618,395 Construction Supervision Equipment Rental Field office expense Mobilization / Demobilization Consumables Spare Parts (5% of Equipment Cost) 918,927 Equipment Insurance & Freight Cost (7% of Direct Cost) 1,832,877




TOTAL 43,403,137

ExclusionsGeotechnical Permits, royalties and licensesMining Taxes, duty and import fees; IVAReclamation and closure costs Research and development costsMetallurgical Testing Preproduction costsOre haulage & placement Local sales & Import taxesProperty acquisition cost Hazardous waste removal Permitting costs Other consultant costsSite work that is not ripable Owners home office expenseEnvironmental costs



Area 10 ‐ Primary Crushing Phase II10‐AF‐208 Apron Feeder 1 VFD ‐ 2.3m x 10m (90" x 33') 15 350,000 350,000 MCS 2010 ‐ AF 60"x8'= 467,500(add 50000 for width differnce) + 65350x add LF10‐HO‐200 Dump hopper 1 200 tonne 0 450,000 450,000 KDE Database ‐ Dump Hopper ‐ 200 Tonne (49,895kgs)10‐RB‐205 Rock Breaker 1 75 201,800 201,800 MCS 2010 ‐ 31 boom 85700 + 50 hp hammer 11610010‐SN‐210 Grizzly Feeder 1 HRVGF 64" x 24' w/VL14 Mechanism MM13011 40.2 159,618 159,618 CCE WB ‐ Metso QUOTE 2008 138,798 + ( 5% ESCLATION X 3YEARS)10‐CR‐215 Primary Crusher 1 C160 Jaw Crusher 335 800,000 800,000 Metso C160 Jaw Crusher10‐CV‐220 Primary Crusher Discharge Conveyor 1 1219mm x 15m (48" x 49') 30 49,000 49,000 $1000 x lf KDE databse10‐MA‐225 Tramp Iron Magnet 1 Suspension type 54" with rectifier 7 16,175 16,175 MCS 2010 ‐ 5KW 10‐CN‐260 Monorail Crane 1 37 ton electric monorail 20 97,550 97,550 MCS 2010 ‐ 50 TON @50' LIFT ‐ 90350 @ 30 + 360X20'TOTAL Area 10 ‐ Primary Crushing Phase II 2,124,143

Area 13 ‐ Secondary Crushing Phase II13‐AF‐112 Apron Feeder 1 VFD, 1829mm x 7m (72" x 23') 10.1 175,000 175,000 CCE WB ‐ Metso QUOTE 2008 307218 + ( 5% ESCLATION X 3YEARS)13‐AF‐113 Apron Feeder 1 VFD, 1829mm x 7m (72" x 23') 10.1 175,000 175,000 CCE WB ‐ Metso QUOTE 2008 307218 + ( 5% ESCLATION X 3YEARS)13‐CV‐220 Coarse Ore Transfer Conveyor 1 1219mm x 105m (48" x 345') 201 431,250 431,250 $1250 x lf KDE databse13‐SC‐225 Secondary Vibrating Screen 1 RiplFlo XH 2400 x 6100 DD Screen 50 65,000 65,000 MCS ‐ 6X16 INCLINED DD + MOTOR13‐CR‐230 Secondary Crusher 1 MP800 Cone Crusher 804.3 2,843,520 2,843,520 KDE databse ‐ METSO MP800 Cone Crusher ‐ budgetaryTOTAL Area 13 ‐ Secondary Crushing Phase II 3,689,770

Area 15 ‐ Tertiary Crushing Phase II15‐BF‐107 Tertiary Screening Belt Feeder 1 VFD; 1677mm x 8m (66" x 26') 10 130,000 130,000 $5000/lf KDE databse15‐BF‐108 Tertiary Screening Belt Feeder 1 VFD; 1677mm x 8m (66" x 26') 10 130,000 130,000 $5000/lf KDE databse15‐SC‐112 Tertiary Vibrating Screen 1 MF 3600 x 7300 SD Screen 74 101,800 101,800 MCS. ‐ 3.0 x 6.1m ‐ Single Deck15‐SC‐113 Tertiary Vibrating Screen 1 MF 3600 x 7300 SD Screen 74 101,800 101,800 MCS. ‐ 3.0 x 6.1m ‐ Single Deck15‐BF‐137 Tertiary Crusher Belt Feeder 1 VFD; 1219mm x 11m (54" x 36') 10 180,000 180,000 $5000/lf KDE databse15‐BF‐138 Tertiary Crusher Belt Feeder 1 VFD; 1219mm x 11m (54" x 36') 10 180,000 180,000 $5000/lf KDE databse15‐CR‐142 Tertiary Crusher 1 HP800 Cone Crusher 737.3 1,397,514 1,397,514 KDE Database ‐ Tertiary Crusher ‐ Metso HP800 Cone Crusher15‐CR‐143 Tertiary Crusher 1 HP800 Cone Crusher 737.3 1,397,514 1,397,514 KDE Database ‐ Tertiary Crusher ‐ Metso HP800 Cone CrusherTOTAL Area 15 ‐ Tertiary Crushing Phase II 3,618,628

Area 17 ‐ Product Storage Loadout Phase II17‐CV‐146 Agglomerator Feed Conveyor 1 Future 108,000 108,000 KDE Database17‐AM‐152 Agglomerator 1 Future 450,000 450,000 KDE Database17‐WX‐156 Belt Scale 1 Future 5,000 5,000 KDE Database17‐CV‐162 Agglomerator Collection Conveyor 1 Future 231,000 231,000 KDE Database17‐CV‐176 Portable Conveyors 16 Future 200,000 3,200,000 KDE Database17‐CV‐171 Extended Tripper Conveyor 1 Future 2,952,000 2,952,000 $900/LF @ 1000mTOTAL Area 17 ‐ Product Storage Loadout Phase II 6,946,000

ADR Plant Phase IICC 111 Carbon Column No 1 BCC 112 Carbon Column No 2 BCC 113 Carbon Column No 3 BCC 114 Carbon Column No 4 BCC 115 Carbon Column No 5 BCP 111 Control Panel CIC ColumnsPC 111 Carbon Transfer PumpSA 111 Wire SamplerSA 112 Wire SamplerPP 3004 Barren Solution PumpTK 302 Strip VesselEC 511 Electrowinning CellEC 512 Electrowinning CellRC 511 RectifierRC 512 RectifierTOTAL ADR Plant Phase II 1 2,000,000 2,000,000




Appendix 4

Drawings

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