For personal use only - ASX

May 2012 Agbaou Gold Project Page 1-1 NI 43-101 Technical Report

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May 2012 Agbaou Gold Project Page I NI 43-101 Technical Report

Agbaou Gold Mine NI 43-101 Technical Report Effective Date: 25 May 2012 Author’s Signatory Page

SRK Authors: Mark Wanless Principal Geologist Pr.Sci.Nat

Mark Sturgeon Principal Mining Engineer Pr Eng

Hendrik Theart Professional Geoscientist Pr.Sci.Nat

Date: 25 May, 2012

_______________ _______________

Mark Wanless Hendrik Theart Principal Geologist Professional Geoscientist

_______________ Mark Sturgeon Principal Mining Engineer

SENET Authors: Neil Senior Corporate Sponsor FSAIMM

Philemon Bundo Process Manager MSAIMM

Thilo Naidoo Studies Manager MSAIMM

John Naismith Project Sponsor

Kevin Miles Project Manager

Date: 25 May, 2012

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May 2012

Knight Piésold Authors:

Date: 25 May, 2012

Agbaou Gold Project NI 43-101

Technical Report

Duncan Grant-Stuart Director

Douglas Dorren Snr Geotech Eng

Angus Rowland Snr Hydrogeologist

Brett Garland Snr Civil Eng

Page II

Pr. Eng

Pr. Eng

Pr.Sci.Nat

Pr. Eng (O.R)

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May 2012 Agbaou Gold Project Page III NI 43-101 Technical Report

CERTIFICATE OF QUALIFIED PERSON

Mark Wanless Pr.Sci.Nat

265 Oxford Road,

Illovo,

2196,

Johannesburg, South Africa

I, Mark Wanless am a Professional Geoscientist, employed as a Principal Geologist within SRK Consulting South

Africa (Pty) Ltd.

This certificate applies to the technical report entitled Agbaou Gold Mine NI 43-101 Technical Report dated 25

May 2012.

I am a member of the South African Council for Natural Scientific Professionals, 400178/05. I graduated from the

University of Cape Town with a BSc (Hons) degree in Geochemistry in 1995.

I have practiced my profession continuously for 16 years. I have been directly involved in the Mineral Resource

estimation of the Agbaou Gold Project.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43-101

Standards of Disclosure of Mineral Projects (NI 43-101).

I did not visit the Agbaou Gold Project.

I am responsible for sections 14, 25 and 26 of the Agbaou Gold Mine NI 43-101 Technical Report.

I am independent of Endeavour Mining Corporation as independence is described by Section 1.4 of NI 43-101.

I have had no previous involvement with the Agbaou Gold Project.

I have read NI 43-101 and this report has been prepared in compliance with that Instrument.

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 the technical report not misleading.

I consent to the filing of the Technical Report with any Canadian stock exchange and other Canadian regulatory

authorities and publication by them for regulatory purposes, including electronic publication in the public company

files on their websites accessible by the public, of the Technical Report.

“Signed and sealed”

Mark Wanless, Pr.Sci.Nat

Dated at Illovo, Johannesburg, South Africa, this 10th

day of July 2012.

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Hendrik Frederik Johannes Theart Pr.Sci.Nat

265 Oxford Road,

Illovo,

2196,


I, Hendrik Frederik Johannes Theart, am a Professional Geoscientist, employed as a Corporate Consultant within

SRK Consulting South Africa (Pty) Ltd.


May 2012.

I am registered as a Professional Natural Scientist in the Geological Science field of practice, with the South

African Council for Natural Scientific Professionals, Registration Number 400069/88. I graduated from the

University of Stellenbosch with a Ph.D. degree in Geochemistry in 1985 and also have a M.Sc. (Geology) 1974

degree from the University of Cape Town and a B.Sc. (Hons) (Geology) 1977 degree from the University of

Stellenbosch.

I have practiced my profession continuously for 35 years. I have been directly involved in the review of the

exploration programme, the geological interpretation, the sampling and quality assurance and quality control of

the analytical results of the Agbaou Gold Project.



I visited the Agbaou Gold Project in Côte d’Ivoire from the 3rd

October to the 7th

October 2011.

I am responsible for sections 4 to 12, 23 and contributed to 25 and 26 of the Agbaou Gold Mine NI 43-101

Technical Report.










H. F. J. Theart, Pr.Sci.Nat


day of July 2012.

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Mark Sturgeon Pr Eng

265 Oxford Road,

Illovo,

2196,


I, Mark Sturgeon am a Professional Engineer, employed as a Principal Mining Engineer within SRK Consulting

South Africa (Pty) Ltd.


May 2012.

I am a member of the Engineering Council of South Africa 20040276. I graduated from the University of the

Witwatersrand with a BSc degree in Mining Engineering in 1975.

I have practiced my profession continuously for 36 years. I have been directly involved in the Mineral Reserve

estimation of the Agbaou Gold Project.



I visited the Agbaou Gold Project in Côte d’Ivoire from the 5th October to the 7

th October 2011.

I am responsible for sections 15, 16, 25 and 26 of the Agbaou Gold Mine NI 43-101 Technical Report.










Mark Sturgeon, Pr Eng


day of July 2012.

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Neil Senior Pr.Eng. FSAIMM

No.1 High Street, Moddercrest office Park,

Modderfontein,

1609,

Gauteng, South Africa

I, Neil senior am a Professional Engineer, employed as a Joint Managing Director of SENET.

This certificate applies to the technical report entitled Agbaou Gold Mine NI 43-101 Technical Report dated 25th

May 2012.

I am a fellow of the South African Institute of Mining and Metallurgy (SAIMM) and a registered Professional

Engineer (registration number 800284). I graduated from Cranfield University, United Kingdom with an MSc.

Engineering (mechanical) in 1972.

I have practiced my profession for 35 years. I have been directly involved as the Sponsor of the Agbaou Gold

Project and reviewer of the report on behalf of SENET.



I did not visit the Agbaou Gold Project site in Côte d’Ivoire.

I am responsible for Sections 1, 2, 3, 13, 17 to 22, 24 to 27 of the Agbaou Gold Mine NI 43-101 Technical Report.










Neil Senior, Msc Mech. Eng. FSAIMM

Dated at Modderfontein, Gauteng, South Africa, this 10th

day of July 2012.

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Duncan Grant-Stuart Pr.Eng.

4 De La Rey Road

Rivonia

2128


I, Duncan Grant-Stuart am a Professional Engineer, employed as a Director within Knight Piésold (Pty) Ltd.


May 2012.

I am a member of South African Institute of Civil Engineers and a registered Professional Engineer (registration

number 900014). I graduated from the University of the Witwatersrand in 1976 with B.Sc. Engineering (civil).

I have practiced my profession for 36 years. I have been directly involved in the management, conceptual design

and report on the Tailings Storage Facility (TSF) portion of the Engineering Optimisation Study.



I visited the Agbaou Gold Project site in Côte d’Ivoire between the 9th

October and 14th

October 2011.

I am responsible for the Tailings Storage Facility aspects of the Agbaou Gold Mine NI 43-101 Technical Report.


I have been involved with the Agbaou Gold Project in preparation of the conceptual design and report for the

Tailings Storage Facility, as part of the Engineering Optimisation Study intermittently from 2007 when the project

was run by Etruscan Resources.







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Angus Rowland, Pr.sci.Nat.

4 De La Rey Road

Rivonia

2128


I, Angus Rowland am a Professional Natural Scientist, employed as a Senior Hydrogeologist within Knight

Piésold (Pty) Ltd.


May 2012.

I am a member of South African Council for Natural Scientific Professions (SACNASP), registration number

400240/10. I graduated from the University of the Free State with an MSc in Geohydrology in 2010.

I have practiced my profession for 5 years. I have been directly involved in the management and compilation of

the hydrogeological feasibility study and report for the Agbaou Gold Project.



I visited the Agbaou Gold Project site in Côte d’Ivoire between the 9th

October and 14th

October 2011.

I am responsible for the hydrogeological aspects of the Agbaou Gold Mine NI 43-101 Technical Report.


I have been involved with the Agbaou Gold Project as part of the preliminary Hydrogeological Feasibility Study

between November 2007 and October 2008 for Etruscan Resources.







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TABLE OF CONTENTS

AUTHOR’S SIGNATORY PAGE ................................................................................ I

CERTIFICATE OF QUALIFIED PERSON ................................................................ III

TABLE OF CONTENTS ............................................................................................ IX

LIST OF TABLES ................................................................................................ XVIII

LIST OF FIGURES................................................................................................. XXI

LIST OF APPENDICES ....................................................................................... XXIII

SECTION 1. SUMMARY ................................................................................... 1-1

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

1.2 Tenure, Underlying Agreements, Permits and Rights ....................................................... 1-1

1.3 Project History ........................................................................................................................ 1-1

1.4 Geology and Mineralization .................................................................................................. 1-2

1.5 Exploration, Drilling and Sampling ...................................................................................... 1-2

1.6 Data Verification ..................................................................................................................... 1-3

1.7 Mineral Processing and Metallurgical Testing .................................................................... 1-3

1.8 Mineral Resource Estimates ................................................................................................. 1-7

1.9 Mineral Reserve Estimates and Mining Methods.............................................................. 1-11

1.10 Recovery Methods ............................................................................................................... 1-13

1.11 Environmental Studies and Social Impacts ...................................................................... 1-13

1.12 Infrastructure, Water Supply, Tailings Storage and Logistics ........................................ 1-14

1.13 Capital and Operating Costs ............................................................................................... 1-15

1.14 Economic Analysis .............................................................................................................. 1-17

1.15 Interpretation and Conclusions .......................................................................................... 1-18

1.16 Recommendations ............................................................................................................... 1-20

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SECTION 2. INTRODUCTION........................................................................... 2-1

2.1 Project Overview .................................................................................................................... 2-1

2.2 Purpose of Technical Report ................................................................................................ 2-1

2.3 Scope of Services .................................................................................................................. 2-2

2.3.1 SRK’s Scope of Services ................................................................................................. 2-2

2.3.2 SENET’s Scope of Services ............................................................................................. 2-2

2.3.3 Knight Piésold’s Scope of Services.................................................................................. 2-4

SECTION 3. RELIANCE ON OTHER EXPERTS .............................................. 3-1

SECTION 4. PROPERTY DESCRIPTION AND LOCATION............................. 4-1

4.1 Mineral Tenure ........................................................................................................................ 4-2

4.2 Underlying Agreements ......................................................................................................... 4-4

4.3 Permits and Authorization .................................................................................................... 4-5

4.4 Environmental and Social Considerations .......................................................................... 4-5

4.5 Mining Rights in Côte d'Ivoire .............................................................................................. 4-6

SECTION 5. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ......................................................... 5-1

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

5.2 Local Resources and Infrastructure .................................................................................... 5-1

5.3 Climate .................................................................................................................................... 5-1

5.4 Physiography.......................................................................................................................... 5-1

SECTION 6. HISTORY ...................................................................................... 6-1

6.1 Prior Ownership of the Property and Exploration .............................................................. 6-1

6.2 Historical Mineral Resource and Mineral Reserve Estimates ........................................... 6-1

SECTION 7. GEOLOGICAL SETTING AND MINERALIZATION ..................... 7-1

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

7.2 Property Geology ................................................................................................................... 7-3

SECTION 8. DEPOSIT TYPES.......................................................................... 8-1

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SECTION 9. EXPLORATION ............................................................................ 9-1

9.1 SRK Comments ...................................................................................................................... 9-2

SECTION 10. DRILLING ................................................................................... 10-1

10.1 Introduction .......................................................................................................................... 10-1

10.2 Drilling Procedures .............................................................................................................. 10-2

10.2.1 Drillhole Collar Location ................................................................................................. 10-2

10.2.2 Down-hole Surveying Procedures.................................................................................. 10-2

10.2.3 Reverse Circulation Procedures .................................................................................... 10-2

10.2.4 Diamond Drilling Procedures ......................................................................................... 10-2

10.2.5 RC and Core Sampling Procedures ............................................................................... 10-2

10.2.6 Drilling Orientation Procedures ...................................................................................... 10-3

10.3 Trenching and Pitting .......................................................................................................... 10-3

10.4 Drilling ................................................................................................................................... 10-3

10.5 Drilling Pattern and Density ................................................................................................ 10-3

10.6 SRK Comments .................................................................................................................... 10-3

SECTION 11. SAMPLE PREPARATION, ANALYSES AND SECURITY ......... 11-1

11.1 Sampling and Logging Procedures ................................................................................... 11-1

11.1.1 Sampling and Logging for Reverse Circulation Drilling.................................................. 11-1

11.1.2 Sampling and Logging for Diamond Drilling ................................................................... 11-1

11.2 Sample Recovery ................................................................................................................. 11-2

11.3 Sample Quality ..................................................................................................................... 11-2

11.4 Specific Gravity Data ........................................................................................................... 11-3

11.5 Quality Assurance and Quality Control Programs ........................................................... 11-4

SECTION 12. DATA VERIFICATION ................................................................ 12-1

12.1 ENDEAVOUR Field Duplicates............................................................................................ 12-1

12.2 Certified Reference Materials.............................................................................................. 12-5

12.3 Blanks .................................................................................................................................... 12-5

12.4 Umpire Laboratory Results ................................................................................................. 12-6

12.4.1 Duplicate Samples Referee Laboratory 1 ...................................................................... 12-6

12.4.2 Certified Reference Materials Submitted to SGS Ghana ............................................. 12-10

12.4.3 Duplicate Samples Referee Laboratory 2 .................................................................... 12-10

12.4.4 Certified Reference Materials Submitted to SGS Canada ........................................... 12-11

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12.5 SRK Comments .................................................................................................................. 12-15

SECTION 13. MINERAL PROCESSING AND METALLURGICAL TESTING .. 13-1

13.1 Background .......................................................................................................................... 13-1

13.2 Review of the Initial Feasibility Metallurgical Testwork ................................................... 13-2

13.3 Limitations of the Initial Feasibility Study ......................................................................... 13-4

13.4 Comminution Circuit Review .............................................................................................. 13-4

13.5 EOS Metallurgical Testwork Sample Selection ................................................................. 13-6

13.6 Metallurgical Test Program (EOS) ...................................................................................... 13-7

13.7 EOS Metallurgical Results ................................................................................................... 13-7

13.7.1 Head Analysis ................................................................................................................ 13-8

13.7.2 Gravity Recoverable Gold .............................................................................................. 13-8

13.7.3 High Shear Reactors (HSR) Testwork ........................................................................... 13-8

13.7.4 Oxygenation Testwork ................................................................................................. 13-11

13.7.5 Composite Leach Kinetic Results by Percentage Solids (w/w) ................................... 13-11

13.7.6 Preg- Robbing Testwork .............................................................................................. 13-12

13.7.7 Variability Preg-Robbing Testwork ............................................................................... 13-13

13.7.8 Thickening and Rheology Testwork ............................................................................. 13-14

13.7.9 Recommendations for Design Values .......................................................................... 13-17

13.7.10 Recommendations for Further Testwork (Post EOS Testwork) ............................... 13-17

SECTION 14. MINERAL RESOURCE ESTIMATES ......................................... 14-1

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

14.2 Resource Estimation Procedures ...................................................................................... 14-1

14.3 Resource Database .............................................................................................................. 14-2

14.4 Solid Body Modelling ........................................................................................................... 14-2

14.4.1 Primary Mineralisation .................................................................................................... 14-2

14.4.2 Surface Mineralisation .................................................................................................... 14-5

14.4.3 Weathering Zones .......................................................................................................... 14-5

14.5 Compositing ......................................................................................................................... 14-5

14.6 Evaluation of Outliers .......................................................................................................... 14-8

14.7 Statistical Analysis and Variography ............................................................................... 14-10

14.8 Block Model and Grade Estimation .................................................................................. 14-19

14.8.1 Block Size Optimization ............................................................................................... 14-19

14.8.2 Sample Number Optimization ...................................................................................... 14-23

14.8.3 Search Orientation Optimization .................................................................................. 14-29

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14.8.4 Kriging Parameters....................................................................................................... 14-32

14.8.5 Model Validation ........................................................................................................... 14-32

14.9 Recoverable Resource Modelling..................................................................................... 14-38

14.9.1 Uniform Conditioning .................................................................................................... 14-39

14.9.2 Multiple Indicator Kriging .............................................................................................. 14-46

14.10 Mineral Resource Classification ................................................................................... 14-57

14.11 Mineral Resource Statement ......................................................................................... 14-58

SECTION 15. MINERAL RESERVE ESTIMATES ............................................ 15-1

15.1 Mining Approach .................................................................................................................. 15-1

15.1.1 Resource Block Model ................................................................................................... 15-1

15.1.2 Geotechnical Investigation ............................................................................................. 15-1

15.1.3 Gap Analysis for Pit Slope Design ................................................................................. 15-2

15.2 Open Pit Optimisation ......................................................................................................... 15-2

15.2.1 Slope Angles .................................................................................................................. 15-3

15.2.2 Mining Costs ................................................................................................................... 15-3

15.2.3 Processing and General Administration Cost ................................................................ 15-3

15.2.4 Mining Factors ................................................................................................................ 15-4

15.2.5 Gold Price and Royalties ................................................................................................ 15-4

15.3 Optimisation Results ........................................................................................................... 15-4

15.3.1 Cut-off Grade Calculation ............................................................................................... 15-4

15.3.2 Optimisation Results – Mining Contractor Option .......................................................... 15-4

15.3.3 Selection of Optimum Pit Shell ....................................................................................... 15-6

15.3.4 Sensitivity Analysis ......................................................................................................... 15-6

15.4 Practical Pit Design .............................................................................................................. 15-7

15.4.1 Description ..................................................................................................................... 15-7

15.4.2 Mineral Reserves ......................................................................................................... 15-12

15.4.3 Comparison with Whittle Results ................................................................................. 15-12

15.5 Waste Dump Design .......................................................................................................... 15-12

15.6 Mine Production Schedule ................................................................................................ 15-14

15.6.1 Description ................................................................................................................... 15-14

SECTION 16. MINING METHODS .................................................................... 16-1

16.1 Introduction .......................................................................................................................... 16-1

16.1.1 Mining Equipment Requirements ................................................................................... 16-1

16.1.2 Mine Work Schedule ...................................................................................................... 16-2

16.1.3 Open Pit Dewatering ...................................................................................................... 16-3

16.1.4 Mine Infrastructure ......................................................................................................... 16-4

16.1.5 Mining Manpower ........................................................................................................... 16-6

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SECTION 17. RECOVERY METHODS ............................................................. 17-1

17.1 Process Design Criteria ....................................................................................................... 17-1

17.1.1 Introduction ..................................................................................................................... 17-1

17.1.2 Ore Characteristics......................................................................................................... 17-1

17.1.3 Operating Schedule ....................................................................................................... 17-2

17.1.4 Plant Recovery ............................................................................................................... 17-3

17.1.5 Air Services and Oxygen .............................................................................................. 17-12

17.2 Process Plant...................................................................................................................... 17-14

17.2.1 Process Plant Description ............................................................................................ 17-14

SECTION 18. PROJECT INFRASTRUCTURE ................................................. 18-1

18.1 Introduction .......................................................................................................................... 18-1

18.2 Mining Facilities ................................................................................................................... 18-3

18.2.1 Mining Administrative Building ....................................................................................... 18-3

18.2.2 Mining Equipment Workshop ......................................................................................... 18-3

18.2.3 Mining Equipment Refueling Facility .............................................................................. 18-3

18.2.4 Explosive Magazine ....................................................................................................... 18-4

18.3 Processing Plant & Administration Facilities ................................................................... 18-4

18.3.1 In-Plant Access Roads ................................................................................................... 18-4

18.3.2 Plant and Administration Buildings................................................................................. 18-4

18.3.3 Plant and Administration Warehousing .......................................................................... 18-5

18.3.4 Process Plant Site Drainage .......................................................................................... 18-6

18.3.5 Sewage Disposal............................................................................................................ 18-6

18.4 Access Road ......................................................................................................................... 18-6

18.5 Water Supply ........................................................................................................................ 18-6

18.5.1 Water Supply Dam ......................................................................................................... 18-6

18.5.2 Potable Water Distribution ........................................................................................... 18-10

18.5.3 Fire Water Distribution ................................................................................................. 18-10

18.6 Communications ................................................................................................................ 18-10

18.7 Security ............................................................................................................................... 18-10

18.8 Accommodation (Main Camp) .......................................................................................... 18-11

18.9 Power Supply (GRID) ......................................................................................................... 18-11

18.9.1 Standby Power Plant .................................................................................................... 18-12

18.9.2 Power Distribution ........................................................................................................ 18-12

18.9.3 Fuel Storage ................................................................................................................. 18-13

18.10 Tailings Storage Facility ................................................................................................ 18-13

18.10.1 Introduction and Design Criteria ............................................................................... 18-13

18.10.2 Site Selection and Access to the Site ...................................................................... 18-15

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18.10.3 TSF Construction and Wall Raising Procedure ....................................................... 18-15

18.10.4 Stability Analysis ...................................................................................................... 18-19

18.10.5 Tailings Distribution and Deposition ......................................................................... 18-20

18.10.6 Tailings Disposal Method ......................................................................................... 18-21

18.10.7 Hydrology and Water Management ......................................................................... 18-21

18.10.8 Geotechnical Conditions at the TSF Site ................................................................. 18-22

18.10.9 Water Balance .......................................................................................................... 18-22

18.10.10 Return Water System ............................................................................................... 18-22

18.10.11 Safety Classification ................................................................................................. 18-22

18.11 Country Infrastructure ................................................................................................... 18-23

18.11.1 Road ......................................................................................................................... 18-23

18.11.2 Port Facilities ............................................................................................................ 18-23

18.11.3 Rail ........................................................................................................................... 18-24

18.11.4 Air ............................................................................................................................. 18-24

18.12 Logistics and Transport ................................................................................................ 18-25

18.12.1 Logistics Transit Time .............................................................................................. 18-25

18.12.2 Routing Abidjan to Agbaou Village........................................................................... 18-25

18.12.3 Documentation ......................................................................................................... 18-26

SECTION 19. MARKET STUDIES AND CONTRACTS .................................... 19-1

SECTION 20. ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT ......................................................................................... 20-1

20.1 Introduction .......................................................................................................................... 20-1

SECTION 21. CAPITAL AND OPERATING COSTS ........................................ 21-1

21.1 Project Requirements .......................................................................................................... 21-1

21.1.1 Introduction ..................................................................................................................... 21-1

21.1.2 Scope of the Estimate .................................................................................................... 21-1

21.1.3 Responsibilities .............................................................................................................. 21-1

21.1.4 Estimate Accuracy.......................................................................................................... 21-1

21.1.5 Exclusions ...................................................................................................................... 21-1

21.1.6 Escalation ....................................................................................................................... 21-2

21.1.7 Exchange Rates ............................................................................................................. 21-2

21.1.8 Taxes, Duties and Fees ................................................................................................. 21-2

21.2 Capital Cost Estimate .......................................................................................................... 21-2

21.2.1 Capital Cost Summary ................................................................................................... 21-2

21.2.2 Mining Capital Estimate ................................................................................................. 21-4

21.2.3 Process Plant and Infrastructure .................................................................................... 21-4

21.3 Sustaining, Rehabilitation and Closure Costs ................................................................ 21-12

21.3.1 Mining Rehabilitation, Mining Contractor Demobilization and Closure Costs .............. 21-12

21.3.2 Tailings Sustaining, Rehabilitation and Closure Costs ................................................ 21-14

21.3.3 General Sustaining Capital .......................................................................................... 21-15

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21.4 Operating Cost Estimate ................................................................................................... 21-15

21.4.1 Introduction ................................................................................................................... 21-15

21.4.2 Processing Plant Operating Costs ............................................................................... 21-16

21.4.3 General and Administration Costs ............................................................................... 21-20

21.4.4 Assay Laboratory Costs ............................................................................................... 21-23

21.4.5 Refining & Royalties ..................................................................................................... 21-24

SECTION 22. ECONOMIC ANALYSIS ............................................................. 22-1

22.1 Marketing .............................................................................................................................. 22-1

22.1.1 Refining Charges, Gold Pricing & Revenue ................................................................... 22-1

22.2 Financial Analysis ................................................................................................................ 22-1

22.2.1 Evaluation Method.......................................................................................................... 22-1

22.2.2 Assumptions ................................................................................................................... 22-1

22.2.3 Financial Analysis Results ............................................................................................. 22-2

22.3 Discussion ............................................................................................................................ 22-7

SECTION 23. ADJACENT PROPERTIES ........................................................ 23-1

SECTION 24. OTHER RELEVANT DATA AND INFORMATION ..................... 24-1

24.1 Implementation ..................................................................................................................... 24-1

24.1.1 Introduction ..................................................................................................................... 24-1

24.1.2 ENDEAVOUR Execution Strategy ................................................................................. 24-1

24.1.3 EPCM ............................................................................................................................. 24-4

24.1.4 Construction ................................................................................................................... 24-5

24.1.5 Commissioning ............................................................................................................... 24-5

24.1.6 Project Schedule ............................................................................................................ 24-5

24.2 Design Criteria ...................................................................................................................... 24-7

24.2.1 Engineering Design Criteria ........................................................................................... 24-7

24.2.2 Electrical Design Criteria ................................................................................................ 24-9

SECTION 25. INTERPRETATION AND CONCLUSIONS ................................ 25-1

25.1 Mineral Resource ................................................................................................................. 25-1

25.2 Mining .................................................................................................................................... 25-3

25.3 Process Plant and Economic Evaluation .......................................................................... 25-4

SECTION 26. RECOMMENDATIONS ............................................................... 26-1

26.1 Mineral Resource ................................................................................................................. 26-1

26.2 Mining .................................................................................................................................... 26-1

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26.3 Metallurgical Testwork ........................................................................................................ 26-1

SECTION 27. REFERENCES ........................................................................... 27-1

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LIST OF TABLES Table 1.1: Summary of Initial Feasibility Metallurgical Testwork ............................................................... 1-4

Table 1.2: Summary of Head Assays and SG ........................................................................................... 1-5

Table 1.3: Summary of Gravity Recoveries ............................................................................................... 1-5

Table 1.4: Optimum Leach Parameters ..................................................................................................... 1-6

Table 1.5: Selected Design Values ........................................................................................................... 1-7

Table 1.6: Mineral Resource Statement, Agbaou Project, SRK Consulting (Pty) Ltd., 30th March 2012

using MIK at a 0.3 g/t cut-off ................................................................................................... 1-10

Table 1.7: Summary of Agbaou Mineral Reserves .................................................................................. 1-12

Table 1.8: Capital Cost Summary ........................................................................................................... 1-16

Table 1.9: Overall Operating LOM Costs ................................................................................................ 1-17

Table 1.10: Assumptions Used in the Financial Evaluation ....................................................................... 1-17

Table 1.11: Summary of Financial Analysis Results .................................................................................. 1-18

Table 4.1: Corner Co-ordinates for Proposed Agbaou Exploitation Permit Area ....................................... 4-2

Table 6.1: Goldivoire, Resource Estimate (ID2. July, 1998) ...................................................................... 6-2

Table 6.2: Goldivoire, Resource Estimate (MIK. February, 1999) ............................................................. 6-2

Table 6.3: DRD, Resource Estimate (April, 2000) ..................................................................................... 6-2

Table 6.4: Coffey Mining, Resource Estimate (February, 2008) ................................................................ 6-3

Table 10.1: Summary of Drilling and Sampling at the Agbaou Gold Project .............................................. 10-1

Table 11.1: Average Densities Applied per Weathering Zone in the Tonnage Calculation ........................ 11-4

Table 12.1: Summary Statistics for Field Duplicate Samples. .................................................................... 12-1

Table 12.2: Summary Statistics for Field Duplicate Samples ..................................................................... 12-2

Table 12.3: Summary Statistics of CRM’s Analysed by BV Laboratory ..................................................... 12-5

Table 12.4: Summary Statistics of Blank Materials Analysed by BV Laboratory ........................................ 12-5

Table 12.5: Summary Statistics for Umpire Laboratory Duplicate Samples ............................................... 12-6

Table 12.6: Descriptive Statistics of Final Dataset of Umpire Laboratory Duplicates ................................. 12-6

Table 12.7: Summary of CRM Results Submitted to Referee Laboratory 1 ............................................. 12-10

Table 12.8: Summary Statistics for Referee Laboratory 2 Duplicate Pulp Samples ................................ 12-10

Table 12.9: Descriptive Statistics of Referee Laboratory 2 Duplicate Pulp Samples (<20g/t) .................. 12-11

Table 12.10: Summary of CRM Results Submitted to Referee Laboratory 2 ............................................. 12-11

Table 13.1: Head Assays ........................................................................................................................... 13-2

Table 13.2: Comminution Results .............................................................................................................. 13-3

Table 13.3: Recovery Results .................................................................................................................... 13-4

Table 13.4: Comminution Results Summary .............................................................................................. 13-5

Table 13.5: Viscosity & Shear Rate Summary ........................................................................................... 13-5

Table 13.6: Head Assays ........................................................................................................................... 13-8

Table 13.7: Gravity Gold Recovery ............................................................................................................ 13-8

Table 13.8: Summary Results: Effect of using a HSR ............................................................................... 13-9

Table 13.9: Effect of Sparging Air / Oxygen on Gold Extraction .............................................................. 13-11

Table 13.10: Saprolite Leach Kinetics at Varied Percent Solids. ............................................................... 13-12

Table 13.11: Bedrock Leach Kinetics at Varied Percent Solids ................................................................. 13-12

Table 13.12: Effect of Preg-Robbing .......................................................................................................... 13-13

Table 13.13: Variability Preg-Robbing Results Summary .......................................................................... 13-14

Table 13.14: Thickening and Rheology Results......................................................................................... 13-14

Table 13.15: Variability Leach Results on Middles and Tails ..................................................................... 13-16

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Table 13.16: Optimum Leach Parameters ................................................................................................. 13-17

Table 13.17: Selected Design Values ........................................................................................................ 13-17

Table 14.1: Gold Grade and Sample Length Statistics for Samples Within Mineralised Zones ................. 14-6

Table 14.2: Statistics of Standard and Non-Standard Length Composites ................................................ 14-8

Table 14.3: Statistics of Composites for Gold by Mineralised Zone ........................................................... 14-8

Table 14.4: Summary of Capping Applied in Ordinary Kriged Estimate Per Zone ................................... 14-10

Table 14.5: Statistics of Gaussian Gold Variables per Zone .................................................................... 14-16

Table 14.6: Modelled Semi-Variogram Parameters per Zone .................................................................. 14-17

Table 14.7: Block Size Optimisation Scenarios ....................................................................................... 14-20

Table 14.8: Summary of Search Neighbourhood Parameters ................................................................. 14-31

Table 14.9: Samples and Composites Versus Kriged Means Per Zone .................................................. 14-33

Table 14.10: Declustering Statistics for Gaussian Anamorphosis Modelling ............................................. 14-39

Table 14.11: Statistics of the Gaussian Variables ...................................................................................... 14-40

Table 14.12: Support Correction Parameters ............................................................................................ 14-40

Table 14.13: Comparison of the OK and UC Estimation Results at a 0.5 g/t Cut-Off ................................ 14-46

Table 14.14: Comparison of the OK and UC Estimation Results at a 1.0 g/t Cut-Off ................................ 14-46

Table 14.15: Zone 1 Semi-Variogram Model Parameters .......................................................................... 14-51



Table 14.18: Volume Variance Ratios for Each MIK Zone ......................................................................... 14-54

Table 14.19: Statistics of OK Estimates and MIK e-type Means ................................................................ 14-55

Table 14.20: Assumptions Considered for Conceptual Open Pit Optimization .......................................... 14-59

Table 14.21: Mineral Resource Statement, Agbaou Project, SRK Consulting (Pty) Ltd., 30th March 2012 using

MIK at a 0.3 g/t cut-off ........................................................................................................... 14-60

Table 14.22: Mineral Resource Reported at Incremental Cut-off Grades, Agbaou Project, SRK Consulting

(Pty) Ltd, 30th

March 2012 using MIK .................................................................................... 14-61

Table 15.1: Mineral Resource Input for Optimisation (Using a 0.5g/t Au cut-off). ...................................... 15-1

Table 15.2: Processing Operating Cost for the Open Pit Optimization ...................................................... 15-3

Table 15.3: Whittle Optimisation Results (Contractor Option) .................................................................... 15-5

Table 15.4: Summary of Agbaou Mineral Reserves ................................................................................ 15-12

Table 15.5: Comparison of Practical Pit with Whittle Shell....................................................................... 15-12

Table 15.6: Details of Waste Dumps ....................................................................................................... 15-13

Table 15.7: Waste Dumping Volumes ..................................................................................................... 15-14

Table 16.1: Equipment Selection for Mine Operation by Contractor .......................................................... 16-2

Table 16.2: Scheduled Working Periods .................................................................................................... 16-3

Table 17.1: Design Criteria – Ore Characteristics ...................................................................................... 17-2

Table 17.2: Design Criteria – Operating Schedule ..................................................................................... 17-2

Table 17.3: Design Criteria – Recovery ..................................................................................................... 17-3

Table 17.4: Design Criteria – Crushing and Ore Stockpiling ...................................................................... 17-4

Table 17.5: Design Criteria – Milling .......................................................................................................... 17-5

Table 17.6: Design Criteria – Classification and Trash Handling ............................................................... 17-6

Table 17.7: Design Criteria – Gravity ......................................................................................................... 17-6

Table 17.8: Design Criteria – CIL ............................................................................................................... 17-7

Table 17.9: Design Criteria – Cyanide Detoxification ................................................................................. 17-8

Table 17.10: Design Criteria – Acid Wash, Elution and Regeneration ......................................................... 17-9

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Table 17.11: Design Criteria – Electrowinning and Smelting ..................................................................... 17-10

Table 17.12: Design Criteria – Lime .......................................................................................................... 17-10

Table 17.13: Design Criteria – Sodium Cyanide ........................................................................................ 17-11

Table 17.14 Design Criteria – Caustic ...................................................................................................... 17-11

Table 17.15: Design Criteria – Sodium Metabisulphite .............................................................................. 17-11

Table 17.16: Design Criteria – Copper Sulphate ....................................................................................... 17-11

Table 17.17: Design Criteria – Hydrochloric Acid ...................................................................................... 17-12

Table 17.18: Design Criteria – Activated Carbon ....................................................................................... 17-12

Table 17.19: Design Criteria – Water Services .......................................................................................... 17-13

Table 18.1: Water Storage Dams – Catchment and Storage Capacities ................................................... 18-7

Table 18.2: Design Criteria – Tailings Storage Facility ............................................................................ 18-14

Table 18.3: TSF Available Capacity per Ore Feed Type ......................................................................... 18-16

Table 18.4: TSF Construction Sequencing .............................................................................................. 18-16

Table 18.5: Material Properties ................................................................................................................ 18-19

Table 18.6: Definitions of the Analyses Scenarios ................................................................................... 18-19

Table 18.7: Results of Analysis ................................................................................................................ 18-20

Table 18.8: Abidjan Port Facilities / Storage Capacities .......................................................................... 18-24

Table 21.1: Exchange Rates ...................................................................................................................... 21-2

Table 21.2: Capital Cost Summary ............................................................................................................ 21-3

Table 21.3: Summarised Construction Quantities ...................................................................................... 21-5

Table 21.4: Structural Steel and Platework Masses .................................................................................. 21-6

Table 21.5: Typical Shipping Costs of Container Types ............................................................................ 21-7

Table 21.6 : Cost Breakdown Summary ..................................................................................................... 21-8

Table 21.7: Cost of Drilling, Pump Installation and Aquifer Testing at the Agbaou Camp ......................... 21-8

Table 21.8: Water Supply and Storage Capital Cost Summary ................................................................. 21-9

Table 21.9: Mining Rehabilitation and Closure Costs .............................................................................. 21-13

Table 21.10: Rehabilitation and Closures Costs ........................................................................................ 21-15

Table 21.11: Overall Operating LOM Costs ............................................................................................... 21-15

Table 21.12: Overall Mining LOM Operating Cost ..................................................................................... 21-16

Table 21.13: Overall Process Plant LOM Operating Costs ........................................................................ 21-16

Table 21.14: LOM Reagents and Consumables Costs .............................................................................. 21-18

Table 21.15: LOM Power Costs ................................................................................................................. 21-19

Table 21.16: LOM Tailings Operating Costs .............................................................................................. 21-20

Table 21.17: LOM G&A Costs ................................................................................................................... 21-21

Table 21.18: LOM Assay Cost Summary ................................................................................................... 21-23

Table 21.19: Refining and Royalty Cost Summary .................................................................................... 21-24

Table 22.1: Assumptions Used in the Financial Evaluation ....................................................................... 22-2

Table 22.2: Summary of LOM Production, Capital and Operating Costs ................................................... 22-3

Table 22.3: Summary of Financial Analysis Results .................................................................................. 22-3

Table 22.4: LOM Project Cashflow ............................................................................................................ 22-5

Table 22.5: Gold Price Sensitivity .............................................................................................................. 22-6

Table 22.6: Capex Sensitivity .................................................................................................................... 22-6

Table 22.7: Operating Costs Sensitivity ..................................................................................................... 22-6

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LIST OF FIGURES Figure 4.1: Map Showing the Location of the Agbaou Project .................................................................... 4-1

Figure 4.2: Mineral Tenure ......................................................................................................................... 4-3

Figure 5.1: Topography of the Project Area ................................................................................................ 5-2

Figure 7.1: Regional Geology ..................................................................................................................... 7-2

Figure 7.2: Local Geology .......................................................................................................................... 7-4

Figure 8.1: Mottled Quartz Vein in Drill Core in Fresh Rock ....................................................................... 8-1

Figure 8.2: Results of the Down-hole Sereoscopic Structural Analysis ...................................................... 8-2

Figure 8.3: Chipboard from RC Drilling Intersection of a Mineralized Quartz Vein ..................................... 8-3

Figure 8.4: Mineralized Quartz Vein Intersected by Diamond Drilling in Saprolite ...................................... 8-4

Figure 9.1: Distribution of Gold in Soil Samples ......................................................................................... 9-2

Figure 10.1: Distribution of Drilling by BHP and Goldivoire ......................................................................... 10-4

Figure 10.2: Distribution of Drilling by ERCI (2005 to 2008) ....................................................................... 10-5

Figure 10.3: Distribution of Drilling by ENDEAVOUR (2010 to 2011) ......................................................... 10-6

Figure 10.4: Final Distribution of Drilling Relative to the Designed Main Pit Outline ................................... 10-7

Figure 12.1: Scatter Plot of Field Duplicates in Reduced Dataset ............................................................... 12-3

Figure 12.2: H.A.R.D Plot of Field Duplicates in Reduced Dataset ............................................................. 12-4

Figure 12.3: Scatter Plot of Umpire Laboratory Duplicate Analyses in the Reduced Dataset ..................... 12-7

Figure 12.4: H.A.R.D Plot of Umpire Laboratory Duplicate Analyses in the Reduced Dataset ................... 12-8

Figure 12.5: QQ Plot of Umpire Laboratory Duplicate Analyses in the Reduced Dataset ........................... 12-9

Figure 12.6: Scatter Plot of the Original Analyses Versus the Results of Referee Laboratory 2 ............... 12-12

Figure 12.7: H.A.R.D Plot of the Original Analyses Versus the Results of Referee Laboratory 2 ............. 12-13

Figure 12.8: QQ Plot of Original and Referee Laboratory 2 Results (<20 g/t Au) ..................................... 12-14

Figure 13.1: EOS Testwork Flowsheet ....................................................................................................... 13-7

Figure 13.2: Bedrock Rate of Gold Dissolution With and Without High Shear Reactor ............................. 13-10

Figure 13.3: Saprolite Rate of Gold Dissolution With and Without High Shear Reactor ............................ 13-10

Figure 13.4: Slurry Yield Stress vs. Percent Solids Curves....................................................................... 13-15

Figure 14.1: Primary Mineralisation Wireframe Models .............................................................................. 14-4

Figure 14.2: Zone 1 Capping Analysis Charts ............................................................................................. 14-9

Figure 14.3: Histograms of Gold Grade for the Zone 1 Group of Zones ................................................... 14-12

Figure 14.4: Histograms of Gold Grade for the Zone 2 Group of Zones ................................................... 14-13

Figure 14.5: Histograms of Gold Grade for the Individual Zones .............................................................. 14-14

Figure 14.6: Histograms of Gold Grade for the Western Area Zones ....................................................... 14-15

Figure 14.7: Gaussian Semi-variograms for Modelled Zones ................................................................... 14-18

Figure 14.8: QKNA Results for Block Size Selection on Zone 1 ............................................................... 14-21

Figure 14.9: QKNA Results for Block Size Selection on Zone 2 ............................................................... 14-22

Figure 14.10: QKNA Results for Sample Number Selection on Zone 1 ...................................................... 14-25

Figure 14.11: QKNA Results for Sample Number Selection on Zone 2 ...................................................... 14-26

Figure 14.12: QKNA Results for Sample Number Selection on Zone 3 and 9 ............................................ 14-27

Figure 14.13: QKNA Results for Sample Number Selection on Zone 11 and 51 ........................................ 14-28

Figure 14.14: Search Domains Main, Main/Central, South/Central and South for Zones 1 to 10 ............... 14-30

Figure 14.15: Search Domains West for Zones 51 to 53 ............................................................................ 14-30

Figure 14.16: Swath Plots of the Main and South Areas for Zones 1 – 10 .................................................. 14-35

Figure 14.17: Swath Plots of the Main and South Areas for Zone 1 ........................................................... 14-36

Figure 14.18: Swath Plots of the Main and South Areas for Zone 2 ........................................................... 14-37

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Figure 14.19: Swath Plots of the Laterite (Zone 11) .................................................................................... 14-38

Figure 14.20: Grade Tonnage Curves for Zone 1 and Zone 2 .................................................................... 14-42



Figure 14.23: Zone 1 Indicator Semi-Variograms for Indicators up to 3 g/t ................................................. 14-48



Figure 14.26: Grade Tonnage Curves Comparison Between MIK and UC for Zones 1 and 2 .................... 14-56

Figure 15.1: Whittle Pit by Pit Graph Optimisation Results (Contractor Option).......................................... 15-5

Figure 15.2: Optimised Pit Shell Selected for Practical Pit Design .............................................................. 15-6

Figure 15.3: Sensitivity Analysis Graph ...................................................................................................... 15-7

Figure 15.4: Haul Road Widths ................................................................................................................... 15-8

Figure 15.5: View of the Agbaou Intermediate Pits ..................................................................................... 15-9

Figure 15.6: View of the Agbaou Final Pits ............................................................................................... 15-10

Figure 15.7: View of the Agbaou Final Pits Showing the Plant Area and the Waste Dump ...................... 15-11

Figure 15.8: Vertical Sections Through Waste Dumps ............................................................................. 15-13

Figure 16.1: Layout of Surface Access Roads, Explosive Magazine and ROM Stockpile Location ............ 16-5

Figure 17.1: Process Flow Diagram .......................................................................................................... 17-14

Figure 17.2: Plant General Layout ............................................................................................................ 17-15

Figure 18.1: Site General Layout ................................................................................................................ 18-2

Figure 18.2: Water Storage Dams – Options 1 to 5 .................................................................................... 18-8

Figure 18.3: Water Storage Dam Option 3 Capacity Curves....................................................................... 18-9

Figure 18.4: TSF Phase 1 ......................................................................................................................... 18-18

Figure 18.5: Map of Country ..................................................................................................................... 18-23

Figure 18.6: Abidjan Port Map .................................................................................................................. 18-24

Figure 18.7: Logistics Transit Time Summary ........................................................................................... 18-25

Figure 18.8: Map Showing Travel Route Options to Agbaou .................................................................... 18-25

Figure 22.1: NPV Sensitivity at 5% Discount Rate ...................................................................................... 22-7

Figure 22.2: IRR Sensitivity at 5% Discount Rate ....................................................................................... 22-7

Figure 24.1: Owner’s Project Implementation Team ................................................................................... 24-3

Figure 24.2: Project Execution Schedule .................................................................................................... 24-6

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LIST OF APPENDICES Appendix 3-A: Opinion from Theodore Heogan and Michel Ette Advocates Associes.

Appendix 4-A: Decree and Exploitation License

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SECTION 1. SUMMARY

1.1 INTRODUCTION Endeavour Mining (ENDEAVOUR) commissioned the following entities to review the

following aspects and optimize the previous feasibility study conducted in September 2009:

• SRK Consulting (South Africa) – Mineral Resource, Mineral Reserve and Mining; • SENET (South Africa) – Process Plant, Infrastructure, Economic Evaluation, and

overall study management; • Knight Piésold (South Africa) – Tailings Management Facility and Water Storage

Dam; • Metallurgical Testwork Consultants – Mintek (South Africa), Maelgwyn Mineral

Services (South Africa), Paterson & Cooke (South Africa), Knelson Africa (South Africa);

• Orway Mineral Consultants (Australia) – Comminution Circuit Design.

The purpose of this study is to demonstrate the viability of the Agbaou Project through:

• Updated mineral resources and reserves; • Economic evaluation taking into account improved resources/reserves and updated

capital and operating costs.

This technical report is a complete summary of the full study as prepared by SRK

Consulting, SENET, Knight Piésold Consulting (2012).

The Agbaou Gold Project is located in Côte d’Ivoire, West Africa on a property 200km north-

west of Abidjan. The concession is reached by tarred and secondary gravel roads and within

10km from the national electrical power grid. The small town of Agbahou is located 12km

north-east of the regional town of Didoko.

1.2 TENURE, UNDERLYING AGREEMENTS, PERMITS AND RIGHTS The mineral right to the project, in the form of a renewable exploration permit, is registered in

the name of Etruscan Resources Côte d’Ivoire SARL, a fully owned subsidiary of

ENDEAVOUR. The current exploration permit expired on 22nd March 2007, but in terms of a

legal opinion (19th August 2011) the permit remains valid since renewal applications have

been submitted and recognized on 22nd March 2007 and 6th October 2009 and with a

Modified Map submitted in October 2009. A ministerial approval to commence with the

development of the mine was given on 16th August 2010.

1.3 PROJECT HISTORY Alluvial gold has been known for some time in this area. Gold mineralisation in bedrock was

first reported in the Agbaou area during the late 1980’s, while the ground was held by a joint-

venture between BHP Minerals and SODEMI. The joint-venture conducted exploration

between 1988 and 1994

Between 1996 and 2000 the property was held by Goldivoire S.A.R.L., controlled by Jandera

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Resources NL (88.4%), a wholly owned subsidiary of Diversified Mineral Resources NL

which was taken over by Hargraves Resources NL in mid-1999. Hargraves Resources were

in turn taken over by Durban Roodepoort Deeps of South Africa in December 1999.

Goldivoire undertook exploration during the period 1996 to 2000. The government of Côte

d’Ivoire withdrew the Agbaou permit and on 27th November 2003, the Ministry of Mines and

Energy for Côte d’Ivoire subsequently granted the Agbaou exploration permit to Etruscan

Resources Côte d’Ivoire (Etruscan) as (Decree Number 2003-464).

After obtaining the project in 2003, Etruscan drilled an additional 179 boreholes and

conducted various studies. The combination of this information with the historic information

formed the basis for a Feasibility Study in 2009. Following this, Etruscan continued with a

sterilization and infill drilling programme from 2010 to 2011 by drilling an additional 85 holes

(7063 m), which required the redefinition of the Mineral Resources.

1.4 GEOLOGY AND MINERALIZATION The shear-zone hosted gold mineralization at the Agbaou deposit is within a sheared

volcano-sedimentary succession that was subjected to lower green-schist facies

metamorphism, forming the Birimian age Oumé Fetekro Greenstone Belt, surrounded by

granodioritic intrusions.

Gold occurs in a mesothermal auriferous sulphide (pyrite + pyrrhotite) assemblage

associated with quartz veins. The quartz veins are characterized by a wide range of quartz-

vein types, brecciation, boudinage, sericitic and carbonate alteration.

The mineralised quartz veins have a very distinctive texture that has been described as

“mottled”. These veins are easily identifiable in the diamond drilling core intersections from

the fresh rock.

1.5 EXPLORATION, DRILLING AND SAMPLING Etruscan’s evaluation of the Agbaou project area began in 2003 with the acquisition of the

Agbaou permit. Exploration has been carried out under the supervision of technically

qualified personnel applying standard industry approaches. Geochemical data quality has

routinely been assessed as part of on-going exploration procedures. All data acquired meets

or exceeds industry standards. All exploration work has been carried out by, or supervised

by technical personnel of the operator. Consultants and contractors have been engaged by

Etruscan for various activities including; sample collection, drilling and assaying.

Etruscan conducted detailed and regional soil geochemical surveys which identified the gold

mineralization at areas known as Agbaou, Agbaou South, Mbazo, Zehiri and Niafouta.

A total of 876 pits and 4 trenches were dug to explore the laterite resource but these results

were not used in the resource estimation.

A total of 514 holes (diamond and reverse circulation) were drilled at the Agbaou Project,

totalling 58,444m.

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Only limited sample preparation was done on site and this pertains mainly to the splitting by

diamond saw of the core samples and the splitting of the percussion drilling chips with two

way riffle splitters. All crushing and milling was completed by independent commercial

laboratories and following standard industry practice. The samples of the last campaign were

submitted to the Bureau Veritas Mineral Laboratory Côte d’Ivoire (BV Laboratory), in Abidjan

for gold analyses using the fire assay method with an atomic absorption finish. An auditable

chain of custody was established for the sample handling, data reporting and database

capturing.

1.6 DATA VERIFICATION The reliability of the gold assay results was based on a well designed and implemented

quality assurance and quality control protocol that entail the analysis of blind blanks,

duplicates and certified reference materials. In addition, samples were also submitted to

umpire laboratories. The apparent coarse nature of the gold results in a high variability in the

field duplicate set. The laboratory returned very good results for the certified reference

materials. Similarly blank material returned acceptable results and SRK accepts the BV

Laboratory results.

The variation in results of the duplicate pulp samples submitted to the SGS laboratory in

Ghana and to the accredited SGS Laboratory in Canada indicates poor but acceptable

replication at the Umpire Laboratories.

SRK believes that the current quality systems in place at Agbaou to monitor the precision

and accuracy of the sampling and assaying, is adequate and that the laboratory returned

acceptable results for use in resource estimation.

1.7 MINERAL PROCESSING AND METALLURGICAL TESTING

Background In 2009, comminution and recovery metallurgical testwork was performed by Mintek in South

Africa on ore samples from the Agbaou deposits. The results were used in the initial

feasibility study to develop the process flowsheet.

In 2011 SENET reviewed the results from the initial feasibility study and with ENDEAVOUR

proposed a metallurgical test program whose results would support the process flowsheet

selection required for an Engineering Optimization Study (EOS).

Review of the Initial Feasibility Metallurgical Testwork The table below includes a summary of the initial feasibility metallurgical results generated

by Mintek in 2009.

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Table 1.1: Summary of Initial Feasibility Metallurgical Testwork

Saprolite Bedrock

Units South North South North

HEAD ASSAY RESULTS

Ore grade g/t 1.88 1.69 1.09 3.67

Moisture content % 5 5 5 5

Specific gravity t/m3 1.69 1.69 2.79 2.83

Bulk density t/m3 1.01 1.01 1.67 1.7

COMMINUTION RESULTS

UCS-Average Mpa - - 102.9 100.5

BBWi(106µm)-Max kWh/t - - 12.54 15.22

BRWi-Max kWh/t - - 16.48 19.22

Ai-Max kWh/t - - 0.1729 0.1851

JK

A × B - - 23.3 22.9

Ta - - 0.40 0.46

RECOVERY RESULTS

Gravity recoverable gold % 23.2 20 22.7 31.4

CIL dissolution (no oxygen) % 90 83.8 90.8 90.6

CIL dissolution (with oxygen) % 92.2 93.1 - -

Overall gold recovery % 94.01 94.48 92.89 93.55

Cyanide consumptions

No oxygen kg/t 0.22 0.18 0.23 0.26

With oxygen kg/t 0.24 0.18 - -

Lime consumptions

No oxygen kg/t 3.01 2.97 0.52 0.53

With oxygen kg/t 2.99 2.97 - -

Comminution Circuit Review In 2011, OMC was requested to review the comminution results produced by Mintek during

the initial feasibility study and make recommendations for a suitable comminution circuit to

treat the Agbaou ore.

From OMC analysis, ENDEAVOUR selected a SAG milling, ball milling and pebble crushing

(SABC) circuit for treating the Agbaou ore based on the following:

• Power is relatively inexpensive in Côte d’Ivoire;

• The SABC circuit has a simple mode of operation, with advantages on operability

and maintenance.

Aim of EOS Metallurgical Testwork Following a review of the initial feasibility metallurgical test results, it was found that there

were limitations to this testwork therefore SENET and ENDEAVOUR proposed a test

program for the EOS. The aims of this testwork program were to:

• Look at areas where the plant design can be improved; • Conduct testwork to address the shortfalls of the initial feasibility study;

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• Complete testwork required to satisfy a final feasibility study.

Saprolite and bedrock samples from the north and south deposit were received for the EOS

metallurgical testwork.

EOS Testwork Results

Head Assays and Specific Gravity (SG) Full elemental analysis and SG determination was conducted on the saprolite and bedrock

composites. The table below shows a summary of the head assays and SG.

Table 1.2: Summary of Head Assays and SG

Testwork Units Saprolite Bedrock

HEAD ASSAYS

Au-composite g/t 2.36 2.99

Au-variability g/t 2.61 2.08

SG-composite

2.79 2.85

SG-variability

2.80 2.89

Gravity Recoverable Gold GRG tests were conducted and plant GRG recoveries were predicted as shown in the table

below.

Table 1.3: Summary of Gravity Recoveries

Testwork Units Saprolite Bedrock

GRAVITY RECOVERY

GRG (lab) % 44.1 31.1

Mass pull % 0.73 0.85

Conc. grade g/t Au 127.1 90.7

GRG (plant predicted) % 30 20

High Shear Reactor (HSR) and Oxygenation Testwork HSR testwork was conducted to evaluate whether a High Shear Reactor could be used to

lower residence time, improve gold recovery and reduce reagent consumptions.

The results indicated that the HSR does lower cyanide consumptions and improve leach

kinetics but does not improve overall gold extraction. When air and oxygen sparging was

investigated to access the effect on gold recovery in was found that oxygen sparging

produced the best gold extraction. Oxygen sparging was therefore selected as part of the

plant design.

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Composite Leach Kinetic Results by Percentage Solids (w/w) Saprolite and bedrock leach kinetic tests were performed at time intervals between 2hrs and

48hrs at varying percentage solids (w/w) to assess the leach kinetics of saprolite and

bedrock ores types.

The testwork indicated the following optimum leach conditions for saprolite and bedrock:

• Saprolite: 24hrs leach time at 40% solids;

• Bedrock: 32hrs leach time at 42% solids. (42% was selected for design

purposes as it is easily achievable with densifying cyclones)

Preg-Robbing Rate and Preg-Robbing Variability Tests Preg-robbing rate test were performed on the composite saprolite and composite bedrock

samples. The results indicated that the preg-robbers are most active in the initial stages of

leach. 28.12% and 15.92% gold was lost to preg-robbers in the first 2hrs of leach for the

saprolite and bedrock ore respectively.

Since the pre-robbing test on the composite sample showed that both ore types have high

preg-robbing characteristic, variability preg-robbing tests were conducted to see if the preg-

robbing issue occurred throughout the ore body. All the variability samples tested showed

the presence of preg-robbers.

Variability Leach Results on Middlings and Tails Variability leach tests were performed at optimum conditions to determine the dissolutions

and reagent consumptions. This was then compared to the optimum results to determine the

ores variability and it was observed that the ore was not highly variable.

Optimum Leach Parameters Selected leach parameters were varied to obtain optimum conditions for the design of the

leach circuit, including the following:

• Effect of cyanide addition.

• Effect of oxygen and air sparging

• Effect of percentage solids.

The table below is a summary of the optimum leach parameters selected.

Table 1.4: Optimum Leach Parameters

Optimum Leach Conditions

Saprolite Bedrock

Cyanide Addition 0.7kg/t 0.7kg/t

Residence Time 24hrs with oxygen sparging 32hrs with oxygen sparging

pH 10.50 10.50

Solids ( m/m) 40% 42%

Gravity Recovery (Lab) 44.1% 31.1%

Au Dissolution (CIL) (Midds and Tails) 97.14% 92.86

Overall Lab Au Recovery - Gravity & CIL 98.40% 95.08%

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Thickening, Rheology and Viscosity Testwork Thickening testwork conducted on the saprolite and bedrock ore indicated that both post

leach slurries can be effectively thickened.

During the initial feasibility study it was observed on Saprolite ore that viscosity issues may

be encountered. In order to possibly increase cyclone overflow densities to higher values,

viscosity modifier testwork were conducted during this phase. The results showed that

viscosity modifiers did reduce the viscosity of the saprolite slurry.

Recommendations for Design Values Design values were selected from the testwork results for the treatment of the Agbaou ore

and are summarized below.

Table 1.5: Selected Design Values Units Saprolite Bedrock

Grind

80%-75µm 80%-75µm

Selected Milling Circuit

SABC SABC

Solids SG

2.79 2.82

Final Residue (Gravity-CIL) g/t Au 0.10 0.16

Cyanide Consumption kg/t 0.33 0.18

Lime Consumption kg/t 3.23 1.50

Leach Time Hrs. 24 32

Oxygen Sparge During Leach

yes yes

CIL Percent Solids % m/m 40 42

Thickening Flux Rate t/(m2.h) 0.2 0.3

1.8 MINERAL RESOURCE ESTIMATES SRK generated a new resource estimate as an update of the previous estimate by Coffey

Mining, completed in 2008, including additional drilling data acquired since then and revised

interpretations. SRK compiled a new database from the source information supplied by

ENDEAVOUR, and generated a new drill hole database independent from that used by

Coffey. SRK used the Coffey wireframes as a guide to the original interpretation of the

mineralized zones. The SRK interpretation, however, honors all the additional drilling

information, aimed to select a higher grade zone within the drill holes (i.e. is less tolerant of

internal dilution), and in places revised the interpretation of the position and continuity of

certain zones based on new information and matching of the gold tenor in the intersections.

SRK conducted a composite optimisation to select the 2m composite length used for all

data. From the 2m composites SRK modelled semi-variograms for six of the total of fourteen

zones, where there was sufficient data to obtain a reasonably structured semi-variogram.

The zones with insufficient data to generate robust semi-variograms borrowed semi-

variograms from zones with similar grade characteristics. SRK completed a Quantitative

Kriging Neighbourhood analysis to select the optimal search and estimation parameters for

each zone.

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An initial Ordinary Kriged estimate was done, and followed by recoverable resource

estimation using two techniques. SRK were requested to use Uniform Conditioning (“UC”)

and Multiple Indicator Kriging (“MIK”) estimates to model the recoverable resources above a

range of cut-off values. The UC estimate was completed using a global change of support

based on the Discrete Gaussian Model and using the Information Effect. Only zones with

sufficient data and their own semi-variograms were subject to the UC.

For the same zones, SRK conducted a MIK estimate, defining semi-variograms for a range

of indicator cut-off’s, and then Kriging the indicators using the same search parameters as

used in the original Ordinary Kriged estimate. The Kriging of the indicators generated

conditional probabilities for a restricted set of specified cut-offs. In order to transform the

probability estimates into grade and tonnage proportions, the local conditional cumulative

density function (ccfd) is rebuilt for each block, using the Indirect Correction through

Permanence of Lognormal Distribution approach. To be consistent with the previous Coffey

MIK estimate no Information Effect was applied in the support correction, however the same

Selective Mining Unit (SMU) size was used in both the MIK and UC estimates (2.5m cubes).

The Ordinary Kriged, UC and MIK models were compared against each other, and also

validated against the source data at zero cut-off. The OK and UC estimates showed good

agreement both globally and locally when compared to the source data. The MIK results

showed very slightly elevated grades at zero cut-off for some zones, but showed good

agreement when compared to the Ordinary Kriged and UC estimates at zero cut-off.

The mineral resources have been classified considering the quality of data, including drilling,

logging, sampling and analytical quality. The slope of regression was used to measure the

quality of the estimates, and blocks with a slope of regression of greater than 0.7, typically in

the densely drilled areas the resource is classified in the measured category. The blocks

estimated in the first search pass, which approximated the semi-variogram range, and

satisfies the search parameter requirements in terms of number of samples, has been

classified in the indicated category. The blocks estimated in the second and third search

passes are classified in the inferred category.

CIM Definition Standards for Mineral Resources and Mineral Reserves (27th November

2010) defines a mineral resource as:

“A Mineral Resource is a concentration or occurrence of diamonds, natural solid inorganic

material, or natural solid fossilized organic material including base and precious metals, coal,

and industrial minerals in or on the Earth’s crust in such form and quantity and of such a

grade or quality that it has reasonable prospects for economic extraction. The location,

quantity, grade, geological characteristics and continuity of a Mineral Resource are known,

estimated or interpreted from specific geological evidence and knowledge.”

The “reasonable prospects for eventual economic extraction” requirement generally implies

that the quantity and grade estimates meet certain economic thresholds and that the mineral

resources are reported at an appropriate cut-off grade taking into account extraction

scenarios and processing recoveries.

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In order to determine the quantities of material offering “reasonable prospects for economic

extraction” by an open pit, SRK used a pit optimizer and reasonable mining assumptions to

evaluate the proportions of the block model (Measured, Indicated and Inferred blocks) that

could be “reasonably expected” to be mined from an open pit.

The optimization parameters were selected based on the parameters defined for the reserve

pit optimisation using contractor mining, aside from the gold price, which was increased from

the US$1,200/t used for the reserve open pit to an intentionally optimistic US$2,000/oz. The

reader is cautioned that the results from this pit optimization are used solely for the purpose

of testing the “reasonable prospects for economic extraction” by an open pit and do not

represent an attempt to estimate mineral reserves. The results were used as a guide to

assist in the preparation of a mineral resource statement and to select an appropriate

resource reporting cut-off grade. From these parameters, a cut-off value of 0.3 g/t was

calculated, and this has been applied in the reporting of the mineral resources. The MIK

mineral resource estimate is summarised in and a range of cut-off grades in tables below.

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Table 1.6: Mineral Resource Statement, Agbaou Project, SRK Consulting (Pty) Ltd., 30th March 2012 using MIK at a 0.3 g/t cut-off

Classification Zone kt Grade (Au (g/t) kozs

Measured 1 1,681 3.60 194

2 870 2.26 63

3 62 2.04 4

4 287 1.44 13

9 687 1.54 34

11 2,989 0.86 83

51 1,384 1.53 68

Total Measured 7,959 1.80 460

Indicated 1 2,318 2.71 202

2 2,369 2.20 168

3 504 1.51 25

4 882 1.24 35

5 617 6.18 123

6 326 3.77 40

7 969 2.28 71

9 325 1.14 12

10 267 1.59 14

11 503 1.23 20

51 313 1.25 13

52 236 0.99 8

53 146 0.94 4

Total Indicated 9,774 2.33 732

Total Measured and Indicated 17,733 2.09 1,192

Inferred 1 54 2.30 4

2 562 1.59 29

3 102 3.22 11

4 110 0.97 3

7 257 1.44 12

8 497 0.61 10

9 232 1.16 9

10 16 1.34 1

Total Inferred 1,830 1.32 78

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Table 1.7: Mineral Resource Statement at a Range of Cut-off Grades, Agbaou Project, SRK Consulting (Pty) Ltd, March 2012 using MIK

Classification Cut-Off Grade (Au

g/t)

kt Grade (Au g/t) kozs

Measured 0.3 7,959 1.8 460

0.5 6,262 2.2 438

0.8 4,570 2.7 403

1.0 3,797 3.1 381

2.0 1,939 4.8 298

Indicated 0.3 9,774 2.3 732

0.5 8,708 2.6 719

0.8 6,981 3.0 683

1.0 6,114 3.3 658

2.0 3,428 4.8 533

Measured and Indicated 0.3 17,733 2.08 1,193

0.5 14,970 2.43 1,157

0.8 11,551 2.88 1,086

1.0 9,911 3.22 1,039

2.0 5,367 4.80 831

Inferred 0.3 1,830 1.3 78

0.5 1,473 1.5 73

0.8 996 2.0 63

1.0 770 2.3 57

2.0 236 4.3 33

The mineral resources are reported in a manner that is consistent with CIM Definition

Standards as set out in NI43-101.

1.9 MINERAL RESERVE ESTIMATES AND MINING METHODS The mining section of the report includes discussion on the open pit optimisation, practical

pit design, scheduling process, options investigated and the reasons behind selections

made. The mineral reserves, the results of the mine design process in terms of production

schedules as well as capital and operating cost estimates are presented.

A block model estimated by MIK (Multiple Indicator Kriging) was used as the basic resource

model for the pit optimization study. The amount of mineral resource using a 0.5g/t Au cut-off

was 15 Mt with a gold content of 1.2Moz. Only mineralized material in the measured and

indicated categories was taken into account.

The Whittle/Gemcom Four-X Analyser software provides guidance to the potential economic

final pit geometries. Whittle-4X was used to identify the optimum pit shell in terms of value

and tonnage. Mining by contractor has been selected by ENDEAVOUR for the mining

operations. This has been the basis cost estimations. For comparison, an owner operated

option has also been prepared. The optimum pit shell in the contractor option contains

approximately 11.4Mt of ore with average Au grade of 2.55g/t. A gold price of US$1200/oz

was selected by ENDEAVOUR which corresponds to other mining studies being undertaken

elsewhere at this time. The cut-off grade determined to be 0.53 g/t for oxide and 0.51 g/t for

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fresh ore within the Whittle algorithm. A sensitivity analysis was prepared by varying the unit

mining cost, process and administration and the gold price by ±10%. The sensitivity results

on cashflow showed the gold price is the most sensitive followed by the mining cost then

process and administration cost.

The practical pit designs were prepared using the optimised pit shells as templates. Surpac

software was used to prepare the practical pit, and to incorporate the haul roads and ramps

together with the appropriate inter-ramp slope angles. A total of five pits have been designed

for North and South parts of Main orebody (interim and final pit) and West orebody. The

Proven and Probable mineral reserves estimated to be contained within the practical pits

(and by SAMREC definition) was approximately 11.1Mt with contained gold of about 0.905

Moz. The mineral reserves are reported in a manner that is consistent with CIM Definition

Standards as set out in NI43-101.

The total amount of waste to be mined and hauled to the waste dumps is 83.9Mt. Three

positions have been provided by ENDEAVOUR for waste dumping. The waste dumps have

been design based on the practical parameters to contain a total of approximately 50 million

cubic meters. Table 1.7 summarizes the Proven and Probable Reserves estimated for the

project.

Table 1.7: Summary of Agbaou Mineral Reserves Reserve Category

Deposit

Tonnes Grade Ounces

(Million) (g/t Au) (Million)

Proven Agbaou 5.407 2.25 0.390

Probable Agbaou 5.668 2.82 0.515

Total Proven and Probable Mineral Reserves 11.075 2.54 0.905

Mine production scheduling also adjusted to meet plant feed criteria. The proposed Agbaou

process plant is based on a 1.6 million tonnes per annum (Mtpa) for saprolite and 1.34Mtpa

for bedrock. The upper portion of the orebody consists of relatively soft rock (Laterite and

Saprolite) and the majority of ore mined in the first 4 years will be soft rock. The proportion

of hard rock will then increase.

The mining operations are based on the use of hydraulic excavators and haul trucks suitable

for conventional open pit mining techniques. The main arterial roads where necessary will be

constructed to a minimum 20 m width, including berms and drainage areas.

Topsoil in mining areas will be recovered during the pit preparation phase and stockpiled for

future use with progressive waste dumps and possible pit rehabilitation. The in-situ materials

in hard and semi-hard rock will require drilling and blasting to assist fragmentation and

subsequent loading. Based upon the information supplied, and experience in similar

operations, it has been assumed that that oxide portion of the orebody is suitable for free

digging but may require light blasting in certain areas.

The mining equipment and infrastructure has been selected based upon the mine production

schedule. This has been prepared by the mining contractor. An owner option in capital and

operating cost was also included as an alternative in the study (SRK Consulting, SENET,

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Knight Piésold Consulting, 2012). The total operating costs for the life of mine, estimated by

contractor are shown to be approximately 335 M$. ENDEAVOUR’s management including

labour costs based on the manpower requirements, administration cost of 0.25$/t.ore and

grade control costs of 1.2$/t.ore are to be to be added to this value.

Based on the Hydrogeological study conducted by Knight Piésold, estimations of the impact

that the proposed mining activities could have on the groundwater environment at various

phases of these activities, as well as recommendations on possible mitigation methods to

contain or minimize these impacts were provided. Additionally, an assessment as to the

magnitude of groundwater inflows to the proposed open pit was conducted along with a

determination of the feasibility of groundwater for process and potable water supply.

1.10 RECOVERY METHODS This section outlines the design criteria and the process description for the Agbaou process

plant. The design criteria were developed on the basis of metallurgical testwork results

obtained from both the previous study and the current EOS. Several references that include

information from ENDEAVOUR, calculations, industry practices and assumptions have been

used to derive data used in the design criteria.

The comminution circuit of the process plant will comprise a primary jaw crusher, followed by

the SAG and ball mills. The milling circuit will include crushing of the pebbles generated from

the SAG mill. A dedicated gravity circuit consisting of a concentrator, intensive cyanidation

package and an electrowinning cell will recover free gold from a portion of the milled product.

The rest of the milled product will be processed in the Carbon in Leach (CIL) circuit where

gold contained in the ore will be leached and adsorbed onto activated carbon. The CIL tails

slurry will undergo cyanide destruction prior to disposal in the tailings dam. Loaded carbon

will be acid washed and rinsed prior to elution. The electrolyte leaving the Anglo American

Research Laboratory (AARL) elution circuit will undergo electrowinning where gold sludge

will be produced. The sludge will be dewatered using a pot filter and dried in a drying oven

ahead of smelting. Fluxes will be added to the dried gold sludge and the mixture placed in

the smelting furnace. After smelting the furnace crucible contents will be poured into

cascading moulds. The gold bars will be cleaned, sampled, labelled and prepared for

shipping.

1.11 ENVIRONMENTAL STUDIES AND SOCIAL IMPACTS An Environmental Impact Assessment (EIA) was undertaken from December 2007 to July

2008 to investigate the local environmental and social situation existing prior to the

development of the Project and to determine the likely positive and negative impacts of the

Project. The timing, extent, intensity and cumulative effects of these impacts were

investigated. The EIA identifies the necessary management measures required to mitigate

the identified impacts. These form the basis of the Environmental Management Plan (EMP)

and the Relocation Action Plan (RAP).

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1.12 INFRASTRUCTURE, WATER SUPPLY, TAILINGS STORAGE AND LOGISTICS

The selected Agbaou site is a greenfields site without any infrastructure except for the

existing Agbaou Exploration Camp. The proposed infrastructure will support the mining,

plant and construction operations. The main infrastructure required for the development of

the project will be:

• Raw water supply system; • New access road of 9.17km to the plant site facilities, as well as 2.47km to the main

camp; • Site roads; • Camp accommodation and catering facilities for 200 people, including: kitchen,

messing, laundry, potable water supply system and sewage disposal unit; • Mining workshops with internal offices, change house, wash bay, refueling station

(external contract), tire change, and explosives storage. Potable water piped form the from process plant;

• Plant workshop; • Warehouses and lay down yards; • Plant administration buildings and medical facilities; • Assay laboratory; • Reagents storage building; • Change house; • Stand-by power plant providing 2.4MW; • Communications; • Security; • Sewage treatment and disposal; • Plant laundry.

Materials and goods will be trucked to the process plant, camp and mining facilities from the

port of Abidjan.

Tailings Storage Facility (TSF) A suitable location for the TSF was identified by considering the general topography, water

course locations, the required size of the TSF based on capacity requirements and the

general geology of the site.

The proposed construction method utilizes a downstream construction method consisting of

three phases utilizing earth fill from local borrow areas for the initial phase and overburden

material from the open pits during the final two phases.

The TSF does not incorporate an HDPE liner as the process plant includes a cyanide

destruction facility. Since seepage through the basin and dams foundations needs to be

controlled, several key elements have been incorporated into the design. All design

considerations of the tailing storage facility have been based on meeting or exceeding the

agreed design criteria which comply with World Bank Standards, Côte d’Ivoire and other

international standards.

The TSF has been designed to retain water from rainwater and from the tailings. To remove

water from the TSF and return the water to the plant, a floating barge will be utilized.

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Water Storage Dam (WSD) During plant start-up a significant water source will be required as reclaim water from the

Tailings Storage Facility will not be available. Also the water balance for the plant operations

indicates that make-up water will be required as the TSF will not be capable of providing

sufficient water during the dry season to meet the plant requirements. Two suitable storage

locations were identified.

1.13 CAPITAL AND OPERATING COSTS The aim of the capital and operating cost estimates (excluding WSD) is to provide costs to

an accuracy level, in the opinion of those listed in Section 1.1, of ±10%. These costs

excluding escalation, financing costs and schedule changes can be utilized to evaluate the

economics of the Agbaou Project when treating 1.6Mtpa. All costs are presented in United

States Dollars (US$).

Capital Cost Estimate The total estimated cost of bringing the Project into production is US$158,939,551 and is

inclusive of US$11,777,985 contingency and US$5,039,749 working capital. Table below

gives a summary of the life of mine capital requirements.

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Table 1.8: Capital Cost Summary US$ % US$

Process Plant Direct Costs

Machinery & Equipment 17 732 002 5% 18 618 603

Civils & Earthworks 11 556 402 12% 12 887 449

Structural Steel & Platework 5 803 972 10% 6 384 369

Piping & Valves 1 433 928 10% 1 577 321

Electrical & Instrumentation 5 094 739 12% 5 686 298

Transportation 5 554 000 15% 6 387 100

Subtotal 47 175 044 51 541 140

Infrastructure Costs

Tailings (Start-up only) 5 622 127 10% 6 184 340

Standby Power Plant 1 283 884 5% 1 348 078

OHL Grid Power to Plant 6 500 000 0% 6 500 000

Access Roads 2 467 329 20% 2 960 795

Main Camp 4 951 394 5% 5 198 964

Onsite Infrastructure Buildings etc 2 649 174 15% 3 046 550

Raw Water Supply 1 565 642 10% 1 722 207

Mining Facilities 727 079 10% 799 787

Communications 53 989 5% 56 689

Vehicles 1 564 400 5% 1 642 620

Mobile Plant 2 308 820 10% 2 539 702

Subtotal 29 693 838 31 999 730

Plant Pre-production

Plant First Fill 899 952 5% 944 950

Spares 3 712 670 5% 3 898 303

Subtotal 4 612 622 4 843 253

Mining Capital Costs

Mining Contractor

Mobilisation 10 372 000 0% 10 372 000

Mining Pre-Strip 4 384 749 0% 4 384 749

Mining Management - Pre-Strip

Period 1 500 000 0% 1 500 000

Subtotal 16 256 749 16 256 749

Other

Insurances 1 511 760 10% 1 662 936

Relocation Cost 5 600 000 7% 6 000 000

Import Tax 1 146 763 0% 1 146 763

Vendor Services 756 553 10% 832 208

Subtotal 9 015 076 9 641 907

Management Costs

Project Management 15 500 000 15% 17 825 000

Owner's Preproduction Costs 11 345 774 5% 11 913 063

Working Capital 5 039 749 10% 5 543 724

Steel, Plate, Mech & Piping

Construction 3 763 419 10% 4 139 761

Electrical Construction 1 757 895 10% 1 933 685

Instrumentation Construction 282 373 10% 310 611

Infrastructure Construction 200 201 10% 220 221

Construction Equipment Hire &

Power 2 518 825 10% 2 770 708

Subtotal 40 408 237 44 656 772

GRAND TOTAL 147 161 567 158 939 551

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Operating Cost Estimate The annual operating costs for the life of the mine (LOM) were estimated for mining,

processing, general and administration (G&A), royalties and refining charges and are

summarized in the table below.

Table 1.9: Overall Operating LOM Costs

1.14 ECONOMIC ANALYSIS The Agbaou Project economic assessment was prepared with the input from ENDEAVOUR,

SRK Consulting, (mining and geology), SENET (processing plant, general administration

and infrastructure) and Knight Piésold (tailings management facility & water storage dam).

The assumptions used in the financial analysis were provided by ENDEAVOUR and are


Table 1.10: Assumptions Used in the Financial Evaluation Parameter Units Assumption

Gold Price US$/oz 1,250

Discount Rate % 5.0

Government Royalty % 3

Tax Rate after 5 years % 25

Fixed Equipment Residual Value % 10

Depreciation % 12.5

A financial analysis was carried out based on the above assumptions, the production

schedule, capital costs and operating costs. The results from the financial analysis are


Unit Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 LOM

Mining US$/t 21.79 27.33 29.87 29.70 48.18 35.60 28.66 29.54 30.94

Plant US$/t 9.26 9.37 9.35 9.59 9.76 9.61 9.69 9.61 9.51

Tailings Storage Facility (TSF) US$/t 0.14 0.15 0.15 0.17 0.15 0.15 0.15 0.15 0.15

G&A US$/t 4.91 5.19 4.97 5.62 6.16 5.72 5.83 5.76 5.46

Assay US$/t 1.01 1.06 1.06 1.20 1.31 0.79 0.81 1.55 1.06

Refining US$/t 0.26 0.35 0.43 0.45 0.32 0.39 0.37 0.48 0.37

Government Royalties US$/t 1.97 2.58 3.22 3.36 2.38 2.88 2.75 3.58 2.77

Total Operating Costs US$/t 39.33 46.03 49.05 50.08 68.27 55.14 48.25 50.67 50.26

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Table 1.11: Summary of Financial Analysis Results

Financial Summary Units 2nd

Q 2012

LOM Tonnage Ore Processed t 11 074,927

LOM Feed Grade Processed g/t 2.50

LOM Gold Recovery % 92.5%

LOM Strip Ratio

7.87

LOM Gold Production oz 821,728

Production Period years 8.00

Gold Annual Production- LOM oz 102,716

LOM Direct Operating Costs US$/oz 635

LOM Total Cash Operating Costs US$/oz 677

LOM Total Cash Operating Costs US$/t 50.3

Total Capital Costs US$/oz 217.8

Total Production Costs US$/oz 895

Post Tax NPV US$ million 184.3

IRR % 27.7%

UnDiscounted Payback Period years 2.53

Project Nett Cash Flow after Tax and Capex US$ million 273.8

LOM cashflow's were calculated and project sensitivity analysis was conducted. These are

shown in Section 22. When sensitivities were ranked, the sensitivity analysis indicated that

the project is most sensitive to gold price fluctuations followed by operating costs and then

capital expenditure.

1.15 INTERPRETATION AND CONCLUSIONS

Mineral Resource The currently defined Mineral Resources coincides with the areas within the original

Exploration License area where the best mineralization was detected in soil samples.

However, recent sterilization drilling has indicated that the potential exists to discover more

mineralization outside the currently modeled areas and also in those areas not explored

during the soil surveys.

In the opinion of SRK the sampling preparation, security and analytical procedures used by

ENDEAVOUR were consistent with generally accepted industry best practices and are

therefore considered satisfactory.

SRK believes that the current QAQC systems in place at Agbaou to monitor the precision

and accuracy of the sampling and assaying is adequate and the BV Laboratory returned


The 2012 mineral resource estimate is an update of the mineral resources estimated by

Coffey Mining in 2008, following the completion of an additional phase of drilling on the

project. The mineralization has been re-modeled to include the new intersections, but the

Coffey interpretation was also revised as follows:

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• Additional zones were modeled parallel to the main mineralized zones where sufficient continuity along strike and down dip was observed;

• The major zones were extended along strike where additional drilling had intersected the mineralization.

The SRK interpretation reduced the low grade dilution within the mineralized envelopes and

this resulted in a trend towards decreased tonnes at a higher grade with some isolated high

grade samples no longer included in the mineralized envelopes.

Following a composite optimization exercise, SRK composited all samples within the

mineralized envelopes to 2m lengths, forcing all samples to be included in the composites by

adjusting the composite length at the base of the mineralisation. Nine of the fourteen zones

required capping of high grade outlier values before estimation to limit the impact of the

unusually high values.

SRK was able to generate robust semi-variograms for the better informed zones, however

for some of the smaller modelled zones there was insufficient data available to generate

reliable experimental semi-variograms. In zones where it was not possible to generate

robust semi-variograms they were grouped together along with better informed zones with

robust semi-variograms, thus borrowing the semi-variograms from the well informed zones.

For the well informed zones the semi-variograms show robust structures with long ranges

between 50m and 100m, which is in excess of the drill hole spacing in the densely drilled

areas.

All zones were estimated by Ordinary Kriging using the optimized search parameters, with a

three phase search strategy. The first pass estimates used the semi-variogram ranges to

define the search. The second and third passes using expanded search ranges were

designed to ensure the best estimates possible beyond the semi-variogram range, but also

ensure that all blocks were estimated. This was required as in some areas the interpreted

mineralization showed abrupt changes in orientation, possibly related to smaller scale

faulting.

In addition to the Ordinary Kriged estimates for the zones with sufficient information and

reliable semi-variograms, SRK generated a UC and MIK estimate. These zones are zones 1,

2, 3, 4, 7, and 9, which (excluding the laterite) make up 90% of the volume of the

North/Central/South area. The UC and MIK estimates were validated against the Ordinary

Kriged estimates at 0g/t cut off and against each other at a range of cut off values. As the

UC estimate is based on the Ordinary Kriged estimate, the estimates correlate exactly, while

the MIK estimates show a higher grade without applying a cut off. The MIK estimates also

typically show a higher grade at cut-off values with the recoverable grades converging at

higher cut-off values. The MIK estimate therefore estimates that a greater degree of

selectivity will be achievable than is estimated by the UC method. The MIK estimate was

selected for the final resource reporting to be consistent with previous resource estimates,

however there is a risk that the selectivity modeled using MIK is over estimated and that the

recoverable tonnes above the selected cut off may be greater than estimated at a lower

grade.

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A number of validations were complete on the Ordinary Kriged, UC and MIK estimates which

indicated relatively good agreement with the source data and the estimates. The mineral

resources are reported above a 0.3g/t cut off value which was calculated using the operation

parameters used in the Reserve calculations and an optimistic gold price of US$2,000/oz.

Mining The updated Geohydrology report and Geotechnical review were not assessed by SRK.

Based on the previous, earlier report there are no reasons that the Agbaou deposits cannot

be mined successfully.

Process Plant and Economic Evaluation Based on the results of all testwork conducted, the Agbaou Gold Plant has been designed to

treat 1.6Mtpa saprolite ore or 1.34Mtpa bedrock ore. The plant design incorporates a

conventional gravity and CIL circuit design.

The key project features are:

• Processed tonnage 11.1 million tonnes • Grade 2.5 g/t • Recovery 92.5% • Life of Mine (LOM) 8 years • Annual gold production 102,716 oz per annum • Total cash costs US$677 per oz. • Capital cost US$159 million

It is in the opinion of SENET that the capital costs provided are within the 10% accuracy

required for this level of study. The operating costs were derived from accepted industry

standards, first principle calculations, SENET in-house database and input from

ENDEAVOUR.

The results of the financial analysis indicate the viability of the Agbaou Gold Project. The

expected after-tax NPV, IRR and payback period based on a gold price of US$1,250 per

ounce are indicated below:

• NPV at 0% discount rate is US$273.8 million; • NPV at 5% discount rate is US$184.3 million; • IRR 27.7%; • Payback period 2.53 years.

1.16 RECOMMENDATIONS

Mineral Resource The current mineral resource estimates contain sufficient material to support reserves for

seven years of mining. No additional exploration is required to support the mining plan at this

time. Peripheral to the reported resources, additional potential mineralization targets have

been intersected in condemnation drilling holes.

It is recommended that an alternative accredited commercial laboratory is selected for the

Umpire Laboratory function to confirm or improve the poor repeatability in the results

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obtained from both SGS Laboratories in Ghana and Canada.

Mining It is recommended that the location of the waste dumps be given more attention. In addition

to this more detailed quotes should be obtained from potential mining contractors in order to

completely evaluate the option of owner mining versus contractor mining.

Metallurgical Testwork Additional testwork is recommended on saprolite samples to ensure the expected recovery

and comminution results used are accurate. Viscosity modifier testwork on the thickener

underflow product is also recommended.

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SECTION 2. INTRODUCTION

2.1 PROJECT OVERVIEW The Agbaou Project is an advanced stage gold project, located in Côte d’Ivoire, West Africa.

The property is 200km north-west of Abidjan, the country’s former capital and major port city.

The exploration permit for the Agbaou Project is fully owned by Etruscan Resources Côte

d’Ivoire SARL (“ERCI”). Endeavour Mining Corporation (Cayman Islands), a TSX-listed and

ASX-listed company, owns 100% of ERCI through its subsidiary companies.

A feasibility study was completed in September 2009 by MDM and Coffey. In September

2011, ENDEAVOUR commissioned various consultants (SRK Consulting, SENET and

Knight Piésold) to review the previous feasibility study and conduct an Engineering

Optimization Study (EOS) incorporating updated resource and reserve estimates and taking

into account some of the limitations that had been noted in the previous study. The findings

of the EOS are summarized this technical report and are valid with an effective date of 25th

May 2012.

This report was prepared following the guidelines of the Canadian Securities Administrators’

National Instrument 43-101, 43-101CP and Form 43-101F1. This technical report documents

a mineral resource and reserves statement for the Agbaou Project prepared by SRK. The

mineral resource and reserves statement reported herein was prepared in conformity with

generally accepted CIM “Estimation of Mineral Resources and Mineral Reserves Best

Practice Guidelines” dated 23 November 2003 and “Definition Standards on Mineral

Resources and Mineral Reserves” dated 27 November 2010.

This report summarizes the technical information available on the Agbaou Project and

demonstrates that the Agbaou Project clearly qualifies as an “Advanced Property” as defined

by the Toronto Stock Exchange. In the opinion of the authors of this report, this property has

merit warranting additional development expenditures. This report summarizes the capital

expenditures required to construct the project. Operating costs and other project parameters

that were used to derive the mineral reserves are summarized within this report.

2.2 PURPOSE OF TECHNICAL REPORT The purpose of this study is to demonstrate the viability of the Agbaou Project through:

• Updated mineral resources and reserves; • Economic evaluation taking into account improved resources/reserves and updated

capital and operating costs.

This technical report is a complete summary of the full study as prepared by SRK

Consulting, SENET, Knight Piésold Consulting (2012).

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2.3 SCOPE OF SERVICES

2.3.1 SRK’s Scope of Services The scope of work, as defined in a letter of engagement executed on 17th November 2011

between ENDEAVOUR and SRK includes the construction of a mineral resource model for

the gold mineralization delineated by drilling on the Agbaou Exploration Permit and the

preparation of an independent technical report in compliance with National Instrument 43-

101 and Form 43-101F1 guidelines.

SRK is responsible for all or part of sections 1 to 12, 14 to 16, 18, 21 to 23, 25 to 27 of this

technical report.

The SRK work included the assessment of the following aspects of the project:

• Topography, landscape, access; • Regional and local geology; • Exploration history; • Audit of exploration work carried out on the project; • Geological modeling; • Mineral resource estimation and validation; • Preparation of a mineral resource statement; • Recommendations for additional work.

The mineral reserve estimates were also prepared by SRK. The preparation of the mining

and reserve estimation portion of the study includes the following activities:

• Mining approach and methods; • Geotechnical data; • Hydrological data; • Open pit optimization and design; • Waste dumps design; • Fleet and haulage requirements; • Scheduling and production schedule; • Manpower requirements; • Mine capital costs; • Mine operating costs.

2.3.2 SENET’s Scope of Services SENET is responsible for all or part of sections 1, 2, 3, 13, 17 to 22, 24 to 27 of this technical

report. The SENET scope of work included the following:

Process Design Review and Optimization

• Review the existing testwork and identify any shortfalls; • Manage the additional testwork program; • Produce a design criteria; • Produce revised datasheets, calculations and schedules (pump/valve/line); • Compare the SENET datasheets, calculations and schedules to the existing process

design datasheets, equipment specifications and schedules, and flag any major differences;

• Review the differences with ENDEAVOUR, and depending on the severity and

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potential impact on other disciplines, amend the process design; • Oversee a comminution specialist to review comminution circuit design with a view to

evaluate and optimise the specified circuit; • Review and update process flow diagrams (PFD) and piping & instrumentation

diagrams (P&ID); • Review and update the process plant operating cost estimate with updated reagent

costs and changes to reagent consumptions where changes to the process circuits were made as a result of the review and optimisation process.

Engineering Design Review and Optimization (Process Plant) The engineering design associated with the process plant was reviewed and optimized

(where deemed necessary). The level of completeness of the existing engineering design

was evaluated against the requirements of the estimating accuracy associated with a

feasibility study. All identified shortcomings were discussed with ENDEAVOUR, and any

additional work associated with addressing these shortcomings was addressed.

The following activities formed the basis of the engineering design review and optimization:

• Review of plant layout and general arrangement drawing; • Review of plant-wide electrical single-line diagrams; • Review of civil/structural bulk earthworks and surface infrastructure drawings; • Review of civil design and civil quantities associated with the process plant; • Review of primary steelwork / structural designs and quantities.

Infrastructure, Utilities and Services Review and Optimization SENET reviewed the existing specified process plant infrastructure and optimized the

infrastructure where possible. This review was characterized by the following key activities:

• Revision of the existing material take-offs for earthworks, civil design and roads based on the existing process plant capacity of 1.2Mtpa, previous geotechnical reports and previous work carried out;

• Revision of process plant infrastructure layout drawings; • Assessment of local availability of bulk construction materials.

Revalidation of the Capital Cost Estimate The capital cost estimate associated with the processing plant and associated infrastructure

was revalidated, generally in accordance with the following activities:

• Re-pricing of the existing material take-offs and schedules for civil works, earthworks, buildings, structural steel, piping, valves and major electrical items;

• Obtaining a minimum of 3 written quotations for all major mechanical equipment and single quotations for minor mechanical equipment;

• Reviewing and updating (re-price only) of existing estimates for freight and construction/installation costs for mechanical, civil, structural, piping and electrical disciplines;

• Estimating EPCM costs including engineering, procurement, construction management and insurances.

Additional Inclusions – To Scope of Work Any activity or work associated with the overall coordination, assembly, formatting and filing

of the NI43-101 report or any press releases associated with the study.

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Development of a Project Implementation Plan A project implementation plan was prepared which included critical activities and labour

resources for the completion of the project for:

• Detailed design; • Procurement – long lead items; • Construction; • Commissioning.

2.3.3 Knight Piésold’s Scope of Services Knight Piésold is responsible for all or part of sections 1, 3, 18, 21, 25 to 27 of this technical

report.

Knight Piésold was retained to complete a Hydrogeological Bankable Feasibility Study

(HBFS) of the Agbaou Project to estimate the impact of the proposed mining activities and

an Engineering Optimization Study (EOS) of the Tailings Storage Facility (TSF) for the

Agbaou Project.

2.3.3.1 Hydrogeological Study (HS)

The following scope of work for the HBFS was included in Purchase Order No. C010401-01 /

E0002 / 2011 that was executed on 30th September 2011 between Endeavour Mining

Corporation and Knight Piésold:

Site Related Tasks

• Site visit; • Drilling supervision; • Aquifer testing;

Office Related Tasks

• Desktop study; • Appointment of drilling subcontractor; • Aquifer test analysis; • Data coding and computerization; • Define conceptual hydrogeological model; • Source and liner optimization modeling (TSF); • Unsaturated zone modeling (TSF); • Saturated zone aquifer characterization modeling; • Contaminant transport modelling; • Preliminary dewatering design.

2.3.3.2 Engineering Optimization Study of Tailing Storage Facility and Water Supply Dam

The following scope of work for the EOS was included in Purchase Order No. C010401-02 /

E0004 / 2011 that was executed on 29th November 2011 between Endeavour Mining

Corporation and Knight Piésold:

Tailings Storage Facility

• Site selection and confirmation of design criteria; • Complete capacity analysis according to design criteria; • Preliminary geotechnical investigation at the TSF site;

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• Hydrology analysis; • Water balance according to hydrology analysis; • Preliminary design including seepage analysis and stability analysis; • Risk assessment; • Closure design; • TSF Capital and operating cost estimates; • Gap analysis of open pit:

o Review of 2008 report produced by Golder Associates Inc.; o Report on review findings.

In addition to the scope of work outlined in the purchase order, ENDEAVOUR requested KP

to identify site locations for potential water storage sites. Five sites were located with two

locations identified as suitable sites by ENDEAVOUR. This of work entailed:

Water Supply Dam (WSD)

• Identify suitable locations and provide estimated water storage capacities; • Preliminary WSD capital

The work for the WSD did not include specific design elements required for the water dams

or any piping facilities to or from the WSD.

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SECTION 3. RELIANCE ON OTHER EXPERTS

SENET, SRK and Knight Piésold have not performed an independent verification of land title

and tenure as summarized in Section 4 of this report. The authors of this report did not verify

the legality of any underlying agreement(s) that may exist concerning the permits or other

agreement(s) between third parties, but have relied on an opinion dated 20th August 2010,

from Theodore Hoegan of Theodore Hoegan and Michel Ette, Advocates Associes

(Appendix 3-A).

The authors were informed by ENDEAVOUR that there are no known litigations potentially

affecting the Agbaou Project.

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SECTION 4. PROPERTY DESCRIPTION AND LOCATION The Agbaou Project is located in Côte d’Ivoire, West Africa, approximately 200km north-west

of the port city of Abidjan. Travelling time between Abidjan and the Agbaou Project area is

approximately 3 hours by car. The concession is reached by tarred and secondary gravel

roads and is within 10km of the national electrical power grid. The small town of Agbahou is

located approximately 12km north-east of the regional town of Divo (Figure 4.1).

Figure 4.1: Map Showing the Location of the Agbaou Project

Agbaou Gold Project Locality Map

Project

No.

439430

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4.1 MINERAL TENURE Figure 4.2 shows the original Agbaou Exploration Permit outline and the subsequent

decrease in the total area to the current Agbaou Exploration Permit after first renewal. The

corner co-ordinates of the area applied for in the application for an Exploitation Permit are

given in Table 4.1.

The project falls within the Agbaou Exploration Permit 177 of 939km2, which was reduced in

the subsequent application for renewal to 469km2. These permit areas refer to the granted

exploration permits whereas the area of the exploitation permit application is 334km2. The

Exploration Permits were granted for all minerals. The permit holder is not entitled to surface

rights and is required to pay compensation for crop damage and other losses to land

holders.

In terms of known royalties, back-in rights, payments, or other agreements and

encumbrances to which the property is subject SRK is only aware of a 10% free carried

interest of the Ivorian government and a 5% free carried interest held by Société pour la

Development Minier de la Côte d’Ivoire (SODEMI).

Table 4.1: Corner Co-ordinates for Proposed Agbaou Exploitation Permit Area Corner UTM Co-ordinates, Zone 30 ,WGS 84

X Y

I 249,349 675,523

II 252,629 680,119

III 252,666 688,907

IV 263,736 688,861

V 267,999 686,002

VI 262,526 678,419

VII 262,465 663,055

VIII 249,298 663,108

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Figure 4.2: Mineral Tenure

Agbaou Gold Project

License Areas

Project No.

439430

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4.2 UNDERLYING AGREEMENTS The security of tenure for the Agbaou Project is based on the legal opinion, provided as a

letter to SRK, by Theodore Hoegan of Theodore Hoegan and Michel Ette, Advocates

Associes, on a Ministerial (Ministere Des Mines et de L’Energie) authorization (No 1625) to

commence with the development work for the project, dated 20th August 2010 (Appendix 4-

A).

The different licenses are discussed above and shown in Figure 4.1 and Figure 4.2 with

corner coordinates of the exploitation permit area listed in Table 4.1. Due to the fact that the

licenses expired and as yet none of the new licenses have been awarded, SRK is reliant on

the legal opinion expressed by Theodore Hoegan in the following terms:

• “Under the provisions of article 2 of the Mining Code all minerals contained in the ground and bedrock are the property of the State de Côte d'Ivoire. The State may, however, authorize any person or entity, to carry out exploration or exploitation of minerals. This authorization is granted under a mining title (exploration or exploitation permit), and this pursuant to the provisions of article 4 of the Mining Code;

• Pursuant to the provisions of article 4 referred to above, the State of Côte d'Ivoire awarded to ERCI, the exploration Permit 177 (Agbaou), listed on the special register of the land conservation of the State of Côte d'Ivoire under number 177, and relates to a surface area of nine hundred and thirty-nine (939) km2;

• The Exploration Permit 177 (Agbaou) was renewed, by decree n0011/MME/DM on 22nd March 2007, by the minister for the Mines and Energy for a period of two (2) years, beginning on November 27, 2006 and ending on 27th November 2008. The surface area was supposed to be reduced by half to equal 469km2, in accordance with the provisions of article 13 of the mining Code. However it was maintained with 939km2, in the First Decree of Renewal;

• In order to conform to the provisions of the Mining Code, ERCI communicated to the director of the Mines, in a letter dated 17th March 2008 new coordinates of the Exploration Permit 177 (Agbaou), reducing its surface to 469km2. The Exploration Permit 177 (Agbaou) coming to expiry, on 26th November 2008, ERCI lodged a request for second renewal of the permit relating to 469km2, which was received and accepted at the same time as it payment of the fixed rights of renewal;

• The decree for the second renewal of Exploration Permit 177 (Agbaou) has not, at present, been issued by the minister of Mines and Energy;

• The provisions of article 25, subparagraph 3 of the Mining Code provides, in such circumstances, that a request for renewal of a mining title, made before the expiry of its period of validity, extends the validity of this mining title automatically, without formality. In addition, under the provisions of article 22 of the Decree of the Mining Code, the renewal of the exploration permit shall be effective on the date of the original validity of the permit. The decree specifies the original date of renewal;

• Exploration work carried out on the Exploration Permit 177 (Agbaou) having identified an economically viable deposit, ERCI filed by letter dated 25th October 2009, an application for an exploitation permit which was attached to a file containing the documents specified by the provisions of article 28 of the Decree of the Mining Code;

• It should be noted that the exploration permit holder is entitled to the exploitation permit granted by Council of Ministers decree on a proposal from the Minister of Mines, after the exploration permit holder provided evidence of the existence of a deposit within its exploration permit. This evidence is furnished with a feasibility study;

• The feasibility study was carried out by ERCI, and was filed with the mining administration, on 30th September 2009. The application for an exploitation permit

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was examined by the Commission Mine Interministérielle (COMINE), which approved the feasibility study, at its meeting of 7th September 2010. ERCI is now awaiting the Decree from the Council of Ministers granting it the exploitation permit;

• The exploitation permit, once granted, will confer on ERCI, the exclusive right of exploration and exploitation of the deposits within, with all rights attached therein including, transport or to cause the transport of extracted minerals, and disposal of extracted minerals on domestic or export markets, and to export them;

• Possession of a land title or ownership of the land, subject of the exploitation permit is not necessary to conduct exploitation operations. However, if the conduct of the exploitation operations requires the occupation of land belonging to third parties, the latter will be entitled to compensation under conditions fixed by the provisions of decree 95-817 of 29th September 1995 establishing the rules of compensation for destruction for cultures and the inter-ministerial decree Na 028 MINAGRNMEF of bearing 12th March 1996 fixing the scale of compensation for destroyed crops.”

The owner of the Agbaou Exploration Permit and, if converted, the Agbaou Exploitation

Permit is ERCI. Although, ERCI has applied for the Exploitation Permit, it has given its

consent for another subsidiary of Endeavour Mining Corporation, Agbaou Gold Operations

SA, to be granted the Exploitation Permit. Endeavour Mining Corporation (Cayman Islands),

a TSX-listed and ASX-listed company, owns 100% of ERCI through its subsidiary companies

and will own 84.8% of Agbaou Gold Operations SA through its subsidiary companies. The

remaining 15.2% of Agbaou Gold Operations SA will be held by the State of Côte d’Ivoire

(10% free carried interest), Société pour le Dévelopement Minier de la Côte d’Ivoire

(“SODEMI”; 5% free carried interest) and two individual shareholders holding each one

qualifying share (0.2%). The two individual shareholders are not free carried interest.

4.3 PERMITS AND AUTHORIZATION The following documents are included in Appendix 4-A:

• The original decree awarding Agbaou to ENDEAVOUR; • Application for the amended Exploitation License.

4.4 ENVIRONMENTAL AND SOCIAL CONSIDERATIONS An environmental and social impact assessment has been undertaken from December 2007

to July 2008 to investigate the local environmental and social situation existing prior to the

development of the project and to determine the likely positive and negative impacts of the

project. The timing, extent, intensity and cumulative effects of these impacts have also been

investigated.

The potential negative environmental and social impacts of the project include: land clearing

that will result in the loss of natural habitat for local flora and fauna; displaced local

subsistence agriculture; contamination and degradation of soils through exposure and land

clearance; contamination and degradation of surface waters through industrial spills; siltation

and flow modifications to local streams will impact aquatic life; dewatering activities around

open pits will impact local groundwater levels; degradation of local air quality resulting from

increased vehicle presence, increased noise associated with mining and processing;

increased traffic volumes; mine blasting in open pits may generate vibrations that could

negatively impact local infrastructure; potential for acid mine drainage; increased prevalence

of HIV and AIDS through the changes in the demography of the local populations, possible

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conflicts between immigrants and local communities for mine jobs; relocation of the affected

communities in the project area to Agbahou Village may lead to economic upheaval,

hardships and tensions between the relocated population and their host communities. The

loss of agricultural land in the project area is expected to impact on 76 farmers whom will

require compensation.

Other significant factors and risks that may affect access, title, or the right or ability to

perform work on the property are:

• Côte d’Ivoire has been experiencing a period of political unrest; • ENDEAVOUR management continues to believe that the political situation in Côte

d’Ivoire will not have a significant impact on the long-term of the project or the recoverability of its investments in the Agbaou property; should the current political unrest continue or worsen, it may have a negative effect on the recoverability of this investment.

4.5 MINING RIGHTS IN CÔTE D'IVOIRE A mining right was applied for on 6th October 2008 as stated above but has not yet been

awarded.

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SECTION 5. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 ACCESSIBILITY The nearest town to the Agbaou site is Divo. Divo has a population of approximately 128,000

and is located 200km from Abidjan on highway A2. The permit is accessed from Divo by

taking an asphalt road north approximately 25km to the village of Didoko and then an

unpaved secondary road east for 12km to the village of Agbahou. The Agbaou gold project

is adjacent to the Agbahou village and is accessed by a series of dirt roads, tracks and

footpaths (Figure 4.2).

5.2 LOCAL RESOURCES AND INFRASTRUCTURE The national electrical grid runs along the main road that connects the town of Divo to

Didoko. A secondary, high voltage electrical line branches off the national grid at Didoko and

passes through the village of Agbahou. Mobile phone coverage has increased significantly in

recent years and a telecommunications tower has been located adjacent to the

ENDEAVOUR Mining Exploration Camp. The nearest medical clinic and police station are

located in Divo, where food and general supplies can be purchased.

5.3 CLIMATE Agbaou is within the Sud Bandama climate region. The Sud Bandama region is located

within the Southern tropical zone which runs inland from the coastline and is characterized

by three seasons; hot and dry (November to February), hot and wet (March to May) and

warm and wet (June to October). The average temperature range in the region is between

21°C and 33°C.

The precipitation estimate was taken from data at Gagnoa, about 80km west of the Agbaou

permit area. The data indicates there are two, distinct wet seasons March to June, and

September to October. June is the wettest month where rainfall can reach 300mm and the

average annual rainfall is estimated at between 950mm and 1,900mm per annum (Canty

and Associates LLC, 2012).

The area is predominantly subsistence farmland producing mainly food crops such as

plantain, corn, cassava, yam, tomatoes and some cash crop such as coffee, cocoa and palm

oil.

5.4 PHYSIOGRAPHY The general topography of the Agbaou permit is undulating hills that range from 130m to

420m above mean sea-level.

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Figure 5.1: Topography of the Project Area

Agbaou Gold Project

General Topography

Project No.

439430

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SECTION 6. HISTORY

6.1 PRIOR OWNERSHIP OF THE PROPERTY AND EXPLORATION Alluvial gold has been known and exploited by local “orpailleur” (artisanal miners) for some

time. Gold mineralization in bedrock was first reported in the Agbaou area during the late

1980’s, while the ground was held by a joint-venture between BHP Minerals and SODEMI.

The joint-venture controlled the ground between 1988 and 1994 and significant exploration

was undertaken during this period including regional and detailed soil sampling, pit sampling,

ground geophysics and a program of eight diamond drillholes (1,680 meters). Based on this

work BHP outlined an unclassified resource of 125,000 ounces and recommended no further

work.

Between 1996 and 2000 the property was held by Goldivoire S.A.R.L. (“Goldivoire”).

Goldivoire was an Australian - Ivoirian owned company controlled by Jandera Resources NL

(88.4%) which was a wholly owned subsidiary of Diversified Mineral Resources NL (“DMR”).

DMR was subsequently taken over by Hargraves Resources NL in mid-1999. Hargraves

Resources were in turn taken over by Durban Roodepoort Deeps (“DRD”) of South Africa in

December 1999. During the period 1996 to 2000, Goldivoire undertook an exploration

program that included semi-regional soil sampling, pit sampling, 36 Rotary-Air-Blast (“RAB”)

drillholes (1,682 meters) and a program of 203 Reverse Circulation (“RC”) drillholes (22,149

meters) of which 25 were extended with a diamond drill hole tail (1,535 meters).

Internal difficulties within DRD subsequently resulted in funds not being available for

Goldivoire to complete further work on the permit and as a result, the government of Côte

d’Ivoire withdrew the Agbaou permit. On 27th November 2003, after the completion of a

bidding process, the Ministry of Mines and Energy for Côte d’Ivoire granted the Agbaou

exploration permit to Etruscan Resources Côte d’Ivoire (Decree Number 2003-464).

6.2 HISTORICAL MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

A number of resource estimations have been completed on the Agbaou Gold deposit since

1998. The initial resource estimate carried out by Resource Services Group (“RSG”) in July

1998 was based on the BHP drilling plus 42 RC drill holes completed by Goldivoire. An

inverse distance squared method (“ID2”) was used and resulted in an estimate totalling

6.229 Mt at a grade of 2.8g/t Au for 564 Koz Au being outlined based on a 1.0 g/t Au cut-off

grade (Table 6.1). This grade estimate was unclassified. Note that in SRK’s opinion the

tonnage and grade estimates and the quantity of gold should be rounded to an appropriate

level of confidence.

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Table 6.1: Goldivoire, Resource Estimate (ID2. July, 1998) Lower Cut-off Grade Unclassified Mineral Inventory

g/t Au Tonnage (Mt) Grade Au g/t Koz Au

1 6.229 2.8 564

A second grade estimate was carried out by RSG in February 1999 and included an

additional 106 RC drill holes (some with diamond tails) completed by Goldivoire. The

Multiple Indicator Kriging (“MIK”) method was used for estimation and resulted in an

Indicated resource of 2.538 Mt at a grade of 2.9g/t Au for 237 Koz Au plus an Inferred

resource of 6.397Mt at 2.3g/t Au for 515 Koz Au being outlined based on a 1.0g/t Au cut-off

grade (Table 6.2). A block size of 10m E x 20m N x 5m RL was selected to reflect the

Selective Mining Unit. This historic resource estimate is not compliant with CIM Definition

Standards or NI43-101 requirements. These numbers should be rounded off to an

appropriate level of confidence.

Table 6.2: Goldivoire, Resource Estimate (MIK. February, 1999) Lower Cut-off Grade Indicated Inferred

Au g/t Tonnage (Mt) Au g/t k oz Au Tonnage (Mt) Au g/t k oz Au

0.5 3.750 2.2 265 11.23 1.7 614

0.8 2.984 2.6 249 8.381 2.1 565

1.0 2.538 2.9 237 6.397 2.3 515

1.5 1.857 3.5 209 4.666 2.9 435

DRD commissioned a resource estimate through RSG in 2000. A total resource was

reported at a 1g/t cut-off of Indicated 9.684Mt at an in-situ grade of 2.1g/t Au for 654Koz Au

plus Inferred 2.562Mt at an in-situ grade of 2.3 for 189Koz Au (Table 6.3). This historic

resource estimate is also not compliant with CIM Definition Standards or NI43-101

requirements. The numbers should be rounded off to an appropriate level of confidence.

Table 6.3: DRD, Resource Estimate (April, 2000) Lower Cut-off Grade Indicated Inferred


0.5 18.22 1.5 879 5.214 1.5 251

0.8 12.23 1.9 747 3.340 2.0 215

1.0 9.684 2.1 654 2.562 2.3 189

1.5 5.950 2.7 517 1.540 3.0 149

These estimates are considered historic and are not current and should not be relied upon.

The issuer is not treating the historical estimates as current mineral resources or mineral

reserves. No verification of the historical resource estimates was undertaken.

More recently an estimate of the Agbaou Mineral Resources was completed by Coffey

Mining Pty Ltd (“Coffey”) 21st February 2008 for ERCI. The Ordinary Kriging method was

used for estimating the laterite resources and Multiple Indicator Kriging was used for

saprolite, sap-rock and fresh rock. A total resource was reported at a 1g/t cut-off of Indicated

10.489Mt at an in-situ grade of 2.6g/t Au for 871koz Au plus Inferred 2.754Mt at an in-situ

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grade of 2.5 for 248Koz Au (Table 6.3). A block size of 5m E x 10m N x 5m RL was selected

to best reflect the Selective Mining Unit. The results of this exercise are presented as a

grade tonnage report in Table 6.4. These numbers should be rounded off in SRK’s opinion.

Table 6.4: Coffey Mining, Resource Estimate (February, 2008) Lower Cut-off Grade Indicated Inferred


0.4 17.785 1.8 1,032 5.514 1.6 278

0.5 16.590 1.9 1,015 5.072 1.7 272

0.6 15.221 2.0 991 4.552 1.8 263

0.7 13.945 2.2 965 3.995 2.0 251

0.8 12.697 2.3 935 3.526 2.1 240

1.0 10.489 2.6 871 2.754 2.5 248

1.2 8.747 2.9 810 2.222 2.8 199

1.5 6.807 3.3 727 1.682 3.3 176

2.0 4.700 4.0 610 1.085 4.1 143

This estimate by Coffey was previously reported under NI43-101 but is no longer current and

should not be relied upon. This report provides a current update of the resource estimate as

prepared by SRK and presented in Section 14.

So far the only gold produced from the area is the small amounts recovered from artisanal

workings on the streambeds draining the Agbaou deposit.

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SECTION 7. GEOLOGICAL SETTING AND MINERALIZATION

7.1 REGIONAL GEOLOGY Coffey (2008) states that the Côte d’Ivoire is largely underlain by a Paleoproterozoic Early

Eburnean plutonic and anatectic granite gneiss terrain with enclosed tectonic fragments of

older Birimian sedimentary and volcano-sedimentary rock types, forming greenstone belts

(Figure 7.1). Collectively these greenstone belts and surrounding granite gneisses form part

of the Man Shield (also known as the Leo Shield), which in turn forms the Southern half of

the larger West African Craton.

The Birimian Supergroup in Côte d’Ivoire have been sub-divided into an older, thick

sequence of metasedimentary basins and a series of greenstone belts consisting of meta-

volcanic rocks with intercalated meta-sediments (Junner, 1946). The greenstone belts occur

as relatively narrow (15-40km wide) but often long (< 200km) features with generally a

NE/SW trend. The effect of the Eburnean Orogeny on the Birimian Supergroup is best

described from studies undertaken in Ghana by Blenkinsop et al. (1984), Eisenlohr (1989),

and Eisenlohr & Hirdes (1992). Here a protracted event initially formed a regionally

penetrative S1 foliation which was followed by formation of high strain shear zones (S2)

along basin/belt contacts. Stress analysis suggests the direction of maximum compression

was sub-horizontal in a NW/SE direction and minimum compressive stress was vertical.

The metamorphic grade of the greenstone belts ranges from lower greenschist to

amphibolite facies, depending on the distance from the enveloping granitoids. Extensive

recent weathering has produced large areas of laterite over the region, which effectively

masks the underlying geology of these areas. As a result of the deep weathering, outcrop is

rare and even when it occurs is often difficult to characterize (Coffey, 2008).

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Figure 7.1: Regional Geology

Agbaou Gold Project

Structural Analysis

Project No.

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7.2 PROPERTY GEOLOGY The Agbaou gold deposit is hosted within the Birimian, Oumé-Fetekro Greenstone Belt and

the rocks within the permit are dominantly deformed, mafic volcanic rocks, metamorphosed

to greenschist to lower amphibolite facies (Figure 7.2). The greenstone belt is folded into an

anticline and the Agbaou deposit lies near the hinge of the fold on the eastern limb

(Eisenlohr, 1998). Bedding, foliation and the dominant vein-set are oriented along the strike

of the fold (roughly north-east/south-west) and dip to the south-east at a moderate to steep

attitude.

Because of the very poor outcrop, a geological map based on a magnetic survey has been

done and the results are shown in Figure 7.2.

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Figure 7.2: Local Geology

Agbaou Gold Project

Local Geology

Project No.

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SECTION 8. DEPOSIT TYPES

The target deposit type being explored for is the mesothermal auriferous sulphide (pyrite

+pyrrhotite) and quartz vein style mineralization. The gold mineralization within Agbaou

deposit is hosted within a specific quartz vein type that occurs along a broad area and can

be characterized by a wide range of quartz-vein types, brecciation, boudinage, sericitic and

carbonate alteration.

The mineralized quartz veins have a very distinctive texture that has been described as

“mottled” (Figure 8.1). Gold mineralization is also associated with variable amounts of

sulphides, mainly pyrrhotite and pyrite. These veins are easily identifiable in the diamond

drilling core intersections from the fresh rock below the saprolite/fresh rock boundary.

Figure 8.1: Mottled Quartz Vein in Drill Core in Fresh Rock

Agbaou Gold Project Mottled Quartz Vein Material

Project No.

439430

A down-hole stereoscopic structural analysis (Terratec Borehole Scanning) of the lithological

layering, foliation, fractures and veins indicate that all these features have a sub-parallel

attitude. The results of the Borehole Scanning analysis are illustrated in Figure 8.2. This

illustrates the sub-parallel nature of all structural features with a predominant north-east to

south-west strike, dipping to the south-east. The fractures and veins appear to have a

shallower average dip than the lithology and foliation. For

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Figure 8.2: Results of the Down-hole Stereoscopic Structural Analysis

Agbaou Gold Project

Structural Analysis

Project No.

439430

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The gold mineralization at Agbaou has been overprinted by strongly developed saprolitic

weathering. Within the saprolite zone even the quartz veins are highly weathered and all

sulphides replaced by iron oxides, seen as limonite stains. However gold was retained in situ

and does not appear to have been redistributed. This is illustrated in reverse circulation

chips and fines as preserved on chip boards that are kept for reference purposes (Figure

8.3) and in the remaining half core of diamond drilling core (Figure 8.4).

Figure 8.3: Chipboard from RC Drilling Intersection of a Mineralized Quartz Vein

Agbaou Gold Project

Quartz Vein Material

Project No.

439430

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Figure 8.4: Mineralized Quartz Vein Intersected by Diamond Drilling in Saprolite

Agbaou Gold Project Quartz Vein Intersection in Saprolite

Project No.

439430

During a paleo weathering event, most probably related to the formation of a paleo land

surface, a ferricrete / laterite cap developed in the soil profile overlying the gold deposits.

Gold derived from the saprolite zone was redistributed in the laterite, but largely retained in

close proximity to the underlying mineralized zones. Subsequent uplift resulted in the most

recent erosion with valleys incised into the earlier land surface. Erosion of the laterite cap

resulted in the redistribution of material derived from the laterite cap and the concentration of

gold in the local streambeds. Gold bearing laterite has been identified in many drill hole

intersections of this zone and the corresponding mineralized zone is delineated

independently of the underlying mineralized zones in the saprolite.

The secondary concentration of gold in the laterite zone is also controlling the gold

distribution observed in soil samples as is shown in Figure 9.1.

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SECTION 9. EXPLORATION

The evaluation by ENDEAVOUR (as Etruscan) of the Agbaou project area began in 2003

with the acquisition of the Agbaou permit. Exploration has been carried out under the

supervision of technically qualified personnel applying standard industry approaches.

Geochemical data quality has routinely been assessed as part of ongoing exploration

procedures. All data acquired meets or exceeds industry standards. All exploration work has

been carried out by, or supervised by technical personnel of the operator (e.g.

ENDEAVOUR, Etruscan, BHP, etc). Consultants and contractors have been engaged by

ENDEAVOUR (or Etruscan) for various activities including; sample collection, drilling and

assaying.

ENDEAVOUR (as Etruscan) conducted detailed (5,911 samples) and regional (1,831

samples) soil geochemical surveys which identified the gold mineralization at Agbaou,

Agbaou South, Mbazo, Zehiri and Niafouta (Figure 9.1).

Detailed pit sampling was completed to aid in the definition of the lateritic cap resource

which was identified and modelled in earlier campaigns. The pitting program covered an

area of approximately 400m2 on all sides of the deposits. The results of this program

indicated anomalous gold values in the laterite along the west flank of the Main deposit, and

to a lesser extent some anomalies were observed along the east flank of the Main deposit.

Pit samples were collected from “hand-dug” pits, 0.2 to 7.5 meters deep, depth being

controlled by the base of the laterite. Pits were systematically completed on a 50 meter by

75 meter grid over the Main and South mineralized zones. Samples are collected from

narrow channels chipped in the floor and wall of the pit and comprise a maximum of 1.5

meters of channel length. Samples were assayed using 50 gram charge fire assay gold

analyses completed by Transworld of Tarkwa, Ghana.

Some 876 Pits (4,056 samples) and 4 trenches (924 samples) were dug to explore the

laterite resource. The pit and trench sampling and analyses were not subjected to the same

QAQC protocol as was used for the drilling samples and as a result this data is not used in

the resource estimation.

During 2008 further studies were conducted to examine the lithology and the structural

geology of the deposit.

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Figure 9.1: Distribution of Gold in Soil Samples

Agbaou Gold Project Soil Geochemical Anomalies in the Original

License Area

Project No.

123456

9.1 SRK COMMENTS The currently defined Mineral Resource coincides with the areas where the best

mineralization was detected in soil samples as can be seen in Figure 9.1. However, recent

sterilization drilling has indicated that the potential exists to discover more mineralization

outside the currently modeled areas, and also in those areas not explored during the soil

surveys.

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SECTION 10. DRILLING

10.1 INTRODUCTION Drilling targets in the license area were identified during the soil surveys and were confirmed

by pitting and trenching. These targets were subsequently subjected to a systematic drilling

program, decreasing the between hole spacing following the identification of the mineralized

zones. A summary of drilling and sampling on the Agbaou Project is provided in Table 10.1.

The drilling methods used are reverse circulation (“RC”), diamond drill coring (“diamond”),

and a combination referred to as diamond tails, where the upper part of the hole is drilled by

RC and thereafter by diamond drilling.

Table 10.1: Summary of Drilling and Sampling at the Agbaou Gold Project Campaign/Company Contractor Type Number Metres Range

BHP 1988 - 1994 - Diamond 8 1680 BC-01 to BC-08

Goldivoire 1997 - 1999 - RC 175 20537 ARC001 to ARC175D

Goldivoire 1997 – 1999

- Diamond Tails

25

1535

ARC002D, ARC016D, ARC037D,








ARC175D

Etruscan 2005 Foraco RC 42 4572 ARC176 to ARC217D

Etruscan 2005 Foraco Diamond Tails 13 649 ARC194D, ARC196D, ARC197D,




ARC217D

Etruscan 2006 - 2007 Boart

Longyear

Diamond 166 22409 ADD218 to ADD383G

ENDEAVOUR 2010 -

2011

Global

Exploration

Services

RC 85 7062 AGBRC421 to AGBRC505

Total 514 58444

Drilling at the Agbaou Gold Project has been managed by both company and contract

geologists since ERCI acquired the project. Successive drilling programs were designed to

increase confidence of grade and continuity within the mineralisation envelopes.

Between 1988 and 1994, BHP drilled 8 diamond drillholes for a total of 1,680m. Goldivoire

(under various ownerships) subsequently drilled 175 drillholes (20,537.4m RC and 1,535.4m

diamond) between 1997 and 1999.

ERCI has completed three drilling campaigns since acquiring the property in 2003. Table

10.1 outlines the drilling prior to the cut-off date of 28th November 2011 for this study (Coffey,

2008).

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10.2 DRILLING PROCEDURES

10.2.1 Drillhole Collar Location Collar surveys were originally completed using an Ashtech (THALES) differential global

positioning system (“DGPS”) and co-ordinates reported in WGS84, UTM Zone 30 North. In

order to improve the accuracy all drill collars have subsequently been independently

surveyed by Envitech using total-station survey techniques. Envitech conducted a traverse

from the nearest control point (RGIR 13) to the Agbaou camp base station. The control point

was provided by the National Cartography Center, and is located in Tiassale some 105km

south-east of Agbaou. The base station was used to set up two field stations within the limits

of the deposit, ST1 and ST2, from which all survey points were measured. A LEICA total

station instrument, with an accuracy of 2mm – 5mm, was used by Envitech. Both UTM and

local grid co-ordinates were recorded. When the previous DGPS coordinates are compared

to total station survey data, some substantial differences are noted. This results in

differences of -9.7m to +8.2m in the easting and northing and +12m to +22.3m in the

elevation. While the total station coordinates are considered extremely accurate in a relative

sense, the differential between it and the previous DGPS data has not been adequately

explained.

10.2.2 Down-hole Surveying Procedures All drillholes completed by ERCI were down-hole surveyed using either a Terratec or a

Flexit© down-hole instrument at a minimum of every 30m and measured relative to magnetic

North. These measurements have been converted from magnetic to UTM Zone 30 North

values. The factor used to convert between the two grids is -6 degrees. No significant hole

deviation is evident in plan or section although a very unusual path has been captured in the

database for one of the holes drilled during an earlier campaign namely Drillhole ARC016D.

10.2.3 Reverse Circulation Procedures The initial ERCI supervised Reverse Circulation (“RC”) drilling was completed by Foraco

International of France using a Reska 30 truck mounted rig. RC rods were 4½ inch diameter

and the drill bit used was a standard 5½ inch diameter. During the last phase of drilling a 450

Schramm rig was used. The Schramm rig has a 350/900 compressor and 4½ inch rods and

the hammer bits are between 5 and 5¼ inch diameter. The rig is capable of drilling to 200

meters depth.

10.2.4 Diamond Drilling Procedures ERCI supervised diamond drilling was completed by Foraco International of France and

Boart Longyear of Canada. Foraco International used a Reska 30 truck mounted rig. Boart

Longyear used an SRS 850-03 skid mounted rig. HQ and NQ drilling was completed and the

core was oriented by a combination of the spear technique and both Flexit and Reflex multi

shot devices.

10.2.5 RC and Core Sampling Procedures The sampling procedures followed during RC and diamond drilling is detailed in Section 11

as is the sample quality assessment.

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10.2.6 Drilling Orientation Procedures The vast majority of drillholes in the Agbaou project were collared with an azimuth of

approximately 300° (UTM). A small number of holes were drilled towards an azimuth of

approximately 120°.

10.3 TRENCHING AND PITTING Trenching and pitting were used to confirm the soil geochemical anomalies, but since the

data does not meet the quality requirements for inclusion in the resources, this data was not

considered any further for estimation purposes.

10.4 DRILLING Both the results of RC and diamond drilling are used in the Mineral Resource estimation.

The increased density of drill hole intersections, with each successive campaign, is shown in

Figure 10.1 to 10-4. The drilling, relative to the latest outline of the main pit, during the first

two campaigns by BHP and Goldivoire is shown in Figure 10.1. Figure 10.2 shows the collar

positions of infill holes drilled during two campaigns in 2005 and in the period 2006-2008 by

ERCI. Infill drilling and sterilization drilling done in the period 2010 to 2011 by ENDEAVOUR

is shown in Figure 10.3.

10.5 DRILLING PATTERN AND DENSITY The final density of drilling achieved in the area of the Main and South deposits and

enclosed in the pit outlines is shown in Figure 10.4. The effect of the drilling density on the

resource classification is discussed in Section 14.

10.6 SRK COMMENTS Considering the statistical and geostatistical analysis alone, for the purposes of assessing

the density of drilling with regards to Mineral Resource confidence classification, the areas of

dense drilling (approximately 20m grid) are typically on a significantly tighter grid than the

ranges of the semi-variograms modeled for most zones. Infill drilling has resulted in an

increased classification of resource in the Main and South zones.

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Figure 10.1: Distribution of Drilling by BHP and Goldivoire

Agbaou Gold Project Drill Hole Collar Positions

Project No.

439430

253020

E

253020

E

253520

E

253520

E

254020

E

254020

E

673600 N 673600 N

674100 N 674100 N

674600 N 674600 N

0 100 200 300

[ABSENT]

[DDH]

[RC]

[RC/DDH]

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Figure 10.2: Distribution of Drilling by ERCI (2005 to 2008)


Project No.

439430

253020

E

253020

E

253520

E

253520

E

254020

E

254020

E

673400 N 673400 N

673900 N 673900 N

674400 N 674400 N

674900 N 674900 N

0 100 200 300

[ABSENT]

[DDH]

[RC]

[RC/DDH]

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Figure 10.3: Distribution of Drilling by ENDEAVOUR (2010 to 2011)


Project No.

439430

252100

E

252100 E

253100

E

253100 E

254100

E

254100 E

674000 N 674000 N

675000 N 675000 N

0 150 300

[ABSENT]

[DDH]

[RC]

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Figure 10.4: Final Distribution of Drilling Relative to the Designed Main Pit Outline


Project No.

439430

253020 E

253020

E

253520 E

253520

E

254020 E

254020

E

673600 N 673600 N

674100 N 674100 N

674600 N 674600 N

0 100 200 300

[ABSENT]

[DDH]

[RC]

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SECTION 11. SAMPLE PREPARATION, ANALYSES AND SECURITY

Only methods applied by ERCI are considered in this section. Documentation of previous

sampling methodologies prior to ERCI’s involvement is not available. The sampling method

has not changed since ENDEAVOUR became the owner of the exploration permit.

11.1 SAMPLING AND LOGGING PROCEDURES Only RC holes were drilled during the most recent drilling campaigns and SRK provides a

description of the sampling and logging procedures used by ENDEAVOUR for that program.

Diamond drilling completed prior to the current study was described by Coffey (2008) and

although it was not observed by SRK it is repeated here.

11.1.1 Sampling and Logging for Reverse Circulation Drilling During the last campaign, RC drill chips were collected as 1m intervals down-hole via a

cyclone into PVC bags, and then weighed prior to splitting. The collected samples were riffle

split using a three tier Jones riffle splitter. A final sample of approximately 2kg was collected

for submission to the laboratory for analysis. The splitter and boxes were cleaned with

compressed air between samples. If the sample was wet, the entire sample was placed in a

large rice bag and allowed to dry in the core shed before the weight was recorded and the

sample was split off. The geologist records the sample number, total weight, if water was

present, and a number of characteristics of the cuttings on a log sheet. RC drilling completed

by ENDEAVOUR has been geologically logged during the drilling program; however the

logging methodology was reviewed at the end of the previous drilling campaign with the

identification of the important association of mottled quartz with the gold mineralization. A re-

logging program was completed and the database updated. Drill cuttings from each sample

interval have been glued onto an A3 size chip-board as a permanent record. The chip

boards are stored at the ENDEAVOUR exploration camp in Agbaou.

11.1.2 Sampling and Logging for Diamond Drilling ERCI supervised diamond drilling was completed by Foraco International of France and

Boart Longyear of Canada. The following description of the diamond drilling sampling is an

excerpt from the Coffey (2008) report:

All drillholes were collared with HQ size single tube core (77.8mm) and once competent rock

is encountered the hole was reduced to NQ size core (60.3mm).

The current core handling procedures are as follows:

• Core laid out into core boxes on the drill site. • Core transported from the drill site, in a specifically designed core tray transport

device, to the logging yard where core is laid out in sequence. • Core marked–up, geotechnically logged, recoveries and RQD’s measured. (RQD’s

are not recorded in the saprolitic segments of the core). • Core geologically logged. • Core photographed with a digital camera.

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• Core marked up for cutting and cut with a conventional diamond core saw. Strongly oxidised core is split using a spatula.

• Bulk density samples are selected and determinations performed on site. Previously, samples were dispatched to TWL in Ghana for the measurements of bulk densities.

Industry standard core sampling methods are used throughout. The orientation mark is

accurately transferred to the side of the core, representing the bottom of the core. The core

is then matched up in the core trays and the orientation line propagated along the length of

the core. The core is then marked off for sampling on a geological basis or to a maximum of

1.5m sample intervals.

Core sampling is undertaken using a diamond saw to cut the core lengthways, ensuring that

the orientation line is preserved on the half of the core that is retained for reference. When

no orientation line can be drawn, a cut line is drawn. Core orientation stubs are clearly

marked and not sampled unless absolutely necessary. The core samples are bagged and

tagged ready for sample preparation as per a prepared cut sheet which is generated from

excel. For best practice, Etruscan endeavors to sample the same half core throughout all

holes.

All core is logged wet by experienced geologists using a simple and consistent code system.

In addition, basic structural and geotechnical information is routinely collected, as this will

reduce the scope of specialist geotechnical consultants and the number of dedicated

geotechnical drillholes, as well as enhancing the quality of the geological interpretations.

However, the quality of the spear generated orientations is not high, and this affects the

quality and the quantity of structural data that can be recorded from such orientated core.

Core is photographed when wet and when dry, two trays at the time.

Upon arrival at the core processing facility, the core is matched up, oriented, logged and

recoveries calculated prior to marking off for sampling and cutting.

11.2 SAMPLE RECOVERY Sample recovery for RC drilling was noted as very good and averages approximately 23.3kg

per metre drilled. Bulk sample weights have been systematically recorded for each metre

drilled. Sample recovery in diamond drill holes was very good although recoveries for core

from the moderate to highly weathered saprolite and highly fractured and brecciated zones

returned poor recoveries. ENDEAVOUR utilized HQ drilling to minimize the core loss in the

weathered zones.

11.3 SAMPLE QUALITY The sampling procedures adopted for drilling are consistent with current industry practise.

Samples collected by diamond drilling within the highly weathered zones are of moderate

quality, with the remainder being high quality. Sample recoveries and quality for the RC

drilling are high.

Coffey (2008) reports on nine sets of twinned holes, but presented data for only eight sets,

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as the data of one set are duplicated. In SRK’s opinion, 3 sets are not true twins as the drill

holes are separated too much in depth. None the less, only one set is rejected as the

repetition is less than 50% and two sets show about 60% agreement. Five sets display

greater than 80% agreement, indicating a strong correlation between RC and diamond

drilling intersections notwithstanding the fact that some of the holes deviated significantly

away from each other

It is difficult to determine if any negative bias has resulted in the diamond drilling due to the

use of water. A number of the diamond drill holes had poor recovery in the highly weathered

zone and there exists potential to wash out the fine gold associated with the fractures and

veining and therefore underestimate the gold content. The proportion of diamond drilling in

the weathered zones is, however, low and the potential for underestimation associated with

this type of negative bias are considered commensurately low.

RC field duplicate samples are routinely collected to allow assessment of the field sampling

error (or bias) once the laboratory error, determined from analysis of pulp duplicates, has

been subtracted. Acceptable reproducibility has been identified during an assessment of RC

field duplicate data (Section 12.1) generated and no distinct bias is evident.

During the latest RC drilling campaign all sampling was done by ENDEAVOUR staff. The

samples were sun dried, bagged, sealed and submitted together with the independent

Quality Assurance and Quality Control (“QAQC”) samples to the Bureau Veritas Mineral

Laboratory Côte d’Ivoire (“BV”), in Abidjan.

At the BV Laboratory the samples are sorted and entered into the laboratory control system.

Sample preparation at the BV Laboratory entails:

• Drying in an oven at 100ºC to 110ºC for 12 hours; • Crushed in a Jaw Crusher to 90% less than 2mm; • The entire sample is then milled in an LM5 mill to 85% less or equal to 75

micrometres; • The gold in the sample is then analyzed by the fire assay method:

o The fire assay uses 40 g of the pulverised sample, fused with fluxing reagents at 1100ºC for one hour;

o The lead button is then separated from the slag; o Cupellation of the sample is done at 990ºC for 45min – 50min in a muffle

furnace; o The silver-gold prill is collected and then digested in aqua regia; o The finish is done by Atomic Absorption Spectroscopy (“AAS”).

SRK visited the BV Laboratory in Abidjan during October 2011 and at that time the

laboratory was not fully accredited in terms of ISO 17025 for the gold analyses, although

parts of the sample preparation was already accredited.

11.4 SPECIFIC GRAVITY DATA All densities used by SRK are based on core from the diamond drilling campaigns. Coffey

(2008) reported as follows on the density measurements. A total of 654 bulk density

determinations have been collected from the Agbaou deposit. A total of 277 determinations

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have been completed by TWL Laboratories, 363 were completed by ENDEAVOUR (of

these, some have been also determined by TWL) and 48 were completed prior to Etruscan’s

involvement. Only determinations completed since Etruscan’s involvement are considered.

All readings were taken over a range of lithological and weathered profiles.

The procedure used by both ENDEAVOUR and TWL is detailed below and is based on the

Archimedes Principle as follows:

• 10cm billet of clean core is weighed (note the core is not oven dried prior to bulk density determination);

• Core is immersed in paraffin wax then reweighed to establish weight of the wax; • Core is then suspended and weighed in water to determine the volume; • The Bulk Density is then calculated using the equation

Bulk Density core = [Mass core] / [(Mass air – Mass water) – (Mass wax / 0.99)].

Only bulk density determinations carried out by ENDEAVOUR (363) were used for the

resource estimation. Results are presented in Table 11.1.

Table 11.1: Average Densities Applied per Weathering Zone in the Tonnage Calculation Weathering Zone Number of measurements Average Value

Laterite 19.0 2.10

Saprolite 274 1.66

Transitional 127 2.00

Fresh 207 2.65

Coffey (2008) recommended that the core should be oven dried to a maximum of 60ºC for a

6 hour period prior to bulk density determination. ENDEAVOUR is of the opinion that the

difference in density between the air dried and oven dried samples would not be significant.

SRK accepts the ENDEAVOUR’s opinion but recommends that when the opportunity

presents itself again the density of especially the saprolitic and lateritic material be tested

before and after oven drying.

11.5 QUALITY ASSURANCE AND QUALITY CONTROL PROGRAMS Coffey (2008) provided a comprehensive review of the QAQC measures employed prior to

the last drilling campaign, and accepted the values in the database and concluded as

follows:

• Use of Certified Standard Reference material has shown a range of relative bias for

TWL, Ghana, which varies between -8.9% to +5.6%. This requires monitoring on an

on-going basis to determine the reason for the range of bias;

• Repeat analyses have confirmed that the precision of sampling and assaying is

within acceptable limits for sampling of gold deposits;

• Umpire assaying at SGS Lakefield, Canada returned a significantly lower mean

grade than the original assaying. This is based on one batch and should be

monitored on an on-going basis to identify the reason why;

• The use of Certified Standard Reference material as blank material is recommended

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at Agbaou.

In Section 12, SRK will review the results of an analysis of the QAQC data for the 2010-2011

drilling campaign.

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SECTION 12. DATA VERIFICATION

During the current exercise no additional samples were collected by SRK for verification

purposes. In the following paragraphs SRK will review the results of an analysis of the

QAQC data for the last drilling campaign.

12.1 ENDEAVOUR FIELD DUPLICATES During the 2010-2011 RC drilling campaign at the Agbaou Gold Project, coarse material

duplicate samples were submitted to the BV laboratory as blind field duplicates. SRK

reviewed the results reported and found a significant scatter between individual duplicate

samples but an overall similarity and no significant bias. In order to explore the scatter the

following subsets of data were investigated:

• The dataset without obvious outliers; • The dataset without the outliers and without samples that returned gold contents of

less than 0.1 g/t gold.

The results of these repeated investigations are summarised in Table 12.1 below.

Table 12.1: Summary Statistics for Field Duplicate Samples.

Subset Number of

Samples

Mean for

Primary

Samples

Mean for

Duplicate

Samples

Error

RMSE%

Full dataset 205 0.87 0.88 73%

Less Outliers 203 0.82 0.81 65%

Greater Than 0.1g/t 75.0 2.08 2.08 41%

In this table the error, as measured by the Root Mean Square Error Percentage (“RMSE %”),

is an indication of the scatter and decreases with each successive exclusion.

Table 12.2 summarises the descriptive statistics for the final dataset, which excludes the two

high grade outliers and the samples with grades less than 0.1g/t gold.

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Table 12.2: Summary Statistics for Field Duplicate Samples Regression statistics

Correlation Coefficient: 0.975

Pearson Product Moment Correlation Coefficient: 0.975

Slope of RMA Line (Reduced Major Axis): 0.958

Error on the slope of RMA Line: 0.025

Rank Correlation Coefficient: 0.932

Univariate Statistics

Original Assay Duplicate Assay

Count 75 75

Arithmetic Mean 2.083 2.082

Minimum 0.1 0.1

Maximum 28.75 26.94

Standard Deviation 3.80186 3.64225

Coefficient of Variation 1.82489 1.74962

Notwithstanding the good general correlation between the original and duplicate sample sets

the scatter is high as is well illustrated by the scatter plot (Figure 12.1) and in the Half

Absolute Relative Difference (H.A.R.D) plot (Figure 12.2).

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Figure 12.1: Scatter Plot of Field Duplicates in Reduced Dataset

Agbaou Gold Project

Field Duplicate Samples

Project No.

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Ass

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Original Assay

Scatter Plot of Au

Au_Scatter

RMA Line

Ideal Correlation

Upper 10% Limit

Lower 10% Limit

Slope = 0.958

Y axis Intercept = 0.086

Error on slope = 0.025

Error on Y axis Intercept = 0.107

y = 0.958x +0.086

Corellation co-efficient = 0.975

Number of Pairs = 75

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Figure 12.2: H.A.R.D Plot of Field Duplicates in Reduced Dataset

Agbaou Gold Project

Field Duplicate Samples

Project No.

439430

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2

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12.2 CERTIFIED REFERENCE MATERIALS Samples of six Certified Reference Materials (“CRM’s”) were added as blind samples by

ENDEAVOUR to the batches of samples submitted to the BV Laboratory during the last

campaign (2010 to 2011). The results of the analyses of these CRM’s are summarised in

Table 12.3.

Table 12.3: Summary Statistics of CRM’s Analysed by BV Laboratory

Standard

Certified

Value

No of

Analyses

Minimum

Maximum

Mean

% Diff

Number

Outside

Range

% Out

OxA71 0.085 6 0.07 0.08 0.08 11.76 0 0

OxF68 0.805 52 0.77 0.84 0.80 0.00 0 0

OxI54 1.868 37 1.74 1.97 1.83 1.61 0 0

OxI67 1.817 35 1.77 1.89 1.84 1.10 0 0

OxJ67 2.366 32 2.19 2.46 2.36 0.42 0 0

OxK79 3.532 34 3.50 3.62 3.57 1.13 0 0

As can be expected the analytical error is significantly larger for the low grade CRM (OxA71)

although the results are still acceptable. This provides additional support for the decision in

Section 12.1 above to consider only coarse field duplicates with grades in excess of 0.1g/t

gold. The very good results obtained for the other CRM’s provides confidence in the BV

Laboratory although they are not accredited in terms of ISO 17025.

12.3 BLANKS ENDEAVOUR submitted blank material to the BV Laboratory as part of the QAQC protocol.

When SRK visited the field operations, sand obtained from a granitic terrain distant from the

project were used as a blank. In addition following the recommendation by Coffey (2008)

certified blank material was also inserted in the sample batches. The results of the analyses

of the blank material are summarized in Table 12.4.

Table 12.4: Summary Statistics of Blank Materials Analysed by BV Laboratory Material No of

analyses Minimum Maximum Mean >0.05 g/t Au %

Field Blank 212 0.005 0.12 0.007 1 0.5

Blank Standard 34B 11 0.005 0.005 0.005 0 0

Only one incident of potential contamination was reported (0.5% of blanks submitted) and

this sample returned a value of 0.12g/t. The field blank in SRK’s opinion provides a better

indication of contamination during the milling stage and are therefore a better indication of

between sample contaminations than the certified blank material. The data is therefore

accepted as no significant contamination was detected.

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12.4 UMPIRE LABORATORY RESULTS Because the BV Laboratory is not ISO 17025 accredited a batch of duplicate pulp samples

were submitted to an unaccredited SGS Lakefield (“SGS”) laboratory in Tarkwa, Ghana

(Referee Laboratory 1). The results of this umpire laboratory comparison will be discussed

below. Following this exercise pulp duplicates were also submitted to an accredited

laboratory, SGS Canada in Lakefield, Ontario (Referee Laboratory 2). The results of this

second umpire laboratory will also be discussed in the subsequent sections.

12.4.1 Duplicate Samples Referee Laboratory 1 The results of a review of the results are summarised in the Table 12.5. The dataset

(excluding QAQC samples) was investigated and two additional subsets were created. The

first of these excludes one outlier that may indicate an entry mistake. The second subset, in

addition, excludes all samples that contain less than 0.1 g/t gold as determined by either of

the laboratories. The statistical comparison of the two laboratories in the reduced dataset is

given in Table 12.6.

Table 12.5: Summary Statistics for Umpire Laboratory Duplicate Samples

Subset Number of

Samples BV Mean

SGS

Mean

∑

Differences Error RMSE%

Full Dataset 347 0.81 0.74 23.31 150.83

Less Outlier 346 0.75 0.74 2.43 63.7

Greater / Equal to 0.1 135 1.87 1.81 6.99 39.25

Table 12.6: Descriptive Statistics of Final Dataset of Umpire Laboratory Duplicates Regression statistics

Correlation Coefficient: 0.958

Pearson Product Moment Correlation Coefficient: 0.958

Slope of RMA Line (Reduced Major Axis): 1

Error on the Slope of RMA Line: 0.025

Rank Correlation Coefficient: 0.879


Original Assay Duplicate Assay

Count 135 135


Minimum 0.1 0.1

Maximum 13.31 14.1



Even after the removal of the low-grade samples the analytical scatter is still very high as

can be seen in the RMSE%, the Scatter plot (Figure 12.3) and the H.A.R.D plot (Figure

12.4). The Quartile – Quartile (“QQ”) plot (Figure 12.5) indicates a slight bias in the range

3g/t to 7g/t gold and this is also reflected in the positive sum of differences (Table 12.5). The

bias for the entire dataset is insignificant as can be seen in the slope and intercept of the

regression line (Figure 12.1 and Table 12.6).

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Figure 12.3: Scatter Plot of Umpire Laboratory Duplicate Analyses in the Reduced Dataset

Agbaou Gold Project Umpire Laboratory Analyses

Project No.

439430

0.00

2.00

4.00

6.00

8.00

10.00

12.00

14.00

16.00

18.00

0.00 2.00 4.00 6.00 8.00 10.00 12.00 14.00 16.00 18.00

Du

pli

cate

Ass

ay

Original Assay

Scatter Plot of Au excl OL

Au excl OL_Scatter

RMA Line

Ideal Correlation

Upper 10% Limit

Lower 10% Limit

Slope = 0.989

Y axis Intercept = 0.001



y = 0.989x +0.001



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Figure 12.4: H.A.R.D Plot of Umpire Laboratory Duplicate Analyses in the Reduced Dataset


Project No.

439430

0

0.5

1

1.5

2

2.5

3

3.5

0%

10%

20%

30%

40%

50%

60%

70%

80%

90%

100%

0% 10% 20% 30% 40% 50% 60% 70% 80% 90% 100%

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Figure 12.5: QQ Plot of Umpire Laboratory Duplicate Analyses in the Reduced Dataset


Project No.

439430

0

2

4

6

8

10

12

14

0 2 4 6 8 10 12 14

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Original Assay

QQ Plot of Au Data

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12.4.2 Certified Reference Materials Submitted to SGS Ghana ENDEAVOUR submitted a set of CRM’s as blind samples within the batch that was

submitted to SGS Ghana. The results of these analyses are summarized in Table 12.7.

Unfortunately an inadequate number of any specific CRM was submitted to justify a

statistical assessment of the quality of the SGS analyses. The CRM results received meets

the certified acceptance criteria.

Table 12.7: Summary of CRM Results Submitted to Referee Laboratory 1 CRM No analysed Certified Values Measured Range

Minimum Maximum

OxA 71 2 0.080 0.07 0.08

OxF65 4 0.805 0.77 0.80

OxI54 6 1.868 1.75 1.97

OxI67 1 1.817 1.78 -

OxJ64 6 2.366 2.30 2.42

OxK79 1 3.532 3.60 -

12.4.3 Duplicate Samples Referee Laboratory 2 A second batch of pulp sample duplicates were specifically selected to span the range of

gold values seen in the Agbaou samples. The objective with this was to verify the results of

the primary laboratory and not to inspect the quality of the SGS laboratory in Ghana

(Referee Laboratory 1). The results obtained from this comparison between the primary

laboratory (BV laboratory in Abidjan) and Referee Laboratory 2 (SGS Canada) are

summarized in Table 12.8. The dataset (excluding QAQC samples) was investigated and an

additional subset was created to exclude the very high grade samples where physical

particulate gold distribution may play a role. The statistical comparison of the two

laboratories in the reduced dataset is given in Table 12.6.

Table 12.8: Summary Statistics for Referee Laboratory 2 Duplicate Pulp Samples Subset Number of

Samples BV Mean

Ref Lab 2

Mean

∑

Differences

Error

RMSE%

Full dataset 190 4.38 4.01 70.19 122.2

Less than 20 g/t 181 2.08 1.96 21.14 43.43

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Table 12.9: Descriptive Statistics of Referee Laboratory 2 Duplicate Pulp Samples (<20g/t) Regression statistics

Correlation Coefficient 0.958

Pearson Product Moment Correlation Coefficient 0.958

Slope of RMA Line (Reduced Major Axis) 0.927

Error on the Slope of RMA Line 0.020

Rank Correlation Coefficient 0.966


Original Assay Ref Lab 2

Count 181 181


Minimum <.01 0.1

Maximum 18.35 16.10



The scatter plot showing the results of the two laboratories is given in Figure 12.3. The

results show approximately 10 (5 %) outliers of variable severity. The analytical scatter is still

very high as can also be seen in the RMSE% and the H.A.R.D plot (Figure 12.4). The

Quartile – Quartile (“QQ”) plot (Figure 12.5) indicates a slight bias in the range 4 g/t to 8 g/t

gold and this is also reflected in the positive sum of differences (Table 12.5).

The bias for the entire dataset is regarded as insignificant and support for this is also evident

when the slope and intercept of the regression line are considered (Figure 12.1, Table 12.6).

12.4.4 Certified Reference Materials Submitted to SGS Canada ENDEAVOUR submitted CRM’s as blind samples with the batch that was submitted to SGS

Canada (Referee Laboratory 2). The results of these analyses are summarized in Table

12.7. Unfortunately an inadequate number of any specific CRM was submitted to justify a

statistical assessment of the quality of the SGS analyses. The CRM results received meets

the certified acceptance criteria although the measured range indicates a bias with the SGS

results consistently lower than the Certified Value, which confirms the bias noted above

although it also appears in values lower than 4g/t Au.

Table 12.10: Summary of CRM Results Submitted to Referee Laboratory 2 CRM No analysed Certified Values Measured Range

Minimum Maximum

OxJ64 3 3.562 3.13 3.49

OxI54 3 1.807 1.69 1.80

STD34B 1 0.613 0.59 For

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Figure 12.6: Scatter Plot of the Original Analyses Versus the Results of Referee Laboratory 2

Agbaou Gold Project

Comparison with Referee Laboratory 2

Project No.

439430

-20

0

20

40

60

80

100

120

140

160

180

0 20 40 60 80 100 120 140 160 180

Du

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Ass

ay

Original Assay

Scatter Plot of Au

Au_Scatter

RMA Line

Ideal Correlation

Upper 10% Limit

Lower 10% Limit

Slope = 1.064

Y axis Intercept = -0.65



y = 1.064x -0.65



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Figure 12.7: H.A.R.D Plot of the Original Analyses Versus the Results of Referee Laboratory 2

Agbaou Gold Project

Original and Referee Laboratory 2 Comparison

Project No.

439430

0

0.5

1

1.5

2

2.5

3

3.5

4

0%

10%

20%

30%

40%

50%

60%

70%

80%

90%

0% 10% 20% 30% 40% 50% 60% 70% 80% 90% 100%

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Figure 12.8: QQ Plot of Original and Referee Laboratory 2 Results (<20 g/t Au)

Agbaou Gold Project

Original and Referee Laboratory 2 Comparison

Project No.

439430

0

5

10

15

20

0 5 10 15 20

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Original Assay

QQ Plot of Au<20ppm Data

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12.5 SRK COMMENTS In the opinion of SRK the sampling preparation, security and analytical procedures used by

ENDEAVOUR are consistent with generally accepted industry best practices and are

therefore adequate.

The apparent coarse nature of the gold results in a high variability in the field duplicate set.

The BV laboratory results of the blind CRM’s inserted by ENDEAVOUR in the field returned

very good results. Similarly blank material returned acceptable results and SRK accepts the

BV Laboratory results.

The relatively poor results of the duplicate pulp samples submitted to the SGS laboratory in

Ghana (Referee Laboratory 1) and to the accredited SGS Laboratory in Canada (Referee

Laboratory 2) indicate poor but acceptable replication. It has to be concluded that the scatter

observed although less in the second referee laboratory, is partly caused by the particle

distribution of the gold, although the same should have been evident in the pulp duplicates

analyzed by the BV laboratory if this was the only cause. The results of the second set of

replicate pulp analyses (Referee Laboratory 2) are however very similar to that of the

independent field coarse duplicates and a slight improvement on the results from the

unaccredited laboratory in Ghana (Referee Laboratory 1).


and accuracy of the sampling and assaying, is adequate and the BV laboratory returned


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SECTION 13. MINERAL PROCESSING AND METALLURGICAL

TESTING

13.1 BACKGROUND Comminution and recovery metallurgical tests were performed on ore samples from the

Agbaou deposit and the results used to develop the process flowsheet are summarized in

this section. Initial feasibility testwork was carried out in 2007 by Mintek under the

supervision of MDM.

SENET reviewed the results from the initial feasibility study and together with ENDEAVOUR,

proposed a metallurgical test program whose results would support the process flowsheet

selection required for the EOS. Various laboratories and or vendors were selected to carry

out these tests, namely:

• Maelgwyn Mineral Services (MMS) - Aachen (high shear reactor-HSR) testwork, gravity and leach tests, cyanide detox testwork;

• Knelson Africa – Gravity (GRG) testwork; • Mintek - Comminution testwork - note that all the samples received were too fine to

carry out any comminution test and even though the comminution testwork was planned for, no comminution tests were conducted;

• Patterson and Cooke - Thickening and rheology testwork and viscosity modifier tests; • OMC - Interpretation of comminution results, design and modelling of the

comminution circuit. The initial feasibility results summarized in the review of the initial testwork section are

contained in the following reports:

• November 2007, Mintek, ˝Laboratory testwork to evaluate processing options for the Agbaou gold deposit” report number 4835;

• May 2008, JK. Tech,˝ Agbaou project” report number 07002; • January 2009, Mintek,” Comminution testwork on drill core samples from Agbaou

deposit” report number 5197; • January 2009, MDM,” Etruscan resources Agbaou gold mine-feasibility study”.

Flowsheet development and process design criteria used in the EOS utilized the results

obtained from tests conducted by various laboratories that were carried out under SENET

supervision. A list of the reports from these various laboratories is provided below and the

reports are available in the Appendices of the EOS (SRK Consulting, SENET, Knight Piésold

Consulting, 2012).

• August 2011, OMC “Agbaou Gold project: comminution testwork and circuit design review” Report number: 8819.30.

• December 2011, OMC “Agbaou Gold project: comminution testwork and circuit design review” Report number: 8857.30.

• December 2011, Patterson & Cooke “Endeavour Corporation’s Agbaou project: Slurry test report” Report number: SEN-AGB 8284 RO1 Rev 1.

• January 2012, BSFSA, “Bulk solids flow report” Report number BSFSA-535. • February 2012, Patterson & Cooke, “Bench–top thickening test report” Report

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number SEN-AGB 8284 RO3, Rev 1. • February 2012, Patterson & Cooke, “Slurry test report” Report number SEN-AGB

8284 RO2, Rev 1. • March 2012, Maelgwyn Mineral Services “Leachox and cyanide detox test on gold

bearing ore samples from Agbaou” Report number: 11/93. • January 2012, Knelson Concentrators Africa, “Gravity Recovery Testwork on

Samples from SENET” Report number SENET 01/12 January 2012.

A list of the samples and the design values selection are also provided in the EOS (SRK Consulting, SENET, Knight Piésold Consulting, 2012 Appendix 13-I and Appendix 13-J respectively).

13.2 REVIEW OF THE INITIAL FEASIBILITY METALLURGICAL TESTWORK This section gives a summary of the feasibility metallurgical results conducted by Mintek in

2009. Details of the origin of samples, test methods and detailed results are contained in the

feasibility study document produced by MDM in January 2009 entitled “Etruscan Resources

Agbaou Gold Mine-Feasibility Study”.

The results obtained from this testwork phase are summarized in Table 13.1:, Table 13.2:

and Table 13.3:.

Table 13.1: Head Assays

Units Saprolite Bedrock

South North South North

Ore grade g/t 1.88 1.69 1.09 3.67

Moisture content % 5 5 5 5

Specific gravity t/m3 1.69 1.69 2.79 2.83

Bulk density t/m3 1.01 1.01 1.67 1.7

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Table 13.2: Comminution Results Saprolite Bedrock


UCS

Min MPa - - 82.2 55.5

Max MPa - - 129.1 194.9

CWi kWh/t - - 10.4 14.16

BBWi(75µm)

Min kWh/t - - 9.99 9.11

Max kWh/t - - 11.79 14.94

BBWi (106µm)

Min kWh/t - - 10.43 9.44

Max kWh/t - - 12.54 15.22

BRWi

Min kWh/t - - 9.99 9.11

Max kWh/t - - 16.48 19.22

Ai

Min - - 0.0544 0.0729

Max - - 0.1729 0.1851

JK Drop Weight

A - - 50.7 58.7

B - - 0.46 0.39

A×B 23.32 22.89

Ta - - 0.40 0.46

Note – saprolite samples were too fine/ friable to allow comminution testing. Typical data was provided by OMC for modeling and circuit design purposes.

Bond Ball indices on the bedrock ore at a limiting screen size of 75µm varied from 9.11 to

14.94kWh/t. At a limiting screen size of 106µm, the BBWi varied from 9.44kWh/t to

15.22kWh/t. indicating that the bedrock ore can be classified in the soft to medium hard

category.

Abrasion indices ranged from 0.0544 to 0.1851 indicating that the bedrock ore has a low

abrasive tendency. Therefore liner wear and ball consumptions are not expected to be

significant.

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Table 13.3: Recovery Results Units

Saprolite Bedrock


Gravity Recovery

Gravity Recoverable Gold % 23.2 20 22.7 31.4

Mass Pull % 0.1 0.1 0.2 0.3

CIL on Gravity Tails

CIL Dissolution (No Oxygen) % 90 83.8 90.8 90.6

CIL Dissolution (With Oxygen) % 92.2 93.1 - -

Cyanide Consumption

No Oxygen kg/t 0.22 0.18 0.23 0.26

With Oxygen kg/t 0.24 0.18 - -

Lime Consumption

No Oxygen kg/t 3.01 2.97 0.52 0.53

With Oxygen kg/t 2.99 2.97 - -

13.3 LIMITATIONS OF THE INITIAL FEASIBILITY STUDY

• No materials handling testwork was conducted on the saprolite ore. The saprolite ore being viscous may present materials handling problems which could hinder production or lower gold recovery.

• No comminution tests were performed on the saprolite ore. • In the milling section, only one circuit option was modelled. The milling section is the

highest energy consumer in the entire plant and therefore warrants proper investigation in selecting a circuit configuration.

• No EGRG tests were performed with the accompanying simulation of the expected plant gravity gold recovery.

• No cyanide detox tests were carried out. • No viscosity modifier tests were performed in light of the saprolite ore being very

viscous. 13.4 COMMINUTION CIRCUIT REVIEW In 2011, OMC was requested to review the comminution results produced by Mintek (as part

of the initial feasibility study) and make recommendations for a suitable comminution circuit

to treat the Agbaou ore.

The tables below show a summary of the testwork results used by OMC in the circuit design.

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Table 13.4: Comminution Results Summary

Ore Type Units

Bed Rock

Saprolite ADD269-

N1U ADDN1L

ADD270-

N2U

ADD270-

N2L

ADD301-

S1U ADD303S2U

ADD303-

S2L

UCS

Average Mpa 97.7 110.8 107 86.4 97.6 104.8 108.2

Range Mpa 79-116 60.-162 41-173 70-103 79-116 95-115 92-125

Abrasion Index 0.110 0.082 0.096 0.174 0.169 0.060 0.133

Crushing Work Index

CWi kWh/t 7 9.8 6.9 6.6 7.5 6.5 7

Rod Mill Work Index

P80 mm 765 738 738 747 735 740 724

Wi kWh/t 19.2 15.7 12.9 13.7 15.6 16.5 15.0 4.5

Ball Mill Work Index

P80 mm 54 56 53 58 53 52 57

Wi kWh/t 14.9 12.3 9.1 11.0 10.0 11.1 11.8 3.5

Rwi : Bwi 1.29 1.27 1.42 1.25 1.56 1.48 1.27

SG

2.73 2.76

2.81 2.82

1.69

SMC

A

69.4 58.7

62.5 50.7

b

0.39 0.67

0.46 0.59

A× b

27.1 39.3

28.8 29.9

ta

0.46 0.53

0.5 0.4

Table 13.5: Viscosity & Shear Rate Summary

Unit Bed Rock Saprolite

North South North South

Viscosity cP/mPa.s 14.04 14.34 900.0 886.8

Shear rate (s-1

) 22.0 22.0 21.2 21.1

Following the review and analysis of the above results, OMC made the following

conclusions:

• There is significant variance in comminution parameters between the saprolite and

bedrock ores. The bedrock ore showed high competency but lower than average

resistance to grinding;

• The saprolite ore was seen to be of high viscosity. As such, all modelling of the

saprolite was carried out at 50% solids;

• The SABC circuit was considered inefficient from a power draw perspective;

• Design of the comminution circuit was primarily based on the competent bedrock ore.

Although the SABC circuit was seen to be inefficient from a power draw perspective, this

option was still selected based on the following:

• Power is relatively cheap in Cote d’Ivoire;

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• The SABC circuit has a simple mode of operation, with advantages on operability

and maintenance.

13.5 EOS METALLURGICAL TESTWORK SAMPLE SELECTION

13.5.1.1 Introduction

Following a review of the initial feasibility metallurgical test results, SENET together with

ENDEAVOUR proposed a test program and the aims were to:

• Look at areas where the plant design can be improved; • Conduct testwork required to satisfy a bankable feasibility study.

13.5.1.2 Samples Received and Sample Compositing

The following samples for the EOS were selected from the North and South deposits:

• Saprolite - South zone and North zone; • Bedrock - North zone. (no samples were available for South zone)

Saprolite South zone and North zone samples were combined to form one composite

saprolite sample for optimisation testwork.

Details of the received samples including depth, where the samples were taken, gold grades

and compositing included in the EOS (SRK Consulting, SENET, Knight Piésold Consulting,

2012 Appendix 13-I).

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13.6 METALLURGICAL TEST PROGRAM (EOS) Testwork conducted during the EOS is shown in Figure 13.1:. Note that no comminution

testwork was conducted on North and South saprolite zones as the samples received were

too fine for any comminution testwork. Therefore only the comminution results obtained for

the bedrock (North zone) in the initial feasibility study were used for the plant design. For

the saprolite ore, assumed values were used by OMC from their data base.

Figure 13.1: EOS Testwork Flowsheet

13.7 EOS METALLURGICAL RESULTS The following tests were completed:

• Head analysis; • Gravity; • Effect of air & sparging of oxygen; • Effect of preg-robbing; • Optimum leach conditions established; • Variability leach;

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• Thickening and rheology.

Details of these results are included in Appendices 13-A to 13-H listed in the background

and described in the sections to follow.

13.7.1 Head Analysis Full elemental analysis was conducted on the saprolite and bedrock composites. A summary

showing Au, Ag, As assays and the specific gravity (SG) is shown in the table below and

detailed results are included in the EOS (SRK Consulting, SENET, Knight Piésold

Consulting, 2012 Appendix 13-G).

Table 13.6: Head Assays

Saprolite Bedrock

Au(g/t) Ag (ppm) As(ppm) SG Au(g/t) Ag (ppm) As(%) SG

Composite 2.36 <1 111 2.79 2.99 <1 0.02 2.85

Variability

Max 3.65 <1 347 2.85 2.57 <1 0.02 2.99

Min 1.71 <1 39 2.75 1.62 <1 0.01 2.77

Average 2.61 <1 148 2.80 2.08 <1 0.02 2.89

P85 3.14 <1 186 2.85 2.45 <1 0.02 2.98

13.7.2 Gravity Recoverable Gold No EGRG testwork was performed because the samples received were too fine for EGRG

testwork. However, GRG tests were conducted and plant GRG recoveries were predicted as

shown in the table below.

Table 13.7: Gravity Gold Recovery

Saprolite Bedrock

Lab

Results

GRG 44.1 31.1

Mass pull (%) 0.73 0.85

Conc grade (g/t Au) 127.1 90.7

GRG

Modelling

GRG (%) – Predicted Plant 30 20

Mass pull (%) 0.02 to 0.03 0.02 to 0.03

13.7.3 High Shear Reactors (HSR) Testwork HSR testwork was conducted to evaluate the following potential benefits:

• Lower leach residence time;

• Improvement in gold recovery; • Reduction in reagent consumptions.

The table below includes a summary of the rate results with and without HSR.

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Table 13.8: Summary Results: Effect of using a HSR

Time

Bedrock Saprolite

Residue DO Reagent

Consumption Dissolution Residue DO

Reagent

Consumption Dissolution

Au NaCN CaO Solid Au NaCN CaO Solid

g/t mg/l kg/t kg/t Au % g/t mg/l kg/t kg/t Au %

HSR 2hr 0.92 8 0.09 0.72 63.72 0.20 32 0.04 3.95 83.37

HSR 4hr 0.82 9 0.10 0.69 67.63 0.16 28 0.02 3.88 86.24

HSR 6hr 0.54 8 0.15 0.69 79.33 0.10 28 0.03 3.96 91.34

HSR 8hr 0.39 8 0.16 0.72 85.15 0.10 25 0.05 3.97 91.25

HSR 12hr 0.29 7 0.24 0.69 88.62 0.11 24 0.08 3.99 91.13

HSR 16hr 0.25 6 0.24 0.75 90.30 0.09 21 0.12 4.16 92.19

HSR 24hr 0.19 6 0.31 0.74 92.52 0.06 18 0.15 4.18 94.58

HSR 32hr 0.14 6 0.37 0.75 94.54 0.06 17 0.15 4.14 94.72

HSR 48hr 0.12 6 0.40 0.74 95.63 0.05 12 0.15 4.17 95.71

No HSR - 2hr 1.16 30 0.18 0.68 54.27 0.45 11 0.09 3.82 61.72

No HSR- 4hr 1.06 28 0.20 0.66 58.41 0.30 11 0.12 3.71 73.63

No HSR - 6hr 0.79 27 0.22 0.65 70.05 0.19 11 0.13 3.83 84.26

No HSR - 8hr 0.71 26 0.23 0.66 73.85 0.17 10 0.15 3.84 85.68

No HSR - 12hr 0.36 26 0.24 0.68 86.29 0.15 10 0.18 3.88 87.78

No HSR - 16hr 0.21 22 0.31 0.68 91.96 0.11 10 0.24 3.83 90.68

No HSR - 24hr 0.19 20 0.41 0.66 93.22 0.06 10 0.26 3.85 94.54

No HSR - 32hr 0.18 18 0.44 0.65 93.46 0.04 9 0.29 3.79 96.58

No HSR - 48hr 0.15 15 0.48 0.64 94.55 0.04 9 0.32 3.84 96.47

The results showed a reduction in the cyanide consumption when HSR was used. This is

shown in the table above.

The results indicate that using a HSR does improve the initial leach kinetics. Figures 13.2

and 13.3 show the leach curves for the bedrock and saprolite ore respectively with and

without use of a HSR. It should be noted that after 20hrs, leach results with and without HSR

were similar to conventional leaching.

Even though the HSR had the benefits as indicated above, it was not included in the design

because oxygenation test described in the next sub section indicated that sparging oxygen

gives comparable gold extraction. Using oxygen spargers requires less capital outlay in

comparison with HSR and would therefore be the preferred method for oxygen introduction.

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Figure 13.2: Bedrock Rate of Gold Dissolution With and Without High Shear Reactor

Figure 13.3: Saprolite Rate of Gold Dissolution With and Without High Shear Reactor

0

10

20

30

40

50

60

70

80

90

100

0 10 20 30 40 50 60

% A

u D

isso

luti

on

Leach Time (hrs)

Au dissolution without HSR Au dissolution with HSR

0

10

20

30

40

50

60

70

80

90

100

0 10 20 30 40 50 60

% A

u d

isso

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Leach time (hrs)

Au dissoltion without HSR Au dissolution with HSRFor

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13.7.4 Oxygenation Testwork Leach tests were performed to assess the use of oxygen or air for improved gold recovery

and included the following tests:

• Without air or oxygen sparging as a base case; • With air sparging; • With oxygen sparging.

The table below shows a summary of the effect of air and oxygen on gold extraction results.

Table 13.9: Effect of Sparging Air / Oxygen on Gold Extraction

Saprolite Bedrock

Gold

Extraction

DO

Level

Final Residue

Value

Gold

Extraction

DO

Level

Final Residue

Value

% mg/l g/t Au % mg/l g/t Au

No Air/Oxygen Sparging 93.08 10 0.08 89.06 7 0.29

With Air Sparging 95.03 10 0.06 90.79 8 0.25

With Oxygen Sparging 97.14 20 0.04 92.86 19 0.20

The results showed that oxygen sparging produced the best gold extraction and was

therefore used in the process design.

The tests indicated that sparging oxygen gives the better gold extraction. Tests to use a High

Shear Reactor (HSR) for oxygen injection showed an improvement in leach kinetics but no

improvement in the final gold extraction. Therefore a HSR was not selected as part of the

plant design. Added to this is the fact that the HSR is high in Capital and Operating cost.

13.7.5 Composite Leach Kinetic Results by Percentage Solids (w/w) Saprolite and bedrock leach kinetic tests were performed at time intervals between 2hrs and

48hrs at the following percentage solids (w/w) to assess the leach kinetics of saprolite and

bedrock ores types:

• 30% (saprolite only);

• 35% (saprolite only);

• 40%;

• 45%.

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Table 13.10: Saprolite Leach Kinetics at Varied Percent Solids.

Residence

Time

30 % solids 35 % solids 40 % solids 45 % solids

Reagent

consumption Au dissolution

Reagent


Reagent


Reagent


NaCN CaO Solid NaCN CaO Solid NaCN CaO Solid NaCN CaO Solid

kg/t kg/t Au % kg/t kg/t Au % kg/t kg/t Au % kg/t kg/t Au %

2hrs 0.13 4.06 65.12 0.12 4.02 68.25 0.09 3.82 61.72 0.06 3.73 56.56

4hrs 0.18 4.07 76.10 0.15 4.02 71.15 0.12 3.71 73.63 0.08 3.72 61.79

6hrs 0.20 4.06 80.06 0.19 3.96 82.15 0.13 3.83 84.26 0.10 3.72 79.70

8hrs 0.20 4.05 85.82 0.23 4.01 84.89 0.15 3.84 85.68 0.14 3.73 84.87

12hrs 0.29 4.04 88.49 0.25 4.04 85.61 0.18 3.88 87.78 0.17 3.73 90.19

16hrs 0.32 4.05 90.45 0.28 3.96 89.79 0.24 3.83 90.68 0.20 3.72 91.23

24hrs 0.41 4.03 92.87 0.30 3.95 90.91 0.26 3.85 94.54 0.22 3.74 93.10

32hrs 0.43 4.04 95.64 0.31 4.03 95.43 0.29 3.79 96.58 0.28 3.74 95.62

48hrs 0.45 4.03 96.51 0.31 3.96 96.46 0.32 3.84 96.47 0.32 3.74 95.50

Table 13.11: Bedrock Leach Kinetics at Varied Percent Solids

Residence

Time

40 % Solids 45 % Solids

Reagent


Reagent


NaCN CaO Solid NaCN CaO Solid

kg/t kg/t Au % kg/t kg/t Au %

2hrs 0.18 0.68 54.27 0.11 0.53 52.73

4hrs 0.20 0.66 58.41 0.20 0.55 65.31

6hrs 0.22 0.65 70.05 0.23 0.55 67.03

8hrs 0.23 0.66 73.85 0.26 0.54 75.13

12hrs 0.24 0.68 86.29 0.33 0.58 77.78

16hrs 0.31 0.68 91.96 0.37 0.57 83.31

24hrs 0.41 0.66 93.22 0.42 0.56 89.56

32hrs 0.44 0.65 93.46 0.44 0.58 91.26

48hrs 0.48 0.64 94.55 0.49 0.56 94.69

The testwork indicated the following optimum leach conditions for saprolite and bedrock:

• Saprolite: 24hrs leach time at 40% solids;

• Bedrock: 32hrs leach time at 42% solids. (42% was selected for design

purposes as it is easily achievable with densifying cyclones)

13.7.6 Preg- Robbing Testwork Preg-robbing tests were performed using leach rate tests with and without carbon addition

on the saprolite composite and bedrock composite samples. The preg-robbing results are

summarized in Table 13.12. The results indicate significant preg robbing tendencies for both

the saprolite and bedrock ore. The effect of pre-robbing was performed using rate tests with

and without carbon addition at the following leach times:

• 2hrs; • 4hrs; • 6hrs; • 8hrs; • 24hrs.

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The results indicated that the preg-robbers were most active during the initial stages of leach

with 28.12% and 15.92% gold lost to preg-robbers in the first two hours for the saprolite and

bedrock ores respectively. Due to the high preg-robbing tendency of the ore, a CIL circuit

was selected in the process design and cyanide detox was included to minimize cyanide

contact with the gold prior to activated carbon being present.

Table 13.12: Effect of Preg-Robbing

Time

Saprolite Bedrock

Au Dissolution (%) Au Dissolution (%)

With

Carbon Without Carbon

Preg-

Robbing

With

Carbon Without Carbon

Preg-

Robbing

2hrs 55.16 27.04 28.12 58.70 42.78 15.92

4hrs 76.34 65.00 11.34 63.77 51.86 11.91

6hrs 81.41 69.62 11.79 69.70 62.61 7.09

8hrs 84.75 72.84 11.91 76.59 70.99 5.60

24hrs 92.47 75.32 17.15 91.36 78.82 12.55

Average 16.06 10.61

13.7.7 Variability Preg-Robbing Testwork Since preg-robbing test on the composite samples indicated that the bedrock and saprolite

ores contain high pre-robbers, variability preg-robbing tests were conducted on the variability

samples to identify if the preg-robing tendency is across the entire ore body. The variability

preg-robbing tests were performed at a residence time of 2hrs since the initial rate tests

indicated that the preg-robbers are most active in the first 2hrs.

The table below includes the variability preg-robbing results on the saprolite and bedrock ore

and the results indicate the following:

• Preg-robbing on the variability saprolite samples ranged from 8.45% to 26.72% with an average of 17.28%;

• Preg-robbing on the variability bedrock samples ranged from 17.25% to 34.55% with an average of 25.02%.

The results confirmed the presence of preg-robbers and showed that the preg-robbing tendency of the ore is variable in nature.

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Table 13.13: Variability Preg-Robbing Results Summary

Au Dissolution (%) Preg-robbing

(%) With carbon Without carbon

Bedrock

Var 1 North 62.83 28.28 34.55

Var 2 North 44.30 24.87 19.43

Var 3 North 47.61 18.74 28.87

Var 4 North 37.94 20.69 17.25

Min 17.25

Max 34.55

Average 25.02

Saprolite

Var 1 South 70.82 48.00 22.82

Var 2 South 58.02 39.41 18.62

Var 3 South 55.59 41.52 14.07

Var 4 South 65.25 56.79 8.45

Var 5 North 58.42 44.86 13.56

Var 6 North 81.90 61.20 20.70

Var 7 North 53.88 27.17 26.72

Var 8 North 73.14 59.81 13.33

Min 8.45

Max 26.72

Average 17.28

13.7.8 Thickening and Rheology Testwork Thickening testwork conducted on the saprolite and bedrock ore indicated that both post

leach slurries can be effectively thickened. However for the bedrock slurry, when raw water

was used to dilute the slurry, the ionic strength changed causing poor settling properties and

higher flocculant consumptions. When the ionic strength was altered back to the original

values as received from leach, the settling properties and flocculant consumptions improved.

This indicates that the bedrock ore is sensitive to changes in ionic strength i.e. pH. A

summary of the thickening and rheology results are shown in the table below.

Table 13.14: Thickening and Rheology Results

Units Saprolite Bedrock

Solid SG

2.78 2.90

Flocculation Type

Magnafloc 6260 Magnafloc 6261

Dosing Concentration % 0.025 0.025

Flocculation Dosage Rate g/t 40 40*

Feed slurry Solid Concentration % m/m 3 7.5

Critical Solids Flux Rate t/(m2.h) 0.2 0.3

Estimated U/F Solids Concentration % m/m 40# 60

*This is provided the slurry is in a coagulated state prior to the flocculation stage. #

The saprolite compacted to a 52% (m/m) solid concentration after a period of 24hrs picket racking.

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13.7.8.1 Rheology and Viscosity Modifier Test

During the initial feasibility study, it was observed on Saprolite ore that viscosity issues may

be encountered. In order to possibly increase cyclone overflow densities to higher values,

viscosity modifier testwork was conducted during this phase.

Viscosity tests were conducted on the saprolite ore at natural and at an elevated pH of 10.5

at various percentage solids. Refer to figure below for the viscosity versus percent solids

relationship.

It was found from the graph that the yield stress at natural and at pH 10.5 was similar.

Viscosity modifiers Antiprex D and DP 725 were investigated to assess if they could reduce

the viscosity issues. The results showed that viscosity modifiers did reduce the viscosity of

the saprolite slurry and at this stage was not included as part of the design as insufficient

tests were performed. In addition to this, the tests were performed on the milled product only

and not on the thickener underflow. Additional viscosity modifier tests have been included in

the recommendations.

Tests were conducted at varied concentrations of the viscosity modifier. From the tests, the

following optimum viscosity additions were selected:

• Antiprex D: 0.3ml of Antiprex D per litre of slurry which translates to 354ml per tonne of dry ore.

• DP 725: 0.9ml of DP 725 per litre of slurry which translates to 1,063ml per tonne of dry ore.

Figure 13.4: Slurry Yield Stress vs. Percent Solids Curves

Reference – Figure 11: Comparison of Yield Stress versus Mass Solids Concentration. (Patterson & Cooke Consulting Engineers, Pty. Ltd., 2012 “Endeavour Corporation’s Agbaou Project: Slurry Test Report” Report number: SEN-AGB 8284 RO1 Rev 1.)

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13.7.8.2 Variability Leach Results on Middlings and Tails

Variability leach tests were performed at optimum conditions to determine the dissolutions

and reagent consumptions. This was then compared to the optimum results to determine the

ores variability and it was observed that the ore was not highly variable.

Table 13.15: Variability Leach Results on Middles and Tails Variability Leach on Midds & Tails

Saprolite Bedrock

Final Residue Au Dissolution Reagent Consumptions Final Residue Au Dissolution Reagent Consumptions

Solid (g/t Au) % Cyanide (kg/t) Lime (kg/t) Solid (g/t Au) % Cyanide (kg/t) Lime (kg/t)

Max 0.12 95.75 0.40 4.86 0.18 93.64 0.21 1.68

Min 0.05 89.06 0.16 2.11 0.09 89.70 0.14 1.03

Ave 0.08 92.42 0.26 2.81 0.13 91.94 0.16 1.31

P85 0.10 93.53 0.33 3.23 0.16 93.53 0.18 1.50

13.7.8.3 Optimum Leach Parameters

Selected leach parameters were varied to obtain optimum conditions for the design of the

leach circuit.

The effect of cyanide addition test was performed by conducting leach tests at the following

cyanide additions:

• 0.3kg/t NaCN;

• 0.5kg/t NaCN;

• 0.7kg/t NaCN;

• 1.0kg/t NaCN.

The effect of sparging oxygen test was performed at the following conditions:

• Without air or oxygen sparging as a base case;

• With air sparging;

• With oxygen sparging;

• With oxygen injection via a High Shear Reactor.

The effect of percent solids test was performed as rate test at the following percent solids:

Saprolite

• 30%;

• 35%;

• 40%;

• 45%;

Bedrock

• 40%;

• 45%.

The table below is a summary of the optimum leach parameters selected.

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Table 13.16: Optimum Leach Parameters Optimum Leach Conditions

Saprolite Bedrock

Cyanide Addition 0.7kg/t 0.7kg/t

Residence Time 24hrs with oxygen sparging 24hrs with oxygen sparging

pH 10.50 10.50

Solids (% m/m) 40 42

13.7.9 Recommendations for Design Values Design values were selected from the testwork results for the treatment of the Agbaou ore

and are summarized below. All the testwork results from the initial feasibility testwork and

the EOS testwork were populated and analyzed to determine the following values:

• Average;

• Maximum;

• Minimum;

• 85th Percentile.

The 85th percentile values for reagent consumptions and dissolution and SG from Paterson

& Cooke results were used as the design values (SRK Consulting, SENET, Knight Piésold

Consulting, 2012 Appendix 13-J).

The table below shows the selected design values used for the process flowsheet

development.

Table 13.17: Selected Design Values Units Saprolite Bedrock

Grind

80%-75µm 80%-75µm

Selected Milling Circuit

SABC SABC

Solids SG

2.79 2.82

Final Residue (Gravity-CIL) g/t Au 0.10 0.16

Cyanide Consumption kg/t 0.33 0.18

Lime Consumption kg/t 3.23 1.50

Leach Time Hrs. 24 32

Oxygen Sparge During Leach

yes yes

CIL Percent Solids % m/m 40 42

Thickening Flux Rate t/(m2.h) 0.2 0.3

13.7.10 Recommendations for Further Testwork (Post EOS Testwork) The following additional testwork is recommended to confirm the assumptions made in the

design:

• Comminution tests on the saprolite ore - the samples received during the EOS were too fine for any comminution testwork. Assumptions in the design were made for the saprolite comminution parameters and therefore need verification;

• EGRG testwork - the samples received during the EOS were too fine for EGRG testwork. It is therefore recommended to perform EGRG and modelling to determine the actual plant gravity recoveries;

• Viscosity modifier testwork on thickener underflow - the viscosity modifier testwork

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performed during the EOS was limited and was only done on the milled product but not on the thickener underflow product;

• Variability viscosity testwork - variability viscosity testwork is recommended to assess the extent of the viscosity concerns across the ore body.

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SECTION 14. MINERAL RESOURCE ESTIMATES

14.1 INTRODUCTION The Mineral Resource Statement presented herein represents an update of the Mineral

Resources prepared for the Agbaou Project by Coffey Mining on behalf of Etruscan

Resources Inc. in March 2008 in accordance with the Canadian Securities Administrators’

National Instrument 43-101.

The mineral resource model prepared by SRK considers 698 RC and core boreholes drilled

by ENDEAVOUR and other owners of the property during the period of 2003 to 2011. The

resource estimation work was completed by Mark Wanless, Pr.Sci.Nat (SACNASP -

400178/05) an appropriate “independent qualified person” as this term is defined in National

Instrument 43-101. The effective date of the resource statement is 25th May 2012.

This section describes the resource estimation methodology and summarizes the key

assumptions considered by SRK. In the opinion of SRK, the resource evaluation reported

herein is a reasonable representation of the recoverable gold mineral resources found in the

Agbaou Project at the current level of sampling. The mineral resources have been estimated

in conformity with generally accepted CIM “Estimation of Mineral Resource and Mineral

Reserves Best Practices” guidelines and are reported in accordance with CIM “Definition

Standards For Mineral Resources and Mineral Reserves” as referenced by the Canadian

Securities Administrators’ National Instrument 43-101. Mineral resources are not mineral

reserves and do not have demonstrated economic viability. There is no certainty that all or

any part of the mineral resource will be converted into mineral reserve.

The database used to estimate the Agbaou Project mineral resources was audited by SRK.

SRK is of the opinion that the current drilling information is sufficiently reliable to interpret

with confidence the boundaries for vein hosted gold mineralization and that the assay data

are sufficiently reliable to support mineral resource estimation.

Datamine Studio Version 3.20 was used to construct the geological solids, prepare assay

data for geostatistical analysis, and construct the block model. The Isatis (version 11.03)

software was used for geostatistical analysis, variography, estimate metal grades and

estimate the recoverable resources using a global change of support and Uniform

Conditioning (“UC”) and Multiple Indicator Kriging (“MIK”). Block models were constructed in

Datamine for Mineral Resource reporting, and replicated in Surpac software, for use in

mineral reserve estimation.

14.2 RESOURCE ESTIMATION PROCEDURES The resource evaluation methodology involved the following procedures:

• Database compilation and verification; • Construction of wireframe models for the boundaries of the gold mineralization; • Definition of resource domains;

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• Data conditioning (compositing and capping) for geostatistical analysis and variography;

• Block modeling and grade interpolation; • Recoverable resource estimation; • Resource classification and validation; • Assessment of “reasonable prospects for economic extraction” and selection of

appropriate cut-off grades; • Preparation of the Mineral Resource Statement.

14.3 RESOURCE DATABASE The exploration data supplied to SRK by ENDEAVOUR was in the form of a Microsoft

Access Database with the following tables included:

• Collar – containing the drill hole collar ID and surveyed co-ordinates, and planned drill hole orientations;

• Survey – containing the down hole orientation survey results; • Geology – containing the geological logs of the drill holes, including fields describing:

o Rock type; o Silicification; o Quartz Veining; o Pyrite; o Oxidation; o Weathering.

• Assay – containing sample ID and gold grade in g/t; and • Geotech – containing geotechnical measurements, and core recovery percentage.

In addition to the drilling database, SRK were supplied with surveyed contours of the surface

topography for the majority of the area surrounding the resources.

The database used in the estimation contained both diamond drill and reverse circulation

drilling results. The data was validated during the import into Datamine to ensure that there

were no unexplained gaps in the data, no overlapping intervals, unusual collar elevations,

survey data was present at the start of each hole and there was consistency of hole names

between input tables. For samples where the analytical laboratory returned a result of less

than detection limit, the value was replaced with half of the detection limit. Datamine

automatically assigns a ‘missing value’ where no assay results are present.

14.4 SOLID BODY MODELLING

14.4.1 Primary Mineralisation The modeling of the mineralized body solids was undertaken in Datamine, using

conventional sectional interpretations. String outlines of the mineralized zones were digitised

on screen along drill lines, and ‘snapped’ onto drill holes to ensure that the mineralized

boundaries were accurately honored.

SRK initially used the wireframes previously generated by Coffey Mining as a reference as

this provided a useful starting point to the interpretation. The approach to the modeling of the

zones was to outline continuously developed zones with an average grade of greater than

0.3 g/t.

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Based on the preference of the ENDEAVOUR geologists and the outcomes of an

independent review of the Coffey Mining estimate by R. Marinho of AMEC Americas Limited

(“AMEC”), the SRK approach was less tolerant of internal dilution at the expense of a

reduction in tonnes. In other words, SRK selected shorter mineralized intervals as the

mineralized zone, and excluded some relatively isolated high grade intersections in order to

reduce dilution and therefore increase the grade of the modeled zones. In many areas the

SRK and Coffey Mining interpretation of the ore zones are near identical, however in other

areas SRK either selected a shorter interval, or revised the interpretation of the mineralized

zones completely. SRK have interpreted some zones differently to Coffey Mining, based on

the mineralization character of the zones and additional information available from the infill

drilling program.

The sectional interpretations along all drill lines linked to form a solid wireframe envelope.

Wireframes were typically extended half the distance between drill hole lines where the

drilling terminated along strike, or where the mineralization was interpreted to terminate.

Plan views at stepped elevations were used as a verification of the interpretation of the

mineralized zones.

A total of ten zones were interpreted in the Main and Southern areas of the project, and an

additional three zones were interpreted in satellite mineralization in the west of the project.

An additional zone of mineralization constrained to the laterite developed at or near surface

was also modeled and is described in section 14.4.2. In the Main and Southern zones, only

two zones are present over the total strike of the interpreted ore body.

The mineralization is displaced by a major fault at the approximate center of the exploration

area, striking close to north-east to south-west, and with a sinistral lateral displacement

which separates the Main area from the Southern area. It is unclear whether or not there is

any component of vertical displacement. The main mineralized zones are interpreted to be

developed on both sides of the fault, and the mineralized zones have been correlated across

the fault. The primary mineralization wireframe models are presented in

Figure 14.1.

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Figure 14.1: Primary Mineralisation Wireframe Models

Oblique view looking approximately North

Oblique v

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Primary Mineralisation Wireframe Models

Oblique view looking approximately North-West

Oblique view looking approximately North-East

Agbaou Gold Project Primary Mineralisation Models

1

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9

5

24 7

8

10

Page 14-4

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14.4.2 Surface Mineralisation Mineralization has also been modeled in the laterite that overlies most of the project. The

mineralization is not limited to the sub-crop position of the primary mineralized zones;

however the tenor of the laterite mineralization above the primary mineralization is generally

enhanced. The laterite unit was modeled as a whole, without grade selectivity. The top and

bottom surface of the laterite intersections were modeled using Leapfrog software, which

enables the derivation of a smoother surface, which, while honoring the location of the

intersections, is extrapolated between data locations using a trend function. The top and

bottom surfaces intersect each other, and this is interpreted as the limit of the laterite

development.

14.4.3 Weathering Zones Composites were also coded by the weathering profile. The profile has been modeled from

drill data and comprises of laterite, saprolite, transition material and fresh units. Weathering

surfaces generally parallel the topographical profile, although weathering can be deeper

within zones of mineralization. The interfaces between weathering zones were modeled in a

similar fashion to the laterite unit, using Leapfrog software. In some drill holes the weathering

layers are more complicated than the simple sequence of saprolite, transitional and fresh, as

there may be intersections of fresh surrounded by transitional material. In these instances,

SRK selected the position which best matched the surrounding intersections. The surface

was also forced to honor intersection points.

14.5 COMPOSITING SRK analyzed the sample lengths before starting a composite optimization analysis for each

zone. Although the sample lengths are variable within the ore zones, the majority of the

samples are 1m lengths, with a second peak of samples at 2m lengths. Approximately 51%

of the samples within the mineralized zones are at 1m lengths, while about 17% are at 2m

lengths. The majority of the irregular length samples are less than 1m, while a small number

are greater than 2m.

The statistics of grade and sample length of samples within each mineralized zone are

reported in Table 14.1. The gold grade values are length weighted.

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Table 14.1: Gold Grade and Sample Length Statistics for Samples Within Mineralised Zones

Variable Zone Count Min Max Mean Std.dev Variance Cov

Au (g/t) 1 2,290 0.0005 74.97 3.01 5.77 33.2 1.92

2 2,060 0.005 74.21 2.09 4.90 24.0 2.34

3 307 0.005 24.90 1.41 2.91 8.50 2.06

4 638 0.005 32.09 1.05 3.01 9.00 2.85

5 282 0.005 94.90 6.75 12.51 156.5 1.85

6 257 0.005 54.27 4.40 7.68 59.0 1.75

7 693 0.005 25.40 1.55 3.07 9.40 1.99

8 76 0.005 4.360 0.64 1.03 1.10 1.60

9 615 0.002 27.97 0.99 2.33 5.50 2.36

10 179 0.005 29.18 2.14 4.94 24.4 2.31

11 1,665 0.005 109.8 0.77 4.15 17.2 5.36

51 353 0.005 25.54 1.64 3.32 11.0 1.91

52 58 0.01 13.63 0.82 1.83 3.40 1.98

53 42 0.03 6.810 0.80 1.18 1.40 1.61

Variable Zone Count Min Max Mean Std.dev Variance

Total

Length

Length

(m)

1 2,290 0.01 3.1 0.52 0.27 2410.2 2,410

2 2,060 0.01 9.0 0.57 0.33 2039.5 2,039

3 307 0.01 2.0 0.60 0.36 310.8 311

4 638 0.01 2.0 0.61 0.37 664.3 664

5 282 0.01 2.0 0.47 0.22 324.7 325

6 257 0.05 2.0 0.36 0.13 264.3 264

7 693 0.01 2.0 0.55 0.30 641.4 641

8 76 0.05 2.0 0.37 0.14 54.7 55.0

9 615 0.01 2.0 0.54 0.29 562.8 563

10 179 0.10 2.0 0.55 0.31 220.3 220

11 1,665 0.10 5.0 1.26 0.54 0.3 2,100

51 353 0.10 2.0 1.17 0.64 0.4 415

52 58 0.15 2.0 1.40 0.63 0.4 81.0

53 42 0.10 2.0 1.49 0.60 0.4 62.0

SRK created composites at a range of lengths from 1m to 5m, and compared the impact on

the grade and variance for each zone. The composites were created at a specific length, but

allowed to vary from the specified length to ensure that all samples were included in the

composites. The adjustment is typically at the base of a mineralized zone. The composites

were allowed to be from half the ideal composite length to 1.5 times the ideal composite

length. For example, for the 2m composite length, the minimum composite length would be

1m and the maximum would be 3m.

The composite optimization was undertaken on data from the Main and Southern areas, and

the final composite length selected from the optimization was applied to the West area

resource estimation. The analysis of the composites showed for most zones, that the grade

of the composites was relatively insensitive to composite length, but for longer composites,

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the grade became more variable. In addition, for some of the narrower zones the average

composite length started to deviate from the standard composite length, indicating that the

zone thickness is close to the composite length. The Coefficient of Variation (“CoV”) of most

of the zones is typically between 2 and 3 on the original sample lengths. Compositing to

longer lengths reduced this value in all zones, however even for relatively long composite

lengths (up to 5m) the CoV remained well above 1. For estimation using Kriging a CoV of

less than 1 is optimal, however not even very long composite lengths achieved this.

SRK elected to retain the 2m composite length used in the previous estimate by Coffey

Mining, as this provided the best balance between regular lengths (i.e. equal support),

retaining the detailed information within the thinner zones, and reducing the variance to

improve the estimation. Considering the eventual estimation block size of 10m by 10m

parallel to the drilling direction, a 1m composite was considered to be too small.

An analysis of the non-standard composite lengths shows that the exclusion of the shorter

lengths was likely to increase the grade of the samples. The statistics of gold grade and

length, of subsets of the composites within the Main and South mineralized zones are

presented in Table 14.2. In Table 14.2 the category “Standard” refers to 2m composites,

“Non-Standard” referred to all composites that are longer or shorter than 2m. “Shorter” and

“Longer” refer to composites shorter and longer than 2m, while “All” includes all composites

regardless of length.

The mean grade of the 2m composites is higher than the mean of all the composites, and

also higher than the mean grades of all the composites that are longer or shorter than 2m.

Excluding the non-standard length composites would therefore most likely have created a

slight high bias in the datasets.

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Table 14.2: Statistics of Standard and Non-Standard Length Composites Domain Variable Count Minimum Maximum Mean Std. Dev. Variance CoV

Standard AU 2833 0.005 109.83 2.31 5.64 31.79 2.44

Non-Standard AU 1983 0.005 59.55 1.77 3.76 14.16 2.12

Longer AU 496 0.005 15.96 1.26 2.06 4.260 1.64

Shorter AU 1487 0.005 59.55 1.94 4.17 17.35 2.14

All AU 4816 0.005 109.83 2.09 4.96 24.60 2.38

Standard LENGTH 2833 2.00 2.000 2.00 0.00 0.000 0.00

Non-Standard LENGTH 1983 1.00 2.950 1.86 0.25 0.060 0.14

Longer LENGTH 496 2.00 2.950 2.13 0.14 0.020 0.06

Shorter LENGTH 1487 1.00 1.990 1.77 0.22 0.050 0.12

All LENGTH 4816 1.00 2.950 1.94 0.18 0.030 0.09

14.6 EVALUATION OF OUTLIERS SRK conducted an analysis of the data for each zone, and assessed the need for capping of

high grade outlier values. The statistics of the raw gold grades (weighted by length) are

presented in Table 14.3 for each zone. The mineralized zones in the Main and Southern

areas are interpreted to be continuous (aside from the fault displacement discussed in

section 14.4) and therefore are grouped into zones named 1 to 10. The laterite is called zone

11, and the West area zones are called zones 51 to 53.

Table 14.3: Statistics of Composites for Gold by Mineralised Zone Zone Count Minimum Maximum Mean Std. Dev. Variance CoV

1 1228 0.005 66.38 3.00 5.11 26.06 1.70

2 1041 0.005 62.83 2.11 4.28 18.36 2.05

3 163 0.005 24.90 1.41 2.66 7.10 1.88

4 338 0.005 17.10 1.03 1.89 3.56 1.83

5 165 0.005 94.90 6.75 13.25 175.67 1.96

6 136 0.005 49.70 4.42 7.66 58.66 1.74

7 325 0.005 25.40 1.63 3.06 9.36 1.98

8 28 0.005 3.85 0.78 1.13 1.27 1.73

9 286 0.005 17.88 0.99 1.89 3.57 1.92

10 115 0.005 29.18 2.14 4.48 20.11 2.10

11 991 0.005 109.83 0.77 3.78 14.27 4.44

51 236 0.005 25.54 1.69 3.17 10.04 1.88

52 48 0.010 13.63 0.83 2.00 4.01 2.42

53 30 0.030 6.81 0.90 1.38 1.91 1.53

The SRK outlier analysis included a visual assessment of the histograms of each zone,

along with a set of charts which assist in determining an appropriate capping level.

Examples of these plots are presented in Figure 14.2. For the first three plots, the samples

are sorted on their gold grade, and then statistics calculated on samples, starting with the

lowest grade two samples, and then repetitively adding samples and recalculating the

statistics. The aim of these charts is to assess the impact on the dataset of individual

samples on the data mean, standard deviation and the CoV. The last two plots are

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histograms of the contribution to the metal content of samples (in 10% bins for the fourth

plot, and in 1% bins from 90% upwards in the last plot). The charts are used to assess the

impact of high grade samples to the statistics of each dataset.

Figure 14.2: Zone 1 Capping Analysis Charts

Agbaou Gold Project Zone 1 Capping Analysis Charts

Project No.

439430

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On the basis of the analysis for each zone, the capping presented in Table 14.4 was applied

to the data for estimation. The capping tried to limit loss of metal to less than 5%.

Table 14.4: Summary of Capping Applied in Ordinary Kriged Estimate Per Zone Uncapped Capped

Zone No.

Comps Mean

Std.

Dev CoV

Cap

Value

No.

Capped Mean

Std.

Dev CoV

Metal

Loss

1 1228 3.00 5.11 1.70 30 6 2.90 4.32 1.48 3.20%

2 1041 2.11 4.28 2.05 28 5 2.04 3.52 1.74 3.26%

3 163 1.41 2.66 1.88 15 1 1.35 2.20 1.63 4.52%

4 338 1.03 1.89 1.83 - 0 1.03 1.89 1.83 0.00%

5 165 6.75 13.25 1.96 40 5 6.12 10.26 1.68 9.44%

6 136 4.42 7.66 1.74 30 2 4.27 6.95 1.63 3.40%

7 325 1.63 3.06 1.98 20 3 1.60 2.87 1.89 1.82%

8 28 0.78 1.13 1.73 - 0 0.78 1.13 1.73 0.00%

9 286 0.99 1.89 1.92 10 2 0.94 1.49 1.58 4.49%

10 115 2.14 4.48 2.10 20 3 1.96 3.55 1.81 8.33%

11 1159 0.77 3.48 4.58 20 2 0.72 1.43 2.10 11.3%

51 236 1.69 3.17 1.88 - 0 1.69 3.17 1.88 0.00%

52 48 0.83 2.00 2.42 - 0 0.83 2.00 2.42 0.00%

53 30 0.90 1.38 1.53 - 0 0.90 1.38 1.53 0.00%

For most zones, the metal loss from capping is relatively small, with less than 5% considered

an acceptable value. For zones 5, 10 and 11 there is a relatively high metal loss. For Zone

5, this is a result of the capping of some very high grade samples (some over 90 g/t) and the

generally high grade of the zone, and the capping of five high grade outliers out of a

relatively small population of samples. Similarly for zone 11 the highest grade samples

capped were over 100 g/t, which has a significant impact on the metal content (and hence

the need for capping). For zone 10, although the capping was relatively conservative, the

small population of composites (115) means that the impact in terms of metal loss is

relatively high.

14.7 STATISTICAL ANALYSIS AND VARIOGRAPHY The summary statistics of the grade of the 2m composites are presented in Table 14.3 for

each mineralized zone. Zones 1 and 2 have a large number of composites, and along with

the laterite zone (Zone 11) are the best informed zones. Zones such as 4, 7 and 9 also have

a significant number of composites, while zones such as zone 8 have too few samples to be

able to characterize the grade population adequately.

The initial testing done by SRK indicated that for the zones with few samples, it was not

possible to generate a semi-variogram with an interpretable structure. Based on an

assessment of the grade characteristics and the geometry and character of the mineralized

zones, and any indication of continuity within the experimental semi-variograms, SRK

grouped some of the sparsely sampled zones together. The estimation has then used a

common semi-variogram for the group, which is calculated on the best sampled zone, and

scaled to the variance of the data within each zone.

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Zone 1 and 2 were used as the primary zones for the grouping in the Main and South areas,

and zone 51 in the West area. The laterite (zone 11), zone 3 and zone 9 has sufficient data

to generate their own semi-variograms. The zones grouped with zone 1 (and sharing the

zone 1 semi-variogram) are zones 4, 5, and 6. The zones grouped with zone 2 (sharing the

zone 2 semi-variogram) are zone 7, 8, and 10. In the West area, only zone 51 had sufficient

data to generate a semi-variogram, and zone 52 and 53 are grouped along with zone 51.

The histograms of the composites for all zones are displayed, (grouped in the same manner

as the semi-variograms) in Figure 14.3 to Figure 14.6 for the Zone 1 group, Zone 2 group,

individual zones, and the western zones respectively. For the laterite (zone 11) two

histograms are presented, one displaying all data, and one restricted to samples below 20

g/t as the high grade outliers in this zone make the histogram display with all data relatively

uninformative.

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Figure 14.3: Histograms of Gold Grade for the Zone 1 Group of Zones

Zone 1 Zone 4

Zone 5 Zone 6

Agbaou Feasibility Study Histogram of Gold Grades for Zone 1 Group of

Zones

Project No.

439430

0

0

10

10

20

20

30

30

40

40

50

50

60

60

70

70

AU

AU

0.0 0.0

0.1 0.1

0.2 0.2

0.3 0.3

0.4 0.4

0.5 0.5

Frequencies

Frequencies

Nb Samples: 1228

Minimum: 0.0040

Maximum: 66.3800

Mean: 3.0024

Std. Dev.: 5.1032

0

0

5

5

10

10

15

15

AU

AU

0.0 0.0

0.1 0.1

0.2 0.2

0.3 0.3

0.4 0.4

Frequencies

Frequencies

Nb Samples: 338

Minimum: 0.0050

Maximum: 17.1000

Mean: 1.0387

Std. Dev.: 1.8837

0

0

50

50

100

100

AU

AU

0.0 0.0

0.1 0.1

0.2 0.2

0.3 0.3

0.4 0.4

0.5 0.5

0.6 0.6

Frequencies

Frequencies

Nb Samples: 165

Minimum: 0.0050

Maximum: 94.9000

Mean: 6.7573

Std. Dev.: 13.2139

0

0

10

10

20

20

30

30

40

40

50

50

AU

AU

0.0 0.0

0.1 0.1

0.2 0.2

0.3 0.3

0.4 0.4

Frequencies

Frequencies

Nb Samples: 136

Minimum: 0.0050

Maximum: 49.7000

Mean: 4.3796

Std. Dev.: 7.6309

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Figure 14.4: Histograms of Gold Grade for the Zone 2 Group of Zones

Zone 2 Zone 7

Zone 8 Zone 10

Agbaou Gold Project Histogram of Gold Grades for Zone 2 Group of

Zones

Project No.

439430

0

0

10

10

20

20

30

30

40

40

50

50

60

60

AU

AU

0.0 0.0

0.1 0.1

0.2 0.2

0.3 0.3

0.4 0.4

0.5 0.5

0.6 0.6

Frequencies

Frequencies

Nb Samples: 1041

Minimum: 0.0050

Maximum: 62.8300

Mean: 2.0988

Std. Dev.: 4.2826

0

0

10

10

20

20

AU

AU

0.0 0.0

0.1 0.1

0.2 0.2

0.3 0.3

0.4 0.4

Frequencies

Frequencies

Nb Samples: 325

Minimum: 0.0050

Maximum: 25.4000

Mean: 1.5424

Std. Dev.: 3.0554

0

0

1

1

2

2

3

3

4

4

AU

AU

0.0 0.0

0.1 0.1

0.2 0.2

0.3 0.3

0.4 0.4

Frequencies

Frequencies

Nb Samples: 28

Minimum: 0.0050

Maximum: 3.8457

Mean: 0.6793

Std. Dev.: 1.1060

0

0

10

10

20

20

30

30

AU

AU

0.0 0.0

0.1 0.1

0.2 0.2

0.3 0.3

0.4 0.4

Frequencies

Frequencies

Nb Samples: 115

Minimum: 0.0050

Maximum: 29.1800

Mean: 2.0822

Std. Dev.: 4.4653

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Figure 14.5: Histograms of Gold Grade for the Individual Zones

Zone 3 Zone 9

Zone 11 All Zone 11 Below 20

Agbaou Feasibility Study Histogram of Gold Grades for Individual Zones

Project No.

439430

0

0

5

5

10

10

15

15

AU

AU

0.00 0.00

0.05 0.05

0.10 0.10

0.15 0.15

0.20 0.20

0.25 0.25

Frequencies

Frequencies

Nb Samples: 163

Minimum: 0.0050

Maximum: 15.0000

Mean: 1.3611

Std. Dev.: 2.1914

0

0

5

5

10

10

15

15

AU

AU

0.0 0.0

0.1 0.1

0.2 0.2

0.3 0.3

0.4 0.4

Frequencies

Frequencies

Nb Samples: 286

Minimum: 0.0025

Maximum: 17.8777

Mean: 0.9904

Std. Dev.: 1.8873

0

0

50

50

100

100

AU

AU

0.00 0.00

0.25 0.25

0.50 0.50

0.75 0.75

1.00 1.00

Frequencies

Frequencies

Nb Samples: 1087

Minimum: 0.00

Maximum: 109.83

Mean: 0.83

Std. Dev.: 3.75

0

0

10

10

20

20

AU

AU

0.0 0.0

0.1 0.1

0.2 0.2

0.3 0.3

0.4 0.4

Frequencies

Frequencies

Nb Samples: 1087

Minimum: 0.00

Maximum: 109.83

Mean: 0.83

Std. Dev.: 3.75

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Figure 14.6: Histograms of Gold Grade for the Western Area Zones

Zone 51 Zone 52

Zone 53

Agbaou Gold Project Histogram of Gold Grades for the Western Area

Zones

Project No.

439430

The semi-variograms on the raw data showed poor or no structure in all cases, and SRK

elected to transform the raw gold grade variables to a Gaussian distribution, with a normal

score transformation, using hermite polynomials. The data were declustered using a rotated

block configuration before transformation. A range of block sizes were tested for each zone,

between 2.5m and 20m in the plane of the mineralization, and between 0.5m and 5m across

the ore bodies. The resulting Gaussian variable statistics have a normal or Gaussian

distribution whereby the minimum and maximum should be very similar in the positive and

negative; the mean should be close to zero; the variance and standard deviation close to

one.

The declustered block size was selected to achieve the closest to the ideal distribution for

the Gaussian variable. The statistics of the Gaussian variables used to generate the

0

0

10

10

20

20

AU

AU

0.0 0.0

0.1 0.1

0.2 0.2

0.3 0.3

0.4 0.4

0.5 0.5

Frequencies

Frequencies

Nb Samples: 236

Minimum: 0.0050

Maximum: 25.5400

Mean: 1.6899

Std. Dev.: 3.0646

0

0

5

5

10

10

15

15

AU

AU

0.0 0.0

0.1 0.1

0.2 0.2

0.3 0.3

Frequencies

Frequencies

Nb Samples: 48

Minimum: 0.0100

Maximum: 13.6300

Mean: 0.8279

Std. Dev.: 1.5945

0

0

1

1

2

2

3

3

4

4

5

5

6

6

7

7

AU

AU

0.0 0.0

0.1 0.1

0.2 0.2

0.3 0.3

Frequencies

Frequencies

Nb Samples: 30

Minimum: 0.0300

Maximum: 6.8100

Mean: 0.9030

Std. Dev.: 1.2851

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experimental semi-variograms are presented in Table 14.5.

Table 14.5: Statistics of Gaussian Gold Variables per Zone Zone Count Minimum Maximum Mean Std. Dev. Variance

1 1,228 -3.25 3.28 0.0000 0.999 0.998

2 1,041 -3.23 3.12 -0.0001 0.999 0.998

3 163 -2.66 2.33 -0.0025 0.989 0.978

4 338 -2.79 2.76 -0.0001 0.996 0.991

5 165 -2.68 2.72 0.0002 0.995 0.989

6 136 -2.17 2.47 0.0029 0.981 0.963

7 325 -2.84 2.84 0.0000 0.996 0.993

8 28 -1.53 1.87 0.0161 0.920 0.846

9 286 -2.90 2.88 0.0002 0.998 0.995

10 115 -2.46 2.60 0.0008 0.991 0.981

11 1,087 -3.18 3.18 0.0000 0.999 0.997

51 236 -2.51 5.57 0.0409 1.246 1.553

The semi-variograms for zones 1, 2, 3, 9, 11 and 51 were modeled in Gaussian space. The

experimental semi-variograms were oriented such that the X and Y axes approximated the

dip and strike directions of the ore bodies. The Z direction is close to the drill hole direction,

and the experimental semi-variogram search parameters were selected such that the Z

direction approximates a true down hole semi-variogram.

The modeled semi-variograms in Gaussian space were scaled to the population variance of

each zone, according to the groupings, retaining the directional ranges and the ratios of

each of the structures. The modeled semi-variograms in Gaussian space are presented in

Figure 14.7, and the semi-variogram parameters used in the estimates (scaled to normal

space) are presented in Table 14.6.

All semi-variograms were modeled with two spherical structures, except the Western area

which was modeled with only one.

As the Zone 3 and 9 experimental semi-variograms have reached a (relatively) stable sill

which is lower than the population variance. SRK have therefore added a very long range

structure to ensure that the total sill of the semi-variogram model matches the population

variance, while still matching the inherent structure in the experimental data. The longer

range portion of the semi-variogram is not accessed for the shorter range estimates. The

very long range ensures that the Kriging weights of any estimates that do access portions of

the semi-variogram beyond the first modeled ranges will not be meaningfully impacted by

the second range.

Zone 1 (and the zones grouped with zone 1) shows a low nugget (modeled on the down hole

semi-variogram) and a short range structure that account for approximately 90% of the

variance. The semi-variogram stabilizes at the population variance at a range of just greater

than 100m. The down hole semi-variogram does not reach stability. This is to be expected

where the mineralization relates to fractures which are approximately perpendicular to the

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direction of drilling.

The zone 2 semi-variogram group shows a similar structure to the zone 1 semi-variogram;

however the range of continuity is significantly shorter, at just over 30 m.

Table 14.6: Modelled Semi-Variogram Parameters per Zone Zone Rotation (Isatis)

Nugget

C(0)

Structure 1 Structure 2

Z Y X

Range (m)

Range (m)

Strike Dip

Down

hole

Sill

(C1) Strike Dip

Down

hole

Sill

(C2)

1 40 45 0 7.687 35 35 4 15.025 105 105 12 6.049

2 30 45 0 4.166 5 5 4 7.137 32 32 15 4.166

3 40 45 0 2.083 56 56 7 2.376 5000 5000 5000 0.343

4 40 45 0 0.951 35 35 4 1.859 105 105 12 0.749

5 40 45 0 43.603 35 35 4 85.227 105 105 12 34.315

6 30 45 0 13.190 35 35 4 25.781 105 105 12 10.380

7 30 45 0 2.560 5 5 4 4.386 32 32 15 1.878

8 30 45 0 0.265 5 5 4 0.453 32 32 15 0.194

9 30 45 0 2.809 60 60 10 2.100 5000 5000 5000 0.546

10 30 45 0 3.904 5 5 4 6.689 32 32 15 2.863

11 30 0 0 1.438 10 10 9 1.926 270 270 22 9.475

51 60 35 0 4.330 135 135 11 0.5670

52 60 35 0 1.702 135 135 11 2.228

53 60 35 0 0.801 135 135 11 1.094

SRK tested for evidence of anisotropy in the plane of mineralization for Zones 1 to 10 and

51, and none of the zones where a meaningful semi-variogram could be modeled showed

any material anisotropy. As a result, the semi-variogram in the plane of the ore body is

isotropic. The down hole direction has a significantly shorter range of continuity however,

and the experimental semi-variograms and models reflect this.

The Zone 11 (laterite) data are strongly affected by topography, within the range of the semi-

variogram. The data were therefore leveled in order to remove the effect of topography. This

was achieved simply by using the top of the laterite in each drill hole as a reference point,

and translating that reference point vertically to a common elevation. In this manner, the

effect of topography on the location of samples in space was reduced, allowing for an

improved analysis of the spatial grade relationships. The experimental semi-variograms

were then calculated on the leveled dataset, with the horizontal plane representing the

direction of best continuity. This resulted in a relatively robust semi-variogram over distances

of over 200m. No significant lateral anisotropy was found in the laterite experimental semi-

variogram, and as such the lateral semi-variogram was modeled isotropically. For

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Figure 14.7: Gaussian Semi-variograms for Modelled Zones

Zone 1 Zone 2

Zone 3 Zone 9

Zone 11 Zone 53

Agbaou Gold Project Gaussian Semi-Variograms for Modeled Zones

Project No.

439430

N40

N130

0

0

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100

100

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200

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250

Distance (m)

Distance (m)

0.00 0.00

0.25 0.25

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0.75 0.75

1.00 1.00

Variogram : Au_Gauss_F


N40

N130

0

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N3014

8921

13046

14209

14600

14047

1306911872

117371111511023

1048010579

9618

1045411007128131216199327854

D-90

84

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Variogram : GAU_Au

Variogram : GAU_Au

str

8 501

502

372297

19

dip4

880

5537

199

5

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Variogram : Gau_Au

Variogram : Gau_Au

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14.8 BLOCK MODEL AND GRADE ESTIMATION After the modeling of the semi-variograms, the next step is to determine the parameters to

be used in the estimation. Quantitative Kriging Neighborhood Analysis (“QKNA”) can used to

optimize the search ranges, block sizes and number samples for Ordinary Kriging. The

QKNA methodology is discussed in Vann et al, 2003, and monitors the sensitivity of the

estimates to Kriging parameters, specifically the Slope of Regression (“SoR”), Kriging

Efficiency (“KE”), and negative weights.

The SoR is a theoretical measure of the correlation between an estimated block value and

the “actual” value. Values will be between 0 and 1, with the ideal being a slope of 1 (no

Conditional Bias). The KE is another theoretical measure calculated during the Kriging

process which indicates the degree of smoothing in the estimation. Values close to 1 are not

smoothed while those close to zero indicate a high degree of smoothing. In the case of

negative values, the global (declustered) mean of the samples is considered more reliable

than Kriging. Ordinary Kriging is based on an algorithm whereby the sum of the weights is

equal to 1 (to minimize error). In certain cases this requires negative weights to meet that

condition. A few small negative weights are acceptable but these should never exceed 5%.

14.8.1 Block Size Optimization The QKNA procedure involves iteratively Kriging selected blocks and varying selected

parameters (block size, sample numbers, search distances etc.), and then monitoring the

selected Kriging statistics and selecting the optimal parameters. SRK used QKNA to

optimize the block size and maximum number of samples. The optimization was undertaken

for zones 1 and 2, as these represent the best informed zones, for which there is a robust

semi-variogram model, and which represent a significant proportion of the volume of the

resources.

Generally a block size of less than half the grid spacing will produce unacceptable QKNA

results and there are numerous of references in the literature to the dangers of Kriging into

small blocks relative to the sample grid. In contrast Kriging into large blocks will reduce the

estimation variance; it also implies very low selectivity, which is often an unrealistic

assumption. Six block sizes were tested with QKNA, which were selected on the basis of the

ore body dimensions, drill hole spacing, and the Smallest Mining Unit (“SMU”) dimension of

2.5m3 indicated by the SRK and ENDEAVOUR Engineers. The block model was not rotated

at the request of the Mining Engineers and remains orthogonal to the main geographical

axes. The block dimensions tested are presented in Table 14.7.

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Table 14.7: Block Size Optimisation Scenarios Block option X size (m) Y size (m) Z size (m) Volume (m

3)

1 5 20 10 1000

2 5 25 10 1250

3 10 20 10 2000

4 15 20 10 3000

5 15 25 10 3750

6 10 40 10 4000

Blocks were tested for four sites in each of zone 1 and 2. The blocks were in densely and poorly samples areas of both the North and South areas of zone 1 and 2. The block Kriged for each block size was located in a similar position in each of the four scenarios, in both zones. The results of the QKNA testing for block size are presented in Figure 14.8 for zone 1 and Figure 14.9 for zone 2. The absolute values of each of the

statistics are not material for this exercise, but rather the comparison of how the block sizes

affect the statistics in each of the four zones. In these figures the values on the X axis

correspond to the ‘Block Option’ in Table 14.7.

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Figure 14.8: QKNA Results for Block Size Selection on Zone 1

Agbaou Gold Project QKNA Results for Block Size Selection on Zone 1

Project No.

439430

0.5

0.6

0.7

0.8

0.9

1

1.1

1 2 3 4 5 6

Slo

pe

of

Rre

gre

ssio

nSlope of Regression

Site 1

Site 2

Site 3

Site 4

-0.1000

0.0000

0.1000

0.2000

0.3000

0.4000

0.5000

1 2 3 4 5 6

Kri

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Kriging Efficiency

Site 1

Site 2

Site 3

Site 4

0

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1.5

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2.5

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4

4.5

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Va

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Variance

Site 1

Site 2

Site 3

Site 4

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Figure 14.9: QKNA Results for Block Size Selection on Zone 2

Agbaou Gold Project QKNA Results for Block Size Selection on Zone 2

Project No.

439430

0.5

0.6

0.7

0.8

0.9

1

1.1

1 2 3 4 5 6

Slo

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Slope of Regression

Site 1

Site 2

Site 3

Site 4

-0.1000

0.0000

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0.2000

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0.5000

1 2 3 4 5 6

Kri

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Kriging Efficiency

Site 1

Site 2

Site 3

Site 4

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Variance

Site 1

Site 2

Site 3

Site 4

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While there is no clear optimal size, the option 3 was selected as a consistently good result

for all sites. In addition, the size of 10m (X) x 20m (Y) x 10m (Z) is a suitable size when

considering the dimensions of the ore bodies, and the typical drill hole spacing.

14.8.2 Sample Number Optimization Using a similar approach to the block size optimization, the number of composites to be

used in the search neighborhood was analyzed. Several sites selected for each of Zones 1,

2, 3, 9, 11 and 51 (for which semi-variograms could be modeled) were estimated and the

results plotted.

For all test sites in all zones tested the pattern of the Kriging statistics was similar, showing

improved SoR and KE as the maximum number of samples increased. In only one site were

there any negative weights recorded, and these were for a high number of samples in the

neighborhood in Zone 1. The results of the optimization testes are plotted in Figure 14.10 to

Figure 14.13.

In selecting the optimum number of samples, one needs to find a balance between

maximizing the Kriging statistics (SoR and the KE), which are shown in the plots in Figure

14.10 to Figure 14.13 to increase consistently with the increase in sample numbers, and

avoiding over-smoothing of the estimate, which tends to occur if too many samples are

selected in the search. A measure of the degree of smoothing is the Kriging variance, which

is also plotted on the sample optimization charts.

The Zone 1 results in Figure 14.10 indicate that using upwards of 40 composites in the

estimate would provide the best estimate for a block, with negative weights only being

returned in one instance. The Kriging variance with this number of samples is low, and SRK

consider that this represents too high a number, which would result in over smoothing of the

estimate. SRK therefore selected 20 samples as the optimum number of samples in the

search, as this represented a good balance between estimation quality and retaining the

inherent variability in the estimate. Twenty composites in all cases returned a SoR of greater

than 0.8.

The Zone 2 results in Figure 14.11 show poorer quality estimation results than Zone 1 using

similar parameters, which is a result of the significantly shorter semi-variogram range (see

section 14.6). SRK selected an optimum number of samples of 15 for this zone as the

shorter search distance will limit the number of samples available in the search

neighborhood, and the reduction in the Kriging variance at sample numbers greater than this

was considered too high.

Zone 3 and Zone 9 results in Figure 14.12 show a similar pattern to the Zone 2 results,

where selecting more than 15 samples in the neighborhood would result in over smoothing,

with little improvement in the Kriging statistics. The Zone 3 and Zone 9 results reflect the

typically wider drill hole spacing in the modeled areas of these zones compared to Zone 1

and Zone 2.

For the remaining zones in the Main and Southern area, which were grouped together with

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one of the four zones discussed above, for which there is not an independent semi-

variogram model, the same parameters were used as the best developed zone in the group.

For zones 11 and 51 in Figure 14.13 an optimum number of samples of 25 was selected as

providing the best estimates.

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Figure 14.10: QKNA Results for Sample Number Selection on Zone 1

Agbaou Gold Project QKNA Results for Sample Number Selection on Zone 1

Project No.

439430

0

1

2

3

4

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Maximum number of samples

Sample number optimisation For Zone 1 North Central

Slope of the regression Z | Z* Kriging Efficiency Negative Weights Estimation variance Estimated value

0

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Sample number optimisation For Zone 1 North Dense


0

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Sample number optimisation For Zone 1 South

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Figure 14.11: QKNA Results for Sample Number Selection on Zone 2

Agbaou Gold Project QKNA Results for Sample Number Selection on Zone 2

Project No.

439430

0

1

2

3

4

5

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-2

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0

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5 10 15 20 25 30 35 40 45 50 55 60 65

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Sample number optimisation For Zone 2 South Dense


0

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Sample number optimisation For Zone 2 North


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Figure 14.12: QKNA Results for Sample Number Selection on Zone 3 and 9

Agbaou Gold Project QKNA Results for Sample Number Selection on Zone 3 and Zone 9

Project No.

439430

0

0.2

0.4

0.6

0.8

1

1.2

1.4

1.6

1.8

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Figure 14.13: QKNA Results for Sample Number Selection on Zone 11 and 51

Agbaou Gold Project QKNA Results for Sample Number Selection on Zone 11 and Zone 51

Project No.

439430

-0.25

-0.2

-0.15

-0.1

-0.05

0

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0.2

0.4

0.6

0.8

1

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5 10 15 20 25 30 35 40 45 50 55 60

Kri

gin

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tati

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s

Number of Samples

Sample Number optimisation for Zone 11 dense

Estimation Variance Slope of Regression Kriging efficiency Negative weight

Kri

gin

g e

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-0.16

-0.14

-0.12

-0.1

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Kri

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s

Number of Samples

Sample Number optimisation for Zone 11 sparse


Kri

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-0.1

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Kri

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Number of Samples

Sample Number optimisation for Zone 51 Dense


Kri

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5 10 15 20 25 30 35 40 45 50 55 60K

rig

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Sta

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Number of Samples

Sample Number optimisation for Zone 51 sparse


Kri

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14.8.3 Search Orientation Optimization Initial testing of the search parameters based on the testing above and the semi-variogram

ranges modeled in section 14.7 revealed that the semi-variogram orientations, while

appropriate as an average for the entire zone, produced sub-optimal results in some areas,

because of local changes in dip and strike. SRK identified a location in the North and

Southern areas the dip and strike of most zones showed some change. Likewise in the

western area two locations were identified where there was a change in the orientation of the

ore body.

In

Figure 14.14 the search domains identified in the Main and South areas are illustrated. Two

polygons are drawn around the Southernmost search area (south end of the South area),

and around the central portion of the Main area. The areas outside the polygons are also

treated as separate search domains, termed the Main domain (North end of the Main area),

and South/Central domains. For each of the four domains, for each of the zones which exist

within the domain, an average ore body orientation was approximated and applied to the

estimates within this search domain. Within a zone, no restrictions were placed on the

composites available to estimate within a search domain (i.e. composites within the Main

domain could be used to estimate samples in the North end of the Main area search

domain) however blocks within a search domain were estimated with the search orientation

for that domain.

After the first estimation, using the optimized parameters, and search ranges which are

based on the semi-variogram range, parts of the ore bodies remained un-estimated,

particularly in the zones with short range semi-variograms such as the Zone 2 group. A

second and third search pass were designed, with longer search ranges, to ensure that all

blocks within the ore bodies were informed.

In Figure 14.15, the search domains used for the West area are illustrated. The three

polygons shown represent the three search domains and as with the North, Central and

South areas the search orientations were optimized for each zone within each search

domain. The three zones are identified as East, West and Central. A second search pass

was necessary to inform all blocks.

The laterite in Zone 11 is strongly affected by the topography, and SRK elected to use a

Datamine feature termed ‘dynamic anisotropy’ which allows the search orientation to be

dynamically altered for each block according to the local orientation of the ore body. The top

surface of the laterite is used to calculate a local orientation for the ore body. This orientation

is then estimated into the block model using inverse distance, and the orientation is used to

rotate the search ellipse and the semi-variograms used in the estimates of each block. Only

one search pass was required to inform all blocks within the zone.

The final search neighborhood parameters used in the Ordinary Kriging estimate, based on

the semi-variogram models and the testing described above are summarized in

Table 14.8. For all estimates, a discretization pattern of 5 (X) by 10 (Y) by 5 (Z) points was

used.

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Figure 14.14: Search Domains

Figure 14.15: Search Domains


Technical Report

Search Domains Main, Main/Central, South/Central and South

Not to scale

Agbaou Gold Project Search Domains Main, Main/Central, South/Central

and South

Search Domains West for Zones 51 to 53

Agbaou Gold Project Search Domains West

Page 14-30

and South for Zones 1 to 10

Search Domains Main, Main/Central, South/Central Project No.

439430

Project No.

439430

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Table 14.8: Summary of Search Neighbourhood Parameters

First Search pass Second Search pass Third Search pass

Rotation (Degrees

clockwise) Search Range (m)

Number of

samples Search Range (m)

Number of

samples Search Range (m)

Number of

samples

Zone Search Domain Z Y X X Y Z Min Max X Y Z Min Max X Y Z Min Max

Zone1 North 50 45 0 110 110 15 5 25 200 200 30 5 20 250 250 30 5 20

Zone1 North/Central 30 60 0 110 110 15 5 25 200 200 30 5 20 250 250 30 5 20

Zone1 South 35 65 0 110 110 15 5 25 200 200 30 5 20 250 250 30 5 20

Zone1 South/Central 55 55 0 110 110 15 5 25 200 200 30 5 20 250 250 30 5 20

Zone2 North 50 55 0 38 38 18 5 20 70 70 40 5 15 200 200 40 5 15


Zone2 South 40 45 0 38 38 18 5 20 70 70 40 5 15 200 200 40 5 15


Zone3 North 45 50 0 60 60 10 5 20 120 120 20 5 15 200 200 40 5 15

Zone3 South 40 60 0 60 60 10 5 20 120 120 20 5 15 200 200 40 5 15


Zone4 North 50 50 0 110 110 15 5 25 200 200 30 5 20 120 120 20 5 15


Zone4 South 45 45 0 110 110 15 5 25 200 200 30 5 20 250 250 30 5 20


Zone5 North 50 60 0 110 110 15 5 25 200 200 30 5 20 250 250 30 5 20

Zone6 North 50 65 0 110 110 15 5 25 200 200 30 5 20 250 250 30 5 20


Zone7 North 50 65 0 38 38 18 5 20 70 70 40 5 15 250 250 30 5 20


Zone8 North 65 75 0 38 38 18 5 20 70 70 40 5 15 200 200 40 5 15

Zone9 North 50 55 0 65 65 15 5 20 130 130 30 5 15 200 200 40 5 15



Zone 10 North 55 55 0 38 38 18 5 20 70 70 40 5 15 200 200 30 5 15

Zone 10 North/Central 55 55 0 38 38 18 5 20 70 70 40 5 15 200 200 30 5 15

Zone 11 Dynamic 280 280 10 5 25 - - - - - - - - - -

Zone 51 East 38 42 0 140 140 11 5 25 200 200 30 5 25 - - - - -

Zone 51 Central 68 35 0 140 140 11 5 25 200 200 30 5 25 - - - - -

Zone 51 West 45 35 0 140 140 11 5 25 200 200 30 5 25 - - - - -

Zone 52 East 38 42 0 140 140 11 5 25 200 200 30 5 25 - - - - -

Zone 52 Central 68 35 0 140 140 11 5 25 200 200 30 5 25 - - - - -

Zone 52 West 45 35 0 140 140 11 5 25 200 200 30 5 25 - - - - -

Zone 53 East 38 42 0 140 140 11 5 25 200 200 30 5 25 - - - - -

Zone 53 Central 68 35 0 140 140 11 5 25 200 200 30 5 25 - - - - -

Zone 53 West 45 35 0 140 140 11 5 25 200 200 30 5 25 - - - - -

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14.8.4 Kriging Parameters Ordinary Kriging was used as the interpolation method for all zones at Agbaou. The search

neighborhood used was determined as described in section 14.8.3 and detailed in

Table 14.8. Kriging was done into the full parent cells, of 10m (X) and 20m (Y) and 1m (Z).

Sub-celling of the block model was done to a minimum size of 2.5m in all directions, to honor

the mineralization boundaries, and to best approximate the mineralized volume.

The 2m composites were used for the estimate, with high values capped as described in

section 14.6, and no restriction of samples within each Zone. Separate search domains were

used to best honor the local orientation of the ore bodies for the North/Central/South and the

West areas. The laterite (Zone 11) used dynamic anisotropy to optimize the local search

orientation. Up to three search passes were required to ensure that all blocks within a

domain were estimated. The search pass used to inform the blocks was coded into the block

model and used as one of the inputs during the resource classification process.

14.8.5 Model Validation A number of checks were conducted on the Kriging results to validate the estimates. These

include visual validations of the estimates on a section by section basis, global statistical

comparisons of the data on a zone by zone basis, as well as swath analyses which compare

the estimates in slices of the ore body with the data within the slice. The checks done are

summarized in this section (section plots are included in SRK Consulting, SENET, Knight

Piésold Consulting, 2012 Appendix 14-A).

14.8.5.1 Input Grades Versus Estimated Grades

The Ordinary Kriged means from the block model were compared to the length weighted

means of the sample grade within each zone, the composite grade and capped composite

grades within each zone. The statistics and comparisons are summarized in Table 14.9.

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Table 14.9: Samples and Composites Versus Kriged Means Per Zone Zone No.

Samples

Sample

Grade

No.

Comps

Comp

Grade

Capped

Grade

Estimate

Grade

Sample

vs. Comp

Comp vs.

Cap

Cap vs.

Estimate

1 2,279 3.00 1,225 3.00 2.90 2.65 0% 3% 9%

2 2,033 2.10 1,028 2.11 2.04 1.80 0% 3% 12%

3 307 1.41 163 1.41 1.35 1.28 0% 5% 5%

4 638 1.05 338 1.03 1.03 0.98 2% 0% 5%

5 282 6.75 165 6.75 6.12 6.18 0% 9% -1%

6 256 4.42 135 4.42 4.27 3.78 0% 3% 12%

7 584 1.63 292 1.63 1.60 1.52 0% 2% 5%

8 67 0.77 23 0.78 0.78 0.59 -2% 0% 24%

9 614 0.99 285 0.99 0.94 1.08 1% 4% -15%

10 179 2.14 115 2.14 1.96 1.53 0% 8% 22%

11 1,665 0.77 1,109 0.77 0.72 0.54 0% 6% 25%

51 353 1.64 236 1.69 1.69 1.63 -3% 0% 4%

52 58 0.82 48 0.83 0.83 0.98 -1% 0% -18%

53 42 0.80 30 0.90 0.90 0.90 -13% 0% 0%

All 7239 2.43 3769 2.43 2.33 2.00 0% 4% 14%

For most zones the compositing has little effect on the mean grade, aside from the West

zone, where the composite grade is slightly higher than the original sample grade. The effect

of capping (where it was applied) is generally less than 5% aside from zones 5, 10 and 11,

which is discussed in detail in section 14.6.

The estimated grade is in all but a few cases lower than the capped composite grade. The

zone 5 estimate is marginally higher than the capped composite grade; however for Zone 9

and Zone 52 the estimated grade is significantly higher than the capped composite grade.

For Zone 52 this is attributed to the very small number of composites, which is insufficient to

adequately inform the estimate. For Zone 9 the drilling spacing is relatively wide, and there is

a large high grade zone in the main area of the zone, which is leading to the higher grade.

In Zone 2, the estimated grade is lower than the capped composite grades by approximately

12%. SRK interpret this to result from the higher grade in the most densely drilled area at the

core of the deposit, and the lower grades on the more poorly drilled periphery. For the other

zones which have a significant underestimate, it is a result of a similar pattern, combined

with a small number of composites, such as Zone 8.

14.8.5.2 Swath Analysis

The swath analysis plots are generated by slicing the composite and block model files in

50m increments along strike, along dip and along the Z elevation per zone, as well as

globally. The strike and dip plots are based on a rotated co-ordinate system in order to align

the slices with the orientation of the ore body. In the plots, the Y axis is equivalent to the

strike direction and the X axis is equivalent to the dip direction.

Swath plots are presented in

Figure 14.16 to 14.20 for all zones in the Main and South areas, for Zones 1 and 2 and for

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Zone 11 as examples, but swath plots were generated and analyzed for all zones. In the

swath plots, the average grade of the 2m composites, before and after capping are plotted

as g/t and g/t_Cap, the estimated grade is plotted as g/t*, along with the total tonnage

(in Mt) and the number of composites in each slice.

The Kriged estimates are by their nature expected to be smoother than the composite

values, however the general trend of the composite grade are expected to be matched by

the estimates. In all zones, the estimate never matches the peaks and troughs of the

composite grades; however the general trend of the estimates is honored. In some of the

relatively poorly sampled zones the correlation is poorer, however the trends are honored,

and the correlation is in line with what is expected given the number of composites available

in the estimate.

SRK consider that overall the estimates are consistent with the source data and adequately

model the grade distributions.

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Figure 14.16: Swath Plots of the Main and South Areas for Zones 1 – 10

Agbaou Gold Project Main and South Area Swath Plots – All Zones

Project No.

439430

858

548

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378 374

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Mt g/t* g/t g/t_Cap No. Composites

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31

33

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Elevation Swath analysis


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Figure 14.17: Swath Plots of the Main and South Areas for Zone 1

Agbaou Gold Project Main and South Area Swath Plots – Zone 1

Project No.

439430

8

51

40

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7

63 34

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Figure 14.18: Swath Plots of the Main and South Areas for Zone 2

Agbaou Gold Project Main and South Area Swath Plots – Zone 2

Project No.

439430

64

36

1

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Figure 14.19: Swath Plots of the Laterite (Zone 11)

Agbaou Gold Project Laterite Swath Plots – Zone 11

Project No.

439430

14.9 RECOVERABLE RESOURCE MODELLING Recoverable resource calculations were completed for zones that were sufficiently well

informed, and for which there was sufficient confidence in the estimates to allow this. This

includes zones 1, 2, 3, 4, 7, and 9, which (excluding the laterite) make up 90% of the volume

of the Main and South area. The remaining domains were too small and/or had a Kriged

1 3 1 2 5 3 4 2 1 3 2

95 7 5 4 3 6

11

6 7

3 2

7 5 4 2 15

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X Axis Swath analysis

Mt g/t g/t_Cap g/t* No. Composites

2 2 3

14

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Mt g/t g/t_Cap g/t* No. Composites

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estimate that was too poor to be used. The SMU used is 2.5m by 2.5m by 2.5m. Two

methods of recoverable resource estimation were used; Uniform Conditioning (“UC”) and

Multiple Indicator Kriging (”MIK”). Ultimately the MIK estimate was selected for reporting as

this was consistent with the previous mineral resource estimates of the Agbaou deposit done

by Coffey Mining.

14.9.1 Uniform Conditioning Geostatistics provide some powerful methods to achieve change of support, i.e. calculate

the block grades distribution from the sample grades distribution. The global grade-tonnage

curves can then be calculated for any SMU support. UC aims to condition the local grade-

tonnage curves for each panel. The discrete Gaussian model is then applied, basically

assuming, first, that the SMUs constitute a partition of the panels, and second, the Gaussian

values (after normal score transform) for SMUs and panels have a bivariate normal

distribution, only depending on change of support coefficients.

The steps undertaken in Isatis to achieve the uniform conditioning are as follows:

• Decluster the domain drill hole samples to the Kriged mean; • Gaussian anamorphosis of the domain drill hole samples; • Rescale variogram model using declustered weights to the domain sample variance; • Support correction calculation using information effect; • Uniform conditioning; • Export files to Surpac for mine planning and pit optimization.

Support correction was repeated for the panel and the block without information effect for

comparison of the theoretical curves in Isatis, though only the ‘block with information effect’

was used in the UC.

14.9.1.1 Declustering

Gaussian anamorphosis modeling requires that the composites have been declustered, as

the declustered weights are needed to correctly reproduce the underlying distribution.

Declustering was undertaken for each zone, aiming to generate a declustered mean that

matches the Kriged mean of the domain. The declustering grid was selected to approximate

the Kriged grade for the zone. The declustering results are presented in Table 14.10.

Table 14.10: Declustering Statistics for Gaussian Anamorphosis Modelling Zone

Kriged mean Declustered mean Declustered

Standard Dev.

Declustered

Variance Declustering grid

1 2.72 2.7247 5.3629 28.7607 100 x 100 x 20

2 1.83 1.8344 3.7893 14.3588 50 x 50 x 12

3 1.33 1.3611 2.1914 4.8022 5 x 5 x 0.7

4 0.98 1.0311 1.8865 3.5589 5 x 5 x 1

7 1.56 1.5298 2.9706 8.8245 10 x 10 x 2

9 1.16 1.1619 2.3356 5.4550 70 x 70 x 15

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14.9.1.2 Gaussian Anamorphosis Modeling

The Gaussian anamorphosis modeling in Isatis transforms the raw gold grade variable into a

gaussian variable using a normal score transformation. The transformation is achieved using

hermite polynomials to model the grade distribution. The Gaussian variables are tested to

ensure that they approximate a Gaussian distribution whereby the mean should be close to

zero; the variance and standard deviation close to one; and the minimum and maximum

should be very similar in the positive and negative. The statistics of the Gaussian variables

are presented in Table 14.11, along with the number of polynomials used in the

transformation. In all cases the attributes of a Gaussian distribution have been approximately

achieved.

Table 14.11: Statistics of the Gaussian Variables

Zone Number of Statistics on the Gaussian Variable

polynomials Min Max Mean Variance Std. Dev.

1 30 -3.70737 3.305584 -0.000051 1.001077 1.000538

2 30 -3.26988 3.102155 -0.000265 0.998151 0.999075

3 30 -2.58464 2.584637 0.000000 0.984066 0.992001

4 30 -2.82041 2.820412 0.000000 0.992113 0.996049

7 30 -2.83822 2.838221 0.000005 0.992546 0.996266

9 30 -2.86427 3.140847 0.002880 1.008022 1.004003

14.9.1.3 Support Correction

The Support Correction functionality is designed to model the histogram of blocks from a

point anamorphosis and a variogram model. The method is based on the Gaussian Discrete

Model, and in this case, the information effect has also been used. The information effect

takes into account the number and spacing of drill holes (grade control drilling or blast holes)

that will be available at the time of mining. The grade control grid assumed for this study is

on a 5m (strike) x 8m (dip) x 1m sample length, which is based on the current grade control

drilling at an operating mine nearby to Agbaou. Seven holes in the X and Y axes, and 15

samples in the Z direction were assumed for the study.

The support correction parameters are presented in Table 14.12.

Table 14.12: Support Correction Parameters

Zone Gamma

(v,v)

Real Block

Variance

Real Block

Support

Correction (r)

Used Kriged

Block

Variance

Kriged Block

Support

Correction (s)

Used

Kriged-Real

Block

Covariance

Kriged-Real Block

Support

Correction

1 12.594 15.791 0.8377 14.717 0.8183 14.740 0.9779

2 7.4934 6.6694 0.8148 5.6880 0.7746 5.6861 0.9504

3 2.4893 2.2935 0.7724 2.1279 0.7504 2.1311 0.9727

4 1.5737 1.9732 0.8311 1.8452 0.8121 1.8456 0.9773

7 4.9407 3.8410 0.7682 3.1628 0.7160 3.1614 0.9317

9 2.9733 2.3010 0.7890 2.1276 0.7681 2.1314 0.9747

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14.9.1.4 Change of Support Grade Tonnage Curves

The change of support grade-tonnage curves for each domain can be calculated within Isatis

to check the relationship between the actual Kriged panel estimates, the Gaussian

anamorphosis of the panel, block and block with information effect. The Ordinary Kriged

panel estimates and the theoretical grade tonnage curves are presented in

Figure 14.20 to

Figure 14.22 for the six UC zones.

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Figure 14.20: Grade Tonnage Curves for Zone 1 and Zone 2

Zone 1

Zone 2

Agbaou Gold Project Grade Tonnage Curves

Project No.

439430

0.0

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Zone_OK

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Zone 3

Zone 4


Project No.

439430

0.0

0.0

0.5

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Zone 7

Zone 9


Project No.

439430

0.0

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The grade tonnage curves show that the OK estimate generally underestimates the selective

grade, but also overestimates the selective tonnage at low grades, and underestimates the

selective tonnage at higher grades. The Information Effect has a small influence in all cases.

14.9.1.5 Uniform conditioning

The Uniform Conditioning was undertaken with a SMU size of 2.5m x 2.5m x 2.5m, at

fourteen cut-off grades: 0.1; 0.2; 0.3; 0.4; 0.5; 0.6; 0.7; 0.8; 1.0; 1.2; 1.4; 1.6; 1.8; and 2.0.

Block anamorphosis with information effect was used and no tonnage corrections were

applied. No tonnage correction was applied in the calculation of the UC, however the results

were post processed to take account of small tonnage proportions within panels. The aim of

this is to provide some practicality to the final block model, which contains small blocks

(minimum block size allowed in the sub-celling of the parent blocks to model the volume

correctly was 2.5m square; this is coincident with the SMU size). Parent blocks which have

small proportions of their tonnes inside the wireframes will still have a UC estimate; however

the practical selectivity that can be applied to these small tonnages is likely to be

overestimated.

In order to correct for this, the UC selectivity of any panel (parent block in Datamine) which

has less than 20% of its volume within the wireframe, was reset. Simply, the proportion of

the panel above the cut-off threshold was set to either one or zero, if the OK grade was

respectively above or below the cut-off, and the grade above cut-off was set to the OK

grade. For the zones that did not have a UC estimate (Zones 5, 6, 8, 10, 11, and 51 to 53) a

similar process was undertaken, whereby the grades and tonnage proportion at each cut-off

were set to the OK grades and to one or zero if the OK grades were above or below the cut-

off. All zones could therefore be reported and processed on a like for like basis, at each cut

off.

14.9.1.6 Uniform Conditioning Validation

In addition to the cut-offs detailed in the section above, the UC was calculated at 0 g/t but

not used in the final model, to determine if there are any inconsistencies between the UC

and the OK estimate. The 0 g/t UC estimates were plotted against the OK estimates in a

scatter plot to test for this. All the UC estimates at 0 g/t were identical to the OK estimates,

for all zones, and no corrections had to be applied.

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Table 14.13: Comparison of the OK and UC Estimation Results at a 0.5 g/t Cut-Off OK Model UC Model Percentage Difference

Zone Mt Au (g/t) Au (kOz) Mt Au (g/t) Au (kOz) Mt Au (g/t) Au (kOz)

1 4.52 2.66 386.2 3.86 3.06 379.9 -14.5% 15.1% -1.6%

2 4.29 1.85 255.7 3.75 2.08 250.5 -12.7% 12.2% -2.0%

3 0.59 1.61 30.4 0.54 1.74 30.5 -7.1% 8.4% 0.6%

4 1.48 1.03 48.8 1.03 1.36 45.3 -30.1% 32.8% -7.2%

7 1.42 1.70 77.8 1.17 2.02 76.1 -17.3% 18.3% -2.2%

9 1.24 1.24 49.5 1.04 1.44 48.3 -16.1% 16.3% -2.4%

Total 13.54 1.95 848.4 11.41 2.27 830.7 -15.8% 16.2% -2.1%

Table 14.14: Comparison of the OK and UC Estimation Results at a 1.0 g/t Cut-Off

14.9.2 Multiple Indicator Kriging MIK works on a probabilistic basis to define the distribution of the grades of samples within

each search window, providing a discrete approximation to the cumulative density function

(“cdf”) for each block. As this distribution is based on the samples found within the search

window centered on any given point, it changes from block to block to reflect local grade

variability.

A series of binary indicator values is calculated for a set of cut-offs for each composite within

the wireframe. The indicator value is one if the grade of the composite is greater than or

equal to the indicator cut-off value, and zero if it is less than that value.

The Kriging of the indicators generates conditional probabilities for a restricted set of

specified cut-offs. Starting from these probabilities, the conditional cdf for each grid node can

be re-built, in order to derive the tonnage and metal quantity for a new set of cut-offs. One

can then calculate the probability for the variable to exceed any cut-off or to calculate the

average value of the variable above or below this cut-off, accounting for a possible change

of support.

OK Model UC Model Percentage Difference

Zone Mt Au (g/t) Au (kOz) Mt Au (g/t) Au (kOz) Mt Au (g/t) Au (kOz)

1 4.18 2.81 377.0 3.05 3.67 360.6 -26.9% 30.8% -4.3%

2 3.24 2.22 231.2 2.67 2.62 225.0 -17.4% 17.8% -2.7%

3 0.34 2.25 24.6 0.33 2.40 25.6 -2.2% 6.4% 4.1%

4 0.60 1.41 27.2 0.52 1.98 33.1 -13.3% 40.3% 21.6%

7 0.92 2.23 65.9 0.76 2.73 66.4 -17.8% 22.6% 0.7%

9 0.70 1.62 36.5 0.57 2.03 37.3 -18.2% 25.0% 2.3%

Total 9.97 2.38 762.4 7.91 2.94 748.0 -20.7% 23.7% -1.9%

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The steps followed in the MIK estimate were as follows:

• Calculate the indicator variables for each cut-off; • Calculate and model the semi-variograms for each indicator for each zone; • Krige each of the indicator variables for each zone; • Post process the estimates to model the conditional cdf and apply the change of

support correction; and • Export files to Surpac for mine planning and pit optimization.

14.9.2.1 Statistical Analysis and Variography

A set of nineteen cut-off indicators were generated in the composite file and these were used

as the basis of the MIK estimation. Experimental semi-variograms were calculated for each

indicator, in each zone and these were modeled, with reference to the grade semi-

variograms modeled for the OK estimate. For Zones 3, 4, and 7 the experimental semi-

variograms did not show an easily interpretable structure for most of the indicators, and after

comparison with the experimental semi-variograms for zones 1, 2 and 9, SRK elected to use

the indicator semi-variograms from better informed zones, in a similar manner to the OK

estimate. Zone 3 was found to be most similar to Zone 9, while Zone 7 was matched with

Zone 2 and Zone 4 with Zone 1.

The modeled semi-variogram parameters are detailed in Table 14.15, Table 14.16 and Table

14.17 and the semi-variograms are presented in Figure 14.23, Figure 14.24 and Figure

14.25. Note that only the semi-variograms for indicators ranging from 0.1 to 3 g/t are

displayed in these figures, for presentation reasons. The remaining semi-variograms show

similar structures, and the models are detailed in Table 14.15, Table 14.16 and Table 14.17.

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Figure 14.23: Zone 1 Indicator Semi-Variograms for Indicators up to 3 g/t

Agbaou Gold Project Zone 1 Indicator Semi-Variograms for Indicators up to 3 g/t

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0.0 0.0

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0.3 0.3 Variogram : Au_MIK{0.500000}


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N3095

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14831157

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0.00 0.00

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0.06 0.06

0.07 0.07

0.08 0.08



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Table 14.15: Zone 1 Semi-Variogram Model Parameters C

ut-

off

Rotation (Isatis)

Nugget

C(0)


Z Y X

Range (m)

Range (m)

Strike Dip Down

hole

Sill

(C1) Strike Dip

Down

hole

Sill

(C2)

Grade Variography

- 40 45 0 7.687 35 35 4 15.025 105 105 12 6.049

Indicator Variography

0.1 40 45 0 0.034 115 115 5 0.037

0.2 40 45 0 0.061 115 115 5 0.049

0.3 40 45 0 0.084 115 115 4 0.051

0.4 40 45 0 0.113 120 120 4 0.061

0.5 40 45 0 0.144 120 120 5 0.061

0.6 40 45 0 0.165 136 136 6 0.054

0.7 40 45 0 0.182 136 136 6 0.049

0.8 40 45 0 0.193 139 139 21 0.045

1.0 40 45 0 0.203 119 119 18 0.042

1.2 40 45 0 0.203 69 69 18 0.047

1.4 40 45 0 0.203 69 69 18 0.047

1.6 40 45 0 0.194 55 55 22 0.051

1.8 40 45 0 0.181 45 45 18 0.061

2.0 40 45 0 0.175 37 37 15 0.061

3.0 40 45 0 0.158 37 37 15 0.061

5.0 40 45 0 0.111 32 32 13 0.047

7.0 40 45 0 0.094 32 32 13 0.022

10 40 45 0 0.094 32 32 13 0.022

15 40 45 0 0.094 32 32 13 0.022

All the zone 1 semi-variograms have been fitted with a single structured spherical semi-

variogram. The ranges of continuity are best for the lower grade indicators, below 1g/t, and

the ranges decrease to 32m for the highest grade indictors. The low grade indicators have a

reasonable nugget proportion of approximately 50% to 70%, while the higher grade

indicators have high nuggets at up to 83% of the total sill.

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ut-

off

Rotation (Isatis)

Nugget

C(0)


Z Y X

Range (m)

Range (m)

Strike Dip Down

hole Sill (C1) Strike Dip

Down

hole

Sill

(C2)

Grade Variography

- 30 45 0 4.166 5 5 4 7.137 32 32 15 4.166


0.1 30 45 0 7.687 8 8 4 0.059 35 35 12 0.012

0.2 30 45 0 4.166 3 3 5 0.087 61 61 15 0.023

0.3 30 45 0 2.083 5 5 6 0.070 81 81 13 0.023

0.4 30 45 0 0.951 3 3 7 0.055 97 97 16 0.037

0.5 30 45 0 43.60 6 6 6 0.079 250 250 18 0.033

0.6 30 45 0 13.19 6 6 7 0.071 263 263 25 0.031

0.7 30 45 0 2.560 6 6 7 0.067 185 185 18 0.033

0.8 30 45 0 0.265 8 8 7 0.095 231 231 22 0.031

1.0 30 45 0 2.809 8 8 8 0.100 150 150 14 0.029

1.2 30 45 0 3.904 8 8 8 0.088 110 110 15 0.035

1.4 30 45 0 1.438 9 9 7 0.094 93 93 21 0.031

1.6 30 45 0 1.438 10 10 8 0.086 73 73 17 0.026

1.8 30 45 0 1.438 9 9 6 0.073 58 58 34 0.042

2.0 30 45 0 1.438 7 7 5 0.079 50 50 30 0.044

3.0 30 45 0 1.438 7 7 5 0.079 51 51 30 0.040

5.0 30 45 0 1.438 7 7 5 0.020 44 44 30 0.040

7.0 30 45 0 1.438 7 7 5 0.002 31 31 21 0.032

10 30 45 0 1.438 7 7 5 0.002 28 28 19 0.023

15 30 45 0 1.438 5 5 3 0.002 35 35 23 0.012

All the zone 2 semi-variograms have been fitted with two structured spherical semi-

variograms. The ranges of continuity are best for the indicators between 0.5 and 1 g/t, and

the high grade indicators again have relatively low ranges of continuity. All the fitted semi-

variograms have a very short first range which accounts for a significant proportion of the

total variance above the nugget. The low grade indicators have a relatively low nugget

proportion of approximately 30% to 50%, while the higher grade indicators have high

nuggets at up to 98% of the total sill for the 10 and 15 g/t indicators, which is a result of

having fewer than ten composites greater than 5 g/t in the zone dataset.

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ut-

off

Rotation (Isatis)

Nugget

C(0)


Z Y X

Range (m)

Range (m)

Strike Dip Down

hole Sill (C1) Strike Dip

Down

hole

Sill

(C2)

Grade Variography

- 30 45 0 2.809 60 60 10 2.100 5000 5000 5000 0.546


0.1 30 45 0 7.687 13 13 8 0.060 103 103 18 0.032

0.2 30 45 0 4.166 4 4 6 0.068 134 134 6 0.032

0.3 30 45 0 2.083 5 5 7 0.070 115 115 7 0.034

0.4 30 45 0 0.951 15 15 9 0.049 95 95 10 0.078

0.5 30 45 0 43.60 14 14 8 0.048 108 108 12 0.071

0.6 30 45 0 13.19 14 14 8 0.048 108 108 12 0.066

0.7 30 45 0 2.560 14 14 8 0.038 105 105 11 0.069

0.8 30 45 0 0.265 14 14 8 0.031 95 95 10 0.069

1.0 30 45 0 2.809 14 14 8 0.031 89 89 9 0.070

1.2 30 45 0 3.904 14 14 8 0.031 59 59 6 0.069

1.4 30 45 0 1.438 23 23 4 0.007 60 60 10 0.095

1.6 30 45 0 1.438 19 19 3 0.016 62 62 8 0.068

1.8 30 45 0 1.438 19 19 6 0.016 65 65 8 0.054

2.0 30 45 0 1.438 19 19 6 0.016 65 65 8 0.054

3.0 30 45 0 1.438 5 5 7 0.008 80 80 12 0.013

5.0 30 45 0 1.438 5 5 7 0.008 80 80 12 0.013

7.0 30 45 0 1.438 5 5 7 0.008 80 80 12 0.013

10 30 45 0 1.438 5 5 7 0.008 80 80 12 0.013

15 30 45 0 1.438 5 5 7 0.008 80 80 12 0.013

All the Zone 9 semi-variograms have been fitted with two structured spherical semi-

variograms. The ranges of continuity are best for the lowest grade indicators below 1 g/t,

while the higher grade indicators show reasonable ranges of continuity. All the fitted semi-

variograms have a very short first range which accounts for a significant proportion of the

total variance above the nugget. The low grade indicators have a relatively high nugget

proportion of approximately 50% to 80%, while the higher grade indicators have high

nuggets at up to 98% of the total sill for the above 3 g/t indicators, which is a result of the

paucity of composites with grades of greater than 5 g/t in the zone dataset.

14.9.2.2 MIK Estimation

The block optimization exercise described in section 14.8.1 was used as the basis of the

selection of the block size. The block size is therefore the same as used in the OK and UC

estimates. Similarly, the search strategy and sample number optimization results described

in sections 14.8.3 and 14.8.2 respectively were considered in selecting the same parameters

as the OK and UC estimates. The search parameters used in each of the six MIK zones are

listed in

Table 14.8. The same search was used for all indicators in order to minimize the

inconsistencies which can arise from using different search strategies for each indicator. As

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with the OK estimate, a three pass search strategy was required to estimate all blocks.

The Kriging of the indicators generates conditional probabilities for a restricted set of

specified cut-offs. In order to transform the probability estimates into grade and tonnage

proportions, the local conditional cfd must be rebuilt for each block. The conditional cdf will

allow the calculation of the probability for the estimated block grade to exceed any cut-off or

to calculate the average value of the variable above or below this cut-off, accounting for a

possible change of support.

Two methods of modeling the conditional cdf are available in Isatis, the Affine correction and

the Indirect Correction through Permanence of Lognormal Distribution. The Affine correction

uses a linear relationship to transform a value of one distribution to a value of another

distribution, and the variance of the initial distribution is reduced as the mean value remains

unchanged. The data distribution is then squashed towards the mean value, but its global

shape is kept unchanged. SRK used the Indirect Correction through Permanence of

Lognormal Distribution approach (which is suitable for positively skewed distributions such

as a typical gold distribution), where the values are transformed from one distribution to a

value of a lognormal distribution using an exponential relationship. The variance of the initial

distribution is reduced as the mean value remains unchanged. The extreme values are

squashed towards the mean value more than the median values, the skewness of the

distribution decreasing and the symmetry increasing.

Two methods for processing the estimates were used for comparative purposes; an e-type

estimate which calculates the mean grade for each block, and does not allow for selectivity

which is proposed for production at Agbaou; and an adjustment of the histogram variance by

a variance adjustment factor. The variance adjustment factor is calculated from the

dispersion variance, which is derived from the semi-variogram models. The variance

adjustment ratios calculated for each zone are presented Table 14.18. To be consistent with

the Coffey MIK estimate, no information effect was applied in the support correction. The

same SMU size of 2.5m x 2.5m x 2.5m used in the UC estimate was applied in the MIK

change of support calculation.

Table 14.18: Volume Variance Ratios for Each MIK Zone Zone Zone1 Zone2 Zone3 Zone4 Zone7 Zone9

Volume Variance Reduction Factor 0.5431 0.3800 0.4728 0.5485 0.4189 0.4010

14.9.2.3 Model Validation

The MIK e-type estimates were compared to the OK grades on a block by block basis for the

estimated zones. The MIK e-type means generally show a good correlation with the OK

estimates, however the MIK means are higher than the OK estimates in all zones. In three of

the zones there are also anomalous values in some of the peripheral blocks estimated with

the second and third search passes, in which the MIK estimate at all cut-off values has been

calculated to the maximum grade defined in the indicator post-processing.

These blocks are all on the margins of the ore bodies, and are poorly informed. The MIK

estimates of these blocks are not considered reliable as the indicators estimated are greater

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than 1 (all indicators should be between one and zero), and the values in these blocks have

been reset to the OK values, with no selectivity applied. These blocks have been masked in

order to calculate the statistics presented in Table 14.19. Zones 2 and 3 were principally

affected by this correction, and this results in a minor decrease in the average grade of these

zones.

In addition, and in a similar manner to the processing of the UC estimates, the blocks which

have only a small proportion (<20%) of the parent block inside the wireframes have had the

MIK estimates reset to the OK grades, and only whole panel selectivity is applied.

Table 14.19: Statistics of OK Estimates and MIK e-type Means Zone Estimate Count Minimum Maximum Mean Variance Correlations

1 OK 4529 0.24 12.91 2.51 2.321 0.968

MIK e-type 4529 0.26 12.87 2.76 3.176

2 OK 3956 0.02 11.18 1.72 1.519 0.964

MIK e-type 3956 0.08 12.53 1.79 1.842

3 OK 1247 0.01 6.110 1.25 1.184 0.958

MIK e-type 1247 0.08 8.650 1.42 1.683

4 OK 1861 0.06 5.660 1.05 0.319 0.903

MIK e-type 1861 0.12 5.660 1.15 0.416

7 OK 1047 0.09 6.780 1.42 1.075 0.979

MIK e-type 1047 0.13 7.820 1.51 1.473

9 OK 1745 0.08 3.600 1.08 0.374 0.834

MIK e-type 1745 0.13 7.010 1.35 1.145

The MIK estimates were compared to the UC estimates for each zone, and grade tonnage

curves calculated for all zones. The MIK estimate typically models a greater degree of

selectivity in most zones, returning a lower tonnage and higher grade than the UC estimates.

At higher cut-off values (over 1g/t) the UC estimate at times reported a lower tonnage than

the MIK estimate, but still at a lower grade. The shape of the grade tonnage curves was

generally similar between the MIK and UC estimates. Grade tonnage curves for zones 1 and

2 are presented in Figure 14.26 for the MIK and UC estimates. They reflect the trends seen

in all zones, with the MIK returning higher grades and lower tonnages.

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Figure 14.26: Grade Tonnage Curves Comparison Between MIK and UC for Zones 1 and 2

Agbaou Gold Project Grade Tonnage Curve Comparisons

Project No.

439430

0.00

1.00

2.00

3.00

4.00

5.00

6.00

0

500,000

1,000,000

1,500,000

2,000,000

2,500,000

3,000,000

3,500,000

4,000,000

4,500,000

5,000,000

0 0.1 0.2 0.3 0.4 0.5 0.6 0.7 0.8 1 1.2 1.4 1.6 1.8 2

Gra

de

To

nn

es

Cut_Off

Zone 1

Tonnes MIK

Tonnes UC

Grade MIK

Grade UC

0.00

0.50

1.00

1.50

2.00

2.50

3.00

3.50

4.00

4.50

0

500,000

1,000,000

1,500,000

2,000,000

2,500,000

3,000,000

3,500,000

4,000,000

4,500,000

5,000,000

0 0.1 0.2 0.3 0.4 0.5 0.6 0.7 0.8 1 1.2 1.4 1.6 1.8 2

Gra

de

To

nn

es

Cut_Off

Zone 2

Tonnes MIK

Tonnes UC

Grade MIK

Grade UC

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14.10 MINERAL RESOURCE CLASSIFICATION Block model quantities and grade estimates for the Agbaou Project were classified according

to the CIM Definition Standards for Mineral Resources and Mineral Reserves (27th

November 2010) by Mark Wanless, Pr.Sci.Nat (SACNASP - 400178/05), an appropriate

independent qualified person for the purpose of National Instrument 43-101.

Mineral resource classification is typically a subjective concept, industry best practices

suggest that resource classification should consider both the confidence in the geological

continuity of the mineralized structures, the quality and quantity of exploration data

supporting the estimates and the geostatistical confidence in the tonnage and grade

estimates. Appropriate classification criteria should aim at integrating both concepts to

delineate regular areas at similar resource classification.

SRK is satisfied that the geological modeling honors the current geological information and

knowledge. The location of the samples and the assay data are sufficiently reliable to

support resource evaluation. The sampling information was acquired primarily by core and

RC drilling on sections spaced at twenty to forty meters along strike. The range of continuity

of the grades is represented by the semi-variograms developed for each zone. For the zones

with well-structured semi-variograms and relatively long ranges such as zones 1, 2, 3, 4, and

9, the quality of the estimates in the densely drilled portions of the ore body is considered

sufficient to be classified in the Measured category within the meaning of the CIM Definition

Standards for Mineral Resources and Mineral Reserves (27th November 2010). Ordinary

Kriging allows a theoretical comparison to be made of actual block estimates and a

theoretically unbiased block estimate as a qualitative way of assessing the robustness of a

Kriged block estimate. This comparison is expressed in terms of slope of regression

between estimated blocks z*(v) and theoretical true blocks z(v), and SRK used this statistic

to select the areas to be classified in the Measured category. Areas where the block

estimates typically have a Slope of Regression of greater than 0.7 were considered as high

quality estimates, and classified in the Measured category.

Generally, for mineralization exhibiting good geological continuity investigated at an

adequate spacing with reliable sampling information accurately located, SRK considers that

blocks estimated during the first estimation run considering full variogram ranges can be

classified in the Indicated category. For the Measured and Indicated blocks, SRK considers

that the level of confidence is sufficient to allow appropriate application of technical and

economic parameters to support mine planning and to allow evaluation of the economic

viability of the deposit. Those blocks can be appropriately classified as Indicated.

Conversely, blocks estimated during the second and third pass considering search

neighborhoods set at twice or more than the variogram ranges should be appropriately

classified in the Inferred category because the confidence in the estimate is insufficient to

allow for the meaningful application of technical and economic parameters or to enable an

evaluation of economic viability.

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14.11 MINERAL RESOURCE STATEMENT CIM Definition Standards for Mineral Resources and Mineral Reserves (27th November

2010) defines a mineral resource as:

“A Mineral Resource is a concentration or occurrence of diamonds, natural solid inorganic

material, or natural solid fossilized organic material including base and precious metals, coal,

and industrial minerals in or on the Earth’s crust in such form and quantity and of such a

grade or quality that it has reasonable prospects for economic extraction. The location,

quantity, grade, geological characteristics and continuity of a Mineral Resource are known,

estimated or interpreted from specific geological evidence and knowledge.”

The “reasonable prospects for eventual economic extraction” requirement generally implies

that the quantity and grade estimates meet certain economic thresholds and that the mineral

resources are reported at an appropriate cut-off grade taking into account extraction

scenarios and processing recoveries.

In order to determine the quantities of material offering “reasonable prospects for economic

extraction” by an open pit, SRK used a pit optimizer and reasonable mining assumptions to

evaluate the proportions of the block model (Measured, Indicated and Inferred blocks) that

could be “reasonably expected” to be mined from an open pit.

The optimization parameters were selected based on the parameters defined for the reserve

pit optimization using contractor mining, aside from the gold price, which was increased from

the US$1,200/t used for the reserve open pit to an intentionally optimistic US$2,000/oz. The

reserve open pit optimization parameters are summarized in Table 14.20. The reader is

cautioned that the results from this pit optimization are used solely for the purpose of testing

the “reasonable prospects for economic extraction” by an open pit and do not represent an

attempt to estimate mineral reserves. The results are used as a guide to assist in the

preparation of a mineral resource statement and to select an appropriate resource reporting

cut-off grade.

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Table 14.20: Assumptions Considered for Conceptual Open Pit Optimization Parameter Value Unit

Gold Price 2,000 US$ per ounce

Mining Cost 3.3 US$ per tonne mined

Processing, General and

Administrative

Oxide: 17.22

Fresh: 16.89 US$ per tonne of feed

Mining Dilution 5 Percent

Mining Loss 3 Percent

Overall Pit Slope

Oxide: 30

Transitional: 45

Fresh: 50

Degrees

Process Rate Oxide: 1.60

Fresh: 1.30 Million tonne feed per year

Gold Process Recovery Oxide: 90.7

Fresh: 91.8 Percent

In Situ Cut-Off-Grade Oxide: 0.32

Fresh: 0.31 Grams per tonne

SRK considers that the blocks located above the maximum depth of the conceptual pit

envelope show “reasonable prospects for economic extraction” and can be reported as a

mineral resource. At Agbaou, the optimistic pit shell using the US$2,000/oz gold price

extends to the approximate depth of the deepest modeled portions of the ore body in Zones

1 and 2. As such, all the modeled mineralized material within the mineralized envelopes,

using a MIK cut-off of 0.3g/t has been reported as a mineral resource. The Mineral

Resources are presented in Table 14.21.

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Table 14.21: Mineral Resource Statement, Agbaou Project, SRK Consulting (Pty) Ltd., 30th March 2012 using MIK at a 0.3 g/t cut-off

Classification Zone Kt Grade (Au (g/t) Kozs

Measured 1 1,681 3.60 194

2 870 2.26 63

3 62 2.04 4

4 287 1.44 13

9 687 1.54 34

11 2,989 0.86 83

51 1,384 1.53 68

Total Measured 7,959 1.80 460

Indicated 1 2,318 2.71 202

2 2,369 2.20 168

3 504 1.51 25

4 882 1.24 35

5 617 6.18 123

6 326 3.77 40

7 969 2.28 71

9 325 1.14 12

10 267 1.59 14

11 503 1.23 20

51 313 1.25 13

52 236 0.99 8

53 146 0.94 4

Total Indicated 9,774 2.33 732

Total Measured and Indicated 17,733 2.09 1,192

Inferred 1 54 1.25 2

2 562 0.99 18

3 102 0.94 3

4 110 1.32 5

7 257 2.30 19

8 497 1.59 25

9 232 3.22 24

10 16 0.97 0.5

Total Inferred 1,830 1.64 97

For the purposes of comparison with the previous mineral resource estimate, the resources

have been reported at a range of grade cut-offs per resource classification in Table 14.22.

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Table 14.22: Mineral Resource Reported at Incremental Cut-off Grades, Agbaou Project, SRK Consulting (Pty) Ltd, 30th March 2012 using MIK

Lower Cut-off

Grade Mt

Average

Grade Kozs

(g/t Au) (g/t Au)

Measured

0.3 7.959 1.8 460.1

0.4 7.119 2.0 450.6

0.5 6.262 2.2 438.2

0.6 5.624 2.4 427.0

0.7 5.060 2.6 415.2

0.8 4.570 2.7 403.4

1.0 3.797 3.1 381.2

1.2 3.234 3.5 361.3

1.4 2.783 3.8 342.6

1.6 2.414 4.2 324.8

1.8 2.159 4.5 310.9

2.0 1.939 4.8 297.5

Indicated

0.3 9.774 2.3 732.3

0.4 9.270 2.4 726.7

0.5 8.708 2.6 718.5

0.6 8.025 2.7 706.3

0.7 7.489 2.9 695.1

0.8 6.981 3.0 682.9

1.0 6.114 3.3 658.0

1.2 5.456 3.6 634.7

1.4 4.843 3.9 609.1

1.6 4.266 4.2 581.3

1.8 3.812 4.5 556.5

2.0 3.428 4.8 533.1

Inferred

0.3 1.830 1.3 77.7

0.4 1.749 1.4 76.8

0.5 1.473 1.5 72.9

0.6 1.246 1.7 68.9

0.7 1.112 1.8 66.1

0.8 0.996 2.0 63.3

1.0 0.770 2.3 56.8

1.2 0.611 2.6 51.2

1.4 0.504 2.9 46.7

1.6 0.370 3.4 40.3

1.8 0.295 3.8 36.2

2.0 0.236 4.3 32.7

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SECTION 15. MINERAL RESERVE ESTIMATES

15.1 MINING APPROACH The following addresses the mining components of the feasibility study and is based on the

geological block model generated by SRK. This section of the report includes discussion on

the open pit optimisation, practical pit design, scheduling process, options investigated and

the reasons behind selections made. The mineral reserves, the results of the mine design

process in terms of production schedules as well as capital and operating cost estimates are

presented.

The proposed Agbaou process plant is based on a 1.6 million tonnes per annum (Mtpa) for

the oxide portion of orebody (Saprolite and Laterite) and followed by 1.34Mtpa in the

transition and fresh (Saprock and bedrock) zones. It has been assumed that a conventional

open pit shovel and truck method will be used over a Life of Mine (LOM) of approximately 8

years. Mining by contractor has been selected by ENDEAVOUR for the mining operations.

This has been the basis of the cost estimations. For comparison, an owner operated option

has also been prepared.

The design methodology involved two processes. In the first, Whittle-4X was used to identify

the optimum pit shell in terms of value and tonnage. In the second, using agreed mining

parameters, a practical pit has been designed to determine general mining requirements

such as dumping capacity, equipment requirements, operating costs and a mineable

tonnage profile that can be considered as a mineral reserve.

15.1.1 Resource Block Model A block model estimated by MIK (Multiple Indicator Kriging) was used as the basic resource

model for the pit optimization study. This resource model was imported to Surpac where the

majority of the mining design was prepared. Only mineralized material in the measured and

indicated categories was taken into account. Table 15.1 shows the Mineral Resource

classifications that have been used for the optimisation process.

Table 15.1: Mineral Resource Input for Optimisation (Using a 0.5g/t Au cut-off).

15.1.2 Geotechnical Investigation The geotechnical investigation and slope stability analysis was prepared by Golder

Associates (“Golder”), September 2008. Recommended slope designs by Golder, were

Mineral Resource category Tonnes

(Million)

Grade

(g/t Au) Ounces (Million)

Measured 6.267 2.2 0.438

Indicated 8.708 2.6 0.718

Total: Measured & indicated 14.975 2.4 1.156

Inferred 1.473 1.5 0.073 For

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based on the de-pressurized slopes by inter-ramp angles (SRK Consulting, SENET, Knight

Piésold Consulting, 2012 Appendix 15-A).

15.1.3 Gap Analysis for Pit Slope Design The Nevada offices of Golder Associates (Golder), completed a pit slope design in 2008, in

support of a feasibility study for Etruscan’s Agbaou gold project in Côte d’Ivoire, Africa.

Knight Piésold Ltd. (KPL) was requested to conduct a gap analysis on the September 2008

Golder Report titled “Pit Slope Recommendations – Agbaou Project, Ivory Coast” reference

number 08059. Observations and comments were strictly with regards to the report text

provided. The reviewer was not provided with additional information, nor was any site visit

made.

It should be noted that no check analyses were carried out as part of this review to verify pit

slope design angles presented in the Golder report. Rather, this is a high-level review to

check the compliance of the current feasibility level pit slope design for the Agbaou Project.

The pit slope design report is well written and comprehensive, with a reasonable discussion

of assumptions, expected failure modes, sensitivity analyses, water management issues,

and pit slope controlling factors. However, it is difficult to endorse the validity of the pit slope

design and some conclusions, as the level of data is not consistent with what would be

expected of a feasibility pit slope geotechnical assessment. In some areas, the analyses are

acceptable only for pre-feasibility level design and scoping but may also be lacking

information in several key areas. This is opposite to the report’s conclusion that “the factual

data collected to characterize the strengths of the saprolite, saprock, and bedrock is

sufficient for feasibility-level pit slope design”.

The complete report on the gap analysis review by Knight Piésold can be found in the

Engineering Optimization Study Report (SRK Consulting, SENET, Knight Piésold

Consulting, 2012 Appendix 15-B).

15.2 OPEN PIT OPTIMISATION The ore body model was exported from Surpac to Whittle/Gemcom Four-X where the open-

pit optimization studies were performed. The Whittle/Gemcom Four-X Analyser software

provides guidance to the potential economic final pit geometries. Whittle 4X compares the

estimated value of the individual mining blocks at the pit boundary versus the cost for waste

stripping. It establishes the pit walls where the ore revenue and waste stripping cost balance

for maximum net revenue.

The selected optimum pit shell is then engineered to generate practical pit designs that

incorporate the design slope angles and access ramps / haul roads for operating open pits.

The ore / waste tonnages in the practical pits are estimated and scheduled to determine the

ore production and the waste stripping requirements.

The Whittle process requires various input data including the resource block model, unit

costs and other physical parameters such as the slope angles at which the pit can be mined.

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Appropriate unit costs specific to the Agbaou operation were provided by ENDEAVOUR or

by other relevant parties. For the purpose of the Whittle optimization, capital costs,

depreciation, amortisation and other interest / finance charges have been excluded.

The following sections describe the methodology and derivation of the initial Whittle input

parameters and assumptions.

15.2.1 Slope Angles Based on the slope parameters derived by Golder, the following overall slope angles were

used for the pit optimization:

• Overall slope angle in Oxide material of 30°; • Overall slope angle in Transition material of 45°; • Overall slope angle in fresh material of 50°.

15.2.2 Mining Costs Two options for the mining cost have been prepared. The base case is for Mining

Contractors to carry out the operation with an Owner Operated option for comparison.

The base case contractor unit mining cost used is US$3.30/t. Various mining contractors

were approached for quotes based on schedules provided. The above unit cost provided by

one of the contractors was selected in conjunction with ENDEAVOUR. After optimizations

and initial scheduling were completed by SRK, detailed unit costs by rock type were received

from the contractor. The updated costs provided were tested in a subsequent optimisation

which showed minimal changes to the optimum pit shell (SRK Consulting, SENET, Knight

Piésold Consulting, 2012 Appendix 15-C). The owner operated unit mining cost used for

optimization purposes was US$2.25/t. This cost was originally provided by ENDEAVOUR.

15.2.3 Processing and General Administration Cost ENDEAVOUR provided the metallurgical processing recovery factors and costs for ore, and

the overall General and Administration (G&A) costs based on a throughput of 1.6Mtpa oxide

ore and 1.34Mtpa fresh ore thereafter. A breakdown of the costs and parameters used in the

Whittle optimization runs is shown in the table below.

Table 15.2: Processing Operating Cost for the Open Pit Optimization Type Oxide Fresh

Recovery 90.7% 91.8%

Processing 12.07 11.74

Dewatering 0.40 0.40

Rehandling 0.30 0.30

Grade control 0.20 0.20

Mine Supervision 1.25 1.25

Administration 3.00 3.00

Total 17.22 16.89

Royalty Oxide - US$ 5/oz Oxide - US$ 5/oz

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SENET provided updated processing parameters after optimisations and mining schedules

were completed. A comparative optimisation was prepared using these updated parameters.

Details of the parameters and results are discussed in the EOS (SRK Consulting, SENET,

Knight Piésold Consulting, 2012 Appendix 15-C).

15.2.4 Mining Factors Considering the nature of the geological contacts and the mining equipment selection, it was

decided that the mining dilution factor be set at 5% (at 0g/t grade) and mining recovery set at

97%. These factors were similar to those used in previous studies.

15.2.5 Gold Price and Royalties A gold price of US$1200/oz was selected by ENDEAVOUR which corresponds to other

mining studies being undertaken elsewhere at this time. An US$5.00/oz Royalty also was

included by ENDEAVOUR.

15.3 OPTIMISATION RESULTS

15.3.1 Cut-off Grade Calculation The metallurgical Cut-off Grade (“COG”) was derived calculating the overall cost of

producing one tonne of ore divided by the recovered value of the gold contained therein

[(process cost + Additional Ore Mining Cost)/(Process Recovery * (Price – Royalty))] . This is

calculated within the Whittle algorithm. This was determined to be 0.53 g/t for oxide and 0.51

g/t for fresh ore.

15.3.2 Optimisation Results – Mining Contractor Option The results of the open pit optimization using mining contractor option unit costs are shown

in Table 15.3 and Figure 15.1.

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Table 15.3: Whittle Optimisation Results (Contractor Option) Whittle

Pit Revenue RoM Inferred Rejected Waste Total Strip

Undiscounted

CashFlow

Au

OutPut

Number Factor Kt Au g/t kt Au g/t kt Au g/t kt kt Ratio M$ koz

1 0.3 1,118 2.75 1.0 1.11 8.0 0.51 1,339 2,466 1.20 79.1 89.7

2 0.4 1,961 2.66 3.0 1.00 13.0 0.51 3,947 5,924 2.01 126.8 152.2

3 0.5 5,108 2.72 23.0 5.37 71.0 0.46 24,278 29,480 4.75 291.1 406.2

4 0.6 6,174 2.65 47.0 3.62 87.0 0.45 30,751 37,059 4.98 329.3 477.1

5 0.7 9,357 2.63 152 1.82 115 0.46 60,817 70,441 6.50 437.9 719.6

6 0.8 10,144 2.60 187 1.68 122 0.46 69,046 79,499 6.81 453.7 773.2

7 0.9 10,819 2.57 227 1.58 125 0.46 76,096 87,267 7.03 460.8 814.1

8 1 11,384 2.55 248 1.56 126 0.46 83,931 95,689 7.37 463.3 851.9

9 1.1 11,965 2.54 272 1.56 143 0.46 93,361 105,741 7.80 460.4 890.7

10 1.2 12,582 2.49 291 1.55 145 0.46 100,436 113,454 7.98 455.3 920.3

11 1.3 12,811 2.47 299 1.53 161 0.46 102,639 115,910 8.01 452.7 929.4

12 1.4 13,032 2.46 302 1.53 184 0.45 105,789 119,307 8.12 448.3 939.8

13 1.5 13,330 2.44 316 1.52 191 0.45 110,950 124,787 8.32 440.0 955.0

14 1.6 13,595 2.43 323 1.54 197 0.45 115,687 129,802 8.51 431.5 967.8

15 1.7 13,932 2.40 337 1.52 237 0.45 119,870 134,376 8.60 422.0 979.6

16 1.8 14,085 2.39 337 1.51 241 0.45 122,540 137,203 8.70 416.1 986.0

17 1.9 14,216 2.38 342 1.50 255 0.45 124,814 139,627 8.78 410.7 991.2

18 2 14,335 2.37 344 1.50 258 0.45 126,684 141,621 8.84 406.0 995.4

The Inferred mineralised material shown lies within the particular pit shell and has not been

used as a revenue source in the optimisation.

Figure 15.1: Whittle Pit by Pit Graph Optimisation Results (Contractor Option)

Agbaou Gold Project Project No.

439430

0

20 000

40 000

60 000

80 000

100 000

120 000

140 000

0

50

100

150

200

250

300

350

400

450

500

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18

Kt

MU

S$

RoM Ore Waste UndisCashflow

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The amount of ore inside the optimum shell is approximately 11.4Mt at a grade of 2.55g/t.

The total material moved is 95.6Mt giving a stripping ratio of 7.4. This includes the Main and

West ore body.

15.3.3 Selection of Optimum Pit Shell As stated, the selection of the optimized final pit shell was based on maximum undiscounted

cash flow. In this instance, pit shell 08 for the Contractor option was selected to be used as a

template for the practical pit design, as shown in Figure 15.2.

Figure 15.2: Optimised Pit Shell Selected for Practical Pit Design

Agbaou Gold Project Pit Shells

Project No. 439430

Maximum undiscounted cash flow is used as an indicator for the optimum pit because

discounted cash flow (NPV) essentially penalises higher grade blocks which are scheduled

to be mined towards the end of the mine life. Depending on the discount rate used, the first

five years of the mining schedule have the greatest influence when NPV is used as an

indicator for the selection of the optimum pit shell. Generally optimum shells selected by

discounted cash flow are smaller than those selected by undiscounted cash flow.

15.3.4 Sensitivity Analysis A sensitivity analysis has been prepared by varying the unit mining cost, process and

administration and the gold price by +-10%. These sensitivities were carried using Whittle

FourX. Undiscounted cash flow was used to demonstrate the sensitivities. Figure 15.3

shows the results of this analysis.

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Figure 15.3: Sensitivity Analysis G

As can be seen the gold price is the

and administration cost. An additional

well as sensitivity of contained ounces to changes in key variables

SENET, Knight Piésold Consulting, 2012

15.4 PRACTICAL PIT DESIGN

15.4.1 Description The practical pit designs were

software was used to prepare the practical pit, and to incorporate the haul roads

together with the appropriate i

The open pit design criteria were:

• A nominal bench height of 10• The bench height was selected based on the optimum shovel size that can be

assumed for the nominal mining production rates;• Inter-ramp slope angles of

material. In the fresh material 49Inclusion of the ramping systems results in the approximate overbeing achieved;

• Haul road widths of 25traffic for the 90t capacity truck fleetis considered based on the manufactures recommendation as shown in Figure 15The recommendation indicates a min


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Sensitivity Analysis Graph

Agbaou Gold Project Mining Report

rice is the most sensitive followed by the mining

An additional analysis was prepared using NPV as

well as sensitivity of contained ounces to changes in key variables

SENET, Knight Piésold Consulting, 2012 Appendix 15-C).

DESIGN

The practical pit designs were prepared using the optimised pit shells as templates.

software was used to prepare the practical pit, and to incorporate the haul roads

together with the appropriate inter-ramp slope angles.

The open pit design criteria were:

A nominal bench height of 10m; The bench height was selected based on the optimum shovel size that can be assumed for the nominal mining production rates;

ramp slope angles of 32º in North pit and 39⁰ in South and West pitmaterial. In the fresh material 49º was used in the North pit and 45nclusion of the ramping systems results in the approximate over

25m including safety berms providing sufficient room for twot capacity truck fleet that was preferred by the contractors

is considered based on the manufactures recommendation as shown in Figure 15The recommendation indicates a minimum of 3.0 - 3.5 truck widths two

Page 15-7

Project No. 439430

ining cost then process

nalysis was prepared using NPV as an indicator, as

(SRK Consulting,

using the optimised pit shells as templates. Surpac

software was used to prepare the practical pit, and to incorporate the haul roads and ramps

The bench height was selected based on the optimum shovel size that can be

in South and West pit in the oxide was used in the North pit and 45⁰ in the South pit.

nclusion of the ramping systems results in the approximate overall slope angles

safety berms providing sufficient room for two-way that was preferred by the contractors. This width

is considered based on the manufactures recommendation as shown in Figure 15-4. 3.5 truck widths two-way traffic in

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straights and 3.5 - 4.0 truck widths in corners. Therefore the road width should be 20m on straights and 26m Minor ramps at the lower elevations otraffic density will be lower

Figure 15.4: Haul Road Widths

A haul road gradient of 10% ha

A total of five pits have been designed as follows:

• North, and South IntermediateWhittle pit number 04 as a template. constructed before mining commences. the initial scheduled stripping ratio. are shown in Figure 15

• North, South and WestWhittle pit number 08 are shown in Figure 15working areas to the primary crusher and the waste dumps. The view of these is shown in Figure 15-7.


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4.0 truck widths in corners. Therefore the road width should be and 26m on corners for this truck size (refer to the figure below)

Minor ramps at the lower elevations of the pits have been reduced in width where ower.

idths

Agbaou Gold Project Pit Design’s Road Widths

has been used throughout.

pits have been designed as follows:

Intermediate Pits – exploiting the upper sectionsWhittle pit number 04 as a template. Initial access to the upper elevations will be

ed before mining commences. These pits have been designed to minimise the initial scheduled stripping ratio. Two and three dimensional views of these cuts

Figure 15-5; , South and West Final Pits – exploiting the lower sections

as a template. Two and three dimensional views of these cuts Figure 15-6. Principal haul roads have been designed to connect the

working areas to the primary crusher and the waste dumps. The view of these is

Page 15-8

4.0 truck widths in corners. Therefore the road width should be (refer to the figure below).

f the pits have been reduced in width where

Project No. 439430

upper sections of ore body using Initial access to the upper elevations will be

been designed to minimise Two and three dimensional views of these cuts

lower sections of ore body using Two and three dimensional views of these cuts

Principal haul roads have been designed to connect the working areas to the primary crusher and the waste dumps. The view of these is

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Figure 15.5: View of the Agbaou


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View of the Agbaou Intermediate Pits

Agbaou Gold Project Interim Pit Design

Agbaou Gold Project Pit Interim 3D View

Page 15-9

Project No. 439430

Project No. 439430

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Figure 15.6: View of the Agbaou Final P


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View of the Agbaou Final Pits

Agbaou Gold Project Final Pit Design

Agbaou Gold Project Final Pit 3D View

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Figure 15.7: View of the Agbaou Final Pits Showing the Plant Area and the Waste Dump

Agbaou Gold Project Mine Surface Plan

Project No. 439430

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15.4.2 Mineral Reserves The practical pits have been used to generate the mineral reserves. The mineral reserve

statement uses the definitions and guidelines given in SAMVAL Definition on Mineral

Resources and Mineral Reserves and is reported in accordance with National Instrument 43-

101 requirements. Table 15.4 shows the mineral reserves estimated to be contained in the

Agbaou practical pit designs and the production schedule.

Table 15.4: Summary of Agbaou Mineral Reserves

Reserve Category Deposit Tonnes Grade Ounces

(Million) (g/t Au) (Million)

Proven Agbaou 5.407 2.25 0.390

Probable Agbaou 5.668 2.82 0.515

Total Proven and Probable Mineral Reserves 11.075 2.54 0.905

The three practical pits at Agbaou are mined simultaneously to provide 11 million tonnes of

ore to the processing plant (SRK Consulting, SENET, Knight Piésold Consulting, 2012

Appendix 15-D).

15.4.3 Comparison with Whittle Results The table below represents a comparison of practical pit design with the Whittle Shell.

Table 15.5: Comparison of Practical Pit with Whittle Shell

ROM

(kt)

Au

(g/t)

Inferred

(kt)

Waste

(kt) Total (kt)

Practical Pit 11,075 2.54 260 86,936 98,271

Whittle Shell 11,384 2.55 248 83,931 95,689

15.5 WASTE DUMP DESIGN The waste dumps have been designed using the following parameters:

• Face slope angle – 27o; • Bench height – 20m-30m; • Berm width – 15m-20m; • Overall slope – approximately 20o.

The waste dump capacities have been based on swell factors of 15% for saprolite and 20%

for fresh rock. No allowance has been made for back-filling of exhausted pits. The positions

of the three waste dumps are shown in

Figure 15.7. These positions have been provided by ENDEAVOUR and have been based on

prospectively sterile ground and take into account existing drainage patterns and proposed

infrastructure. Details of the three waste dumps are shown in

Figure 15.7.

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Table 15.6: Details of Waste

Waste Dump Average Height (m)

1

2

3

Figure 15.8: Vertical Sections Through Waste D

Figure 15-8 shows sections through the waste dumps showing the configuration. As waste

dump 1 is positioned on relatively steeply dipping terrain, a set sequence of dumping has

been planned. Initially dumping will start at the 340m elevation (upper) followed by the 300m

elevation (mid) then the 20m elevation (lower). This will enable the planned dump profile to

be achieved with minimal additional shaping. Waste dumps 2 and 3 are located on relatively

flat ground and will be constructed conventionally.


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aste Dumps

Average Height (m) Area Covered (ha) Capacity (Mm³)

70 65

60 35

60 41

Vertical Sections Through Waste Dumps

Agbaou Gold Project Waste Dumps

8 shows sections through the waste dumps showing the configuration. As waste


y dumping will start at the 340m elevation (upper) followed by the 300m



flat ground and will be constructed conventionally.

Page 15-13

Capacity (Mm³)

24

10

17

Project No. 439430

8 shows sections through the waste dumps showing the configuration. As waste


y dumping will start at the 340m elevation (upper) followed by the 300m



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Initially waste will be hauled and dumped on waste dump 3 which is closest to the North pit.

Table 15.7 shows the annual dumping volumes. This table shows the volumes of waste (in

m³) dumped, and the key displayed below the table indicates the corresponding waste

dump.

Table 15.7: Waste Dumping Volumes

Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Grand

Total (m3)

MNorth1 2 874 326 3 583 543 3 720 034 1 779 039 1 248 843 1 054 920 - - 14 260 705

MNorth2 - 918 360 1 204 940 1 196 304 3 691 765 4 063 315 4 381 620 2 876 733 18 333 037

MSouth1 - 2 554 143 2 026 266 790 064 - - - - 5 370 473

MSouth2 - - - 2 812 114 2 374 912 506 997 - - 5 694 023

West Pit 1 913 309 98 045 - - - - - - 2 011 354

Grand

Total (m3)

4 787 635 7 154 091 6 951 240 6 577 521 7 315 520 5 625 232 4 381 620 2 876 733 45 669 592

Key: Waste dump 1 Waste dump 2 Waste dump 3

15.6 MINE PRODUCTION SCHEDULE

15.6.1 Description The scheduling process consists of developing a mine plan and using the inventory included

in the practical pits. Scheduling also adjusted for plant feed criteria. The proposed Agbaou

process plant is based on a 1.6Mtpa in the oxide portion of the ore body (Saprolite and

Laterite) and 1.34Mtpa in the transition and fresh (Saprock and bedrock) zones. An amount

of three weeks feed ore (about 75% of one month capacity) as a ramp-up production plus

two months quantity of plant feed ore that have to be stockpiled near the plant are also

considered as pre-production schedule.

The mining schedule has been generated monthly for year 1, quarterly for year 2 and

annually for the remaining of the life of the operation. Four months pre-production before

gold pour is also considered in year zero to maintain a 4 week plant ramp-up as well as

stockpiling a quantity of 2 months plant feed. In this year mining will be started on outcrops

in the western parts of the North, South and West pits. Preparing access roads and

vegetation removal have to be done before the four months pre-production.

The schedule has been based on the sequential mining of ore and waste blocks in a

practical manner. Scheduling of waste material has been based on hauling to the dumping

area having the lowest cycle time. Growth of waste dumps is taken into account.

A maximum vertical advance rate of 60m per year has been applied to the scheduling. This

is considered by SRK to be the upper practical limit for open pits of this nature.

All mineralized material below the cut-off grade of 0.5g/t has been considered to be waste.

Only ore in the Measured and Indicated category is scheduled to the plant. Some Inferred

mineralised material occurs within the practical pits, which has been accounted for

separately.

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The upper portion of the ore body consists of relatively soft rock (Laterite and saprolite).

Therefore the majority of ore mined in the first 4 years will be soft rock. The amount of hard

rock portion (Saprock and bedrock) will then increase. Table 15-8 presents the proposed

Life-of-Mine (LOM) schedule (SRK Consulting, SENET, Knight Piésold Consulting, 2012

Appendix 15-D).

Table 15-8: LOM Production Schedule

YEAR/period Date Total

Ore (t) Au(g/t) Inferred (t) Au(g/t) Waste (t) Total (t)

Pre-Product (PP)

01/09/2013 93,193 1.21 159 20.0 433,196 526,547

01/10/2013 96,299 1.27 301 13.8 364,106 460,706

01/11/2013 93,193 1.55 543 20.5 348,965 442,700

PP Total

282,685 1.34 1,002

1,146,267 1,429,953

Commissioning 01/12/2013 96,299 1.47 1,393 11.4 192,681 290,373

Y1M1* 01/01/2014 138,923 1.81 264 1.0 515,961 655,148

Y1M2 01/02/2014 125,479 1.88 3,360 11.4 345,087 473,926

Y1M3 01/03/2014 138,923 1.69 684 1.4 553,658 693,266

Y1M4 01/04/2014 134,442 1.64 1,717 1.1 424,155 560,314

Y1M5 01/05/2014 138,923 1.35 0

572,530 711,453

Y1M6 01/06/2014 134,442 1.79 5,632 5.5 960,292 1,100,365

Y1M7 01/07/2014 138,923 1.80 2,154 1.5 1,106,285 1,247,362

Y1M8 01/08/2014 138,923 1.47 26 0.7 974,270 1,113,220

Y1M9 01/09/2014 134,442 2.11 3,105 1.2 1,161,853 1,299,400

Y1M10 01/10/2014 138,923 1.67 850 1.2 921,460 1,061,233

Y1M11 01/11/2014 134,442 1.44 0

753,285 887,727

Y1M12 01/12/2014 138,923 2.76 988 1.1 991,400 1,131,312

Y1 Total

1,635,711 1.78 18,780

9,280,236 10,934,727

Y2Q1 01/01/2015 403,359 2.28 793 1.3 2,601,365 3,005,517

Y2Q2 01/04/2015 407,807 2.14 2,238 1.6 2,896,802 3,306,848

Y2Q3 01/07/2015 412,289 2.41 2,620 4.4 3,453,844 3,868,753

Y2Q4 01/10/2015 412,289 2.39 5,958 3.0 3,308,731 3,726,979

Y2 Total

1,635,744 2.31 11,609

12,260,743 13,908,097

Y3 01/01/2016 1,644,002 2.89 23,214 0.9 12,664,611 14,331,827

Y4 01/01/2017 1,454,229 3.02 12,003 1.4 11,803,451 13,269,684

Y5 01/01/2018 1,326,569 2.14 95,326 1.0 17,951,845 19,373,740

Y6 01/01/2019 1,348,689 2.62 36,175 1.1 11,761,638 13,146,502

Y7 01/01/2020 1,040,651 2.81 32,268 0.9 7,578,532 8,651,451

Y8 01/01/2021 610,347 3.27 41,028 1.1 3,240,502 3,891,877

GRAND TOTAL**

11,074,927 2.50 272,799 1.5 87,880,505 99,228,231

*Y1M1 refers to Year 1, Month 1. This nomenclature follows throughout the table above.

**Grand Total is the added amount of: Pre-Product; Commissioning, and Y1 – Y8.

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An investigation was carried out using a raised cut-off in order to process higher grade ore

initially and stockpile the lower grade ore for processing towards the end of the life of the

operation. The limiting factor to this was that the vertical advance rate was exceeded in the

initial years of operation. This was not considered in the scheduling process for this study.

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SECTION 16. MINING METHODS

16.1 INTRODUCTION SRK understands that contractor-based mining is preferred by ENDEAVOUR. The selected

mining contractor will thus be responsible for the site preparation (including removal of

vegetation) haul road construction, excavation and haulage of ore to the Run of Mine (ROM)

pad and waste to the waste dumps, oversize breakage and equipment maintenance.

The mining operations are based on the use of a hydraulic excavator and haul truck fleet

engaged in conventional open pit mining techniques. The excavators load the

blasted/broken material into the haul trucks, with the ore being transported to the processing

area and the waste to the designated local waste dumps. Access roads will be developed as

required for access into new areas. The main arterial roads where necessary will be

constructed to a minimum 20m width, including berms and drainage areas.

Topsoil in mining areas will be recovered during the pit preparation phase and stockpiled for

future use with progressive waste dumps and possible pit rehabilitation.

The majority of the gold mineralised material will be transported to the ROM area, and

discharged directly into the feed hopper. An ROM stockpile will be available for surplus

material at the ROM pad area.

All of the waste material from the excavation area will be hauled to the external waste dump,

adjacent to the operational pits. The total waste to be moved is about 87Mt, which

approximates to 50 million cubic meters of dumping volume.

The in-situ materials in hard and semi-hard rock will require drilling and blasting to assist

fragmentation and subsequent loading. Based upon the information supplied, and

experience in similar operations, it has been assumed that that oxide portion of the ore body

is free digging or may be in need of light blasting in certain areas. It has also been assumed

that in the owner-operated option, on-site staff at Agbaou will also undertake mine planning,

mine scheduling, grade control and performance monitoring.

Four local and internationally experienced mining contractors were requested to submit an

estimate of the cost of mining based upon the annual mine production schedule. One of

these submissions was suited to the operation. This submission was used as a basis to

generate operating and capital costs.

16.1.1 Mining Equipment Requirements The mining equipment shown in Table 16.1has been selected based upon the annual mine

production schedule. This has been prepared by the mining contractor.

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Table 16.1: Equipment Selection for Mine Operation by Contractor Equipment Type & Model Quantity

Drill Rig ATLAS COPCO ROCL8 1

Breaker MONTABERT BRV65 1

Excavator

CAT 390c 1

CAT 336DL 1

LIEBHERR R9250 1

LIEBHERR R984 4

Dozer CAT D8R 4

Grader CAT 14G 1

Wheel Dozer CAT 824H 1

Truck CAT 777G 8 5 8-12

CAT DUMP BOX 1 1

Water Tank CAT 773+MTT13 1

Wheel Loader CAT 988H 1 1

CAT 966H 1 1

Compactor BOMAG BW219 1 1

Water Tank CLAAS 1

Mobile Crane GROOVE RT890 1 1

Tyrehandler CAT 966TH 1 1

Service truck RVI SERVICE 9 9

Telehandler CAT TH414 1 2

Lowbed 1

16.1.2 Mine Work Schedule Table 16.2 summarises the working schedule, the scheduled shifts per year and the

expected shifts available, accounting for holidays (based on data received by

ENDEAVOUR). The mine is initially scheduled to work 329 days per year, less time for

holidays and inclement weather. It will operate two shifts per day for ore and three shifts per

day for waste, 8 hours per shift for 658 available shifts each year for ore and 987 shifts for

waste for the mine life. A 3-crew roster system will be required to maintain this schedule.

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Table 16.2: Scheduled Working Periods

Parameter Units Value

Calendar Days Days 365

Days per Week Days 7

Holidays Days 15

Weather Days 21

Scheduled Days Days 329

Shifts/Day-Ore Shifts 2

Shifts/Day-Waste Shifts 3

Annual Work Shifts-Ore Shifts 658

Annual Work Shifts-Waste Shifts 987

Hours per Shift-Ore Hours 8

Hours per Shift-Waste Hours 8

Scheduled Hours_Ore Hours 5,264

Scheduled Hours_Waste Hours 7,896

The operating time per shift will be the actual time during the shift that the equipment is

“productively” working at its capacity. This is equal to total scheduled time less all scheduled

and unscheduled delays.

16.1.3 Open Pit Dewatering The mine planning/schedule information obtained from ENDEAVOUR indicates that the

open pits will reach a maximum depth of 150m to 175m deep at the end of the life of mine.

The maximum inflow to the proposed open pits is expected to be approximately 27L/s.

However, it should be noted that significantly increased inflows could be encountered during

mining owing to the presence of previously unmapped structural features which may be

water bearing.

As a result of the groundwater conditions present on site, open pit dewatering is

recommended through the use of sub-horizontal gravity drainage boreholes (drains) drilled

into the walls of the pit in areas where seepage is observed. These drains are recommended

to be 50m in length and drilled at an upward angle of ±5° perpendicular to the pit wall in

every stage of the pit at 20m – 30m apart. The first 15m - 20m of each hole should be cased

with plain PVC casing to prevent collapse. The drains can be drilled by the blasting rig thus

decreasing the cost of drilling and could be even more effective for the dewatering of

vertical/subvertical faults. Water drained from these boreholes will be collected in sumps at

the base of the pit and be pumped out using a centrifugal pump.

In the event that dewatering through the use of drains is not effective, or inflow rates are

higher than expected, it is recommended that a number of dewatering boreholes are drilled

along structural features which have been shown to contribute to inflow. This can reduce the

groundwater inflow to the pits by intersecting the flows before they reach the walls.

Groundwater abstracted from the open pits should be discharged to the raw water supply

dam to avoid release to the environment. As such, the abstracted groundwater will augment

to process water supply requirement to some degree, although it is not sufficient to meet all

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of the process water requirements.

Dewatering of the open pits will affect the groundwater quantity by lowering the water table

elevation and this influence has to be monitored so that there is minimal or no impact on

nearby boreholes and surface water (such as dams and base flow in streams). Groundwater

quality should also be monitored to control any impact of mining activities such as leakage

from the TSF, WRD, the discharged water, and the storage and handling of chemicals, such

as reagents and fuels.

If a significant groundwater quantity and/or quality impact is detected outside of the site

boundary and this impact results in significant de-watering of the water supply boreholes of

adjacent landowners, provision should be made to provide alternative water supplies to

affected parties. This is likely to consist mainly of the drilling of deeper boreholes in order to

provide alternative water supplies.

A comprehensive network of monitoring boreholes should be established at strategic

locations around the site to monitor pit dewatering impact on groundwater quantity and

quality.

16.1.4 Mine Infrastructure The mining infrastructure is covered in Section 18.2.

A layout of surface access roads has been shown in Figure 16.1. These roads allow access

between the pits, plant, explosive magazine, ROM Stockpile and waste dumps. The total

distance amounts to approximately 5km. These roads will be constructed by the mining

contractor under ENDEAVOUR’s supervision.

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Figure 16.1: Layout of Surface Access Roads, Explosive Magazine and ROM Stockpile Location

Agbaou Gold Project Surface Access Road

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The bulk explosives facility will be located away from the mining facilities, at a location to be

decided during the construction phase. Figure 16.1 shows a location suggested by SRK. A

safety aspect of the site will need to be investigated.

A detonator magazine either of concrete construction or with double brick thickness and

surrounded by earthen screening mounds and security fencing, and a small industrial

explosive magazine will be required as separate units situated within close proximity of the

main magazine complex, but all within an enclosed security fencing and managed

accordingly.

An allowance has been made in the mining costs for an adequate supply of hand held and

vehicle mounted radio sets. SRK recommends the fitting of in-vehicle radios to all of the

hydraulic excavators, drill rigs, graders and wheel loaders, to the haul trucks and to the

service vehicles and supervisors transport.

16.1.5 Mining Manpower The ENDEAVOUR mining personal includes mining manager and technical services (survey,

geology/grade control and mine planning) as per below:

• Manager – 2 persons; • Mining supervisor – 5 persons; • Geology and planning engineers – 3 persons; • Surveyor – 2 persons; • Artisans – 5 persons.

The remainder of the mining operational labour will be provided by the mining contractor.

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SECTION 17. RECOVERY METHODS

17.1 PROCESS DESIGN CRITERIA

17.1.1 Introduction Metallurgical testwork results and industry norms, where necessary, were used to define the

process design criteria for the Agbaou Gold Project. The process plant is designed to treat

either 1.6 million tonnes per annum of saprolite ore or 1.34 million tonnes of bedrock ore, or

a combination of the two ore types if required.

The proposed process design criteria consist of crushing, ore stockpiling, milling and

classification, gravity and ILR, CIL, cyanide detoxification, tailings disposal, acid wash,

elution, electrowinning and gold room, carbon regeneration, consumables, oxygen and air

services; and water services.

Several references have been used to derive data used in the process design criteria and

are:

• ENDEAVOUR; • Metallurgical testwork; • Calculated data; • Vendor data or recommendations; • SENET; • Industry standards or practices; • Handbook (engineering handbook); • Assumptions based on experience; • External consultants.

17.1.2 Ore Characteristics The comminution testwork results (obtained from the previous study), specific gravity

determinations (obtained from the current study’s testwork) and information obtained from

ENDEAVOUR were used to characterise the ore as shown in the table below.

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Table 17.1: Design Criteria – Ore Characteristics

* The saprolite samples were too fine to have been subjected to comminution testing.

17.1.3 Operating Schedule The primary crushing circuit will be operated on a 16-hour basis for bedrock and 18-hour

basis for saprolite, with overall utilisation of 83% and 85%, respectively. The remaining hours

in both cases will be set aside for maintenance. The table below summarises the operating

schedule.

Table 17.2: Design Criteria – Operating Schedule

ORE CHARACTERISTICS Units Saprolite Bedrock Source

Ore Source Open Pit Open Pit Client

Max Lump Size mm 800 800 Client

Ore Head Grades (Au) g/t 2.10 2.40 Client

Moisture Content % 18% 5% Client

Specific Gravity 2.79 2.82 Testwork

Bulk Density of Crushed Ore t/m³ 1.40 1.69 Calculated

Rod Mill Work Index: Mpa Not tested * 15.5 Testwork

Ball Mil l Work Index: kWh/t Not tested * 11.8 Testwork

Abrasion Index: kWh/t Not tested * 0.185 Testwork

JK Tech Parameters

A Not tested * 62.3 Testwork

b Not tested * 0.54 Testwork

A x b Not tested * 33.8 Calculated

Ta Not tested * 0.48 Testwork

OPERATING SCHEDULE Units Saprolite Bedrock Source

General

Annual Tonnage Treated Mtpa 1 610 000 1 340 280 Client

Ore Processing Tonnes per Month t/month 134 167 111 690 Calculated

Primary Crushing

Operating Hours per Day hrs 18 16 Client

Scheduled Maintenance per Week hrs 12 12 Client

Scheduled Maintenance per Annum hrs 720 720 Calculated

Operating Hours per Annum hrs 5 558 4 864 Calculated

Util isation % 95% 95% SENET

Overall Utilization % 85% 83% SENET

Selected Crushing Throughput t/hr 290 290 SENET/Client

Milling

Operating Hours per Day hrs 24 24 Client

Scheduled Maintenance per Week hrs 7 7 Industry Practice

Scheduled Maintenance per Annum hrs 364 364 Calculated

Number of Mill Relines per Anuum # 2 2 Industry Practice

Time Required per Mil l Reline hrs 24 24 Industry Practice

Time Lost per Annum due to Rel ines hrs 96 96 Calculated

Operating Hours per Annum hrs 7 885 7 885 Calculated

Util isation % 95% 95% SENET

Overall Utilization % 90% 90% SENET

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17.1.4 Plant Recovery The overall plant recoveries are shown in the table below. In the metallurgy section the

gravity recoveries obtained from testwork are 30% for saprolite and 20% for bedrock. Taking

these recoveries into account the CIL feed grade was calculated. The CIL laboratory

dissolutions were calculated using this feed grade and the CIL tails grade. A confidence level

of 98.5% was applied to the laboratory dissolutions to obtain the CIL plant dissolutions of

91.9% for saprolite and 90.4% for bedrock. Ultimately the total overall recoveries were

obtained as 93.2% for saprolite and 91.4% for bedrock.

Table 17.3: Design Criteria – Recovery

17.1.4.1 Crushing and Ore Stockpiling

There will be one crushing circuit for both saprolite and bedrock ores. The circuit will produce

a SAG mill feed of 100% passing 240mm that will be stockpiled prior to milling. The

maximum lump size of the feed to the crushing plant will be 800mm.

The bedrock comminution testwork results, from the previous feasibility study, were

reviewed by OMC and the outcomes of that review were used as a basis for the design of

the crushing circuit. The crushing design criteria are shown in the table below.

RECOVERY Units Saprolite Bedrock Source

Gold (Au) Recovery

Head Grade g/t head 2.10 2.40 Client

Gravity

Gravity Recovery % head grade 30.0% 20.0% Testwork

Gravity Recovery g/t 0.61 0.46 Calculated

CIL

CIL Feed Grade g/t head feed 1.49 1.94 Calculated

CIL Laboratory Dissolution % 93.3% 91.7% Calculated

Confidence Level % 98.5% 98.5% Industry Practice

CIL Plant Dissolution% of CIL Feed

Grade91.9% 90.4% Calculated

CIL Dissolution g/t head feed 1.37 1.75 Calculated

CIL Solution Tai ls ppm 0.015 0.015 Industry Practice

CIL Solids % m/m % 40.0% 43.0% Testwork

CIL Solution Tai ls g/t head feed 0.023 0.020 Calculated

CIL Solid Tails g/t head feed 0.100 0.160 Testwork

Combined CIL Tails g/t head feed 0.123 0.180 Calculated

CIL Overal l Recovery % head grade 65.3% 72.9% Testwork

Total Overal l Recovery % head grade 93.2% 91.4% Calculated

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Table 17.4: Design Criteria – Crushing and Ore Stockpiling

* The dry crusher feed rate shown in the table above is that of the vibrating grizzly feeder oversize material.

17.1.4.2 Milling

The milling circuit will comprise a single stage SAG mill and a ball mill, as well as a pebble

crusher for the SAG mill oversize. The circuit will produce the CIL feed of 80% passing

75µm. The SAG mill discharge slurry will have a solids density of 55% w/w solids for

saprolite and 70% w/w solids for bedrock. The oversize from the SAG mill, with a top size of

60mm will be crushed further in a pebble crusher with a closed side setting of 15mm.

The bedrock ore characteristics and comminution testwork results were used as a basis for

the design of the milling and classification circuit as summarized in the two tables below.

Comminution characteristics for saprolite were estimated from the OMC database as the

material was too fine to support the required testing and no further testwork has been

conducted at the time of producing this report.

The comminution testwork results were sent to OMC, who modelled several mill

configurations to determine the most suitable and cost effective installation. OMC’s

interpretation for mill configuration and sizing of both mills can be found in the following

report – December 2011, OMC, “Agbaou Gold Project Comminution Testwork and Circuit

Design Review”, Report No. 8857.30 Rev 0 these are included in Appendices 13-A and 13-

B.

The milling and classification design criteria are shown in the table below.

CRUSHING & ORE STOCKPILING Units Saprolite Bedrock Source

ROM Bin

Maximum Lump Size mm 800 800 Client

Selected Grizzly Spacing mm x mm 800 800 Industry Practice

Method of Feeding ROM Bin Truck & FEL Truck & FEL Client

Truck Type CAT 777 CAT 777

Rom Bin Capacity m³ 100 100 Calculated

Withdrawal Method from Bin APRF & GRIZ APRF & GRIZ Client

Apron Feeder Nominal Capacity t/hr 290 290 SENET/Client

Primary Crusher

Type of Crusher Jaw Jaw Industry Practice

Dry Feed Tonnage t/hr 181 * 181 * Calculated

Vibrating Grizzly Undersize as % of ROM Feed % 38% 38% Handbook

Crusher Closed Side Setting mm 150 150 OMC

Primary Crusher P80 mm 148 148 OMC

Primary Crusher P100 mm 240 240 OMC

Stockpile

Design Live Capacity days 1.0 1.0 Industry Practice

Selected Stockpile Live Capacity tonnes 5 000 5 000 Calculated

Reclaim Method Apron Feeders Apron Feeders Client

Number of Reclaim Feeders # 2.0 2.0 Industry Practice

Reclaim Feeder Withdrawal Rate t/hr 205 170 SENET/Client

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Table 17.5: Design Criteria – Milling

* The OMC report has a SAG mill discharge density of 50% solids. However, when the milling circuit water balance was done for saprolite

ore at this density a surplus of water was encountered. Thus, a density of 55% was selected. This will be confirmed during the post-study

testwork campaign.

**Ball loads 12% and 15% for the SAG as per OMC report.

MILLLING Units Saprolite Bedrock Source

Milling Circuit Selection

Circuit type SABC SABC SENET/Client

SAG Milling

Number of Mills # 1.0 1.0 OMC

Wet or Dry Mill ing Wet Wet OMC

Open or Closed Circuit Open Open OMC

Overflow or Grate Grate Grate OMC

Feed Size (F80) mm 148 148 OMC

Product Size (P80) µm 857 857 OMC

Discharge Slurry % Solids % 55% * 75% SENET

Selected SAG Mill Size:

Diameter (Inside Shell) m 6.71 6.71 OMC

Effective Grinding Length (EGL) m 3.54 3.54 OMC

Mill Speed as % of Critical Speed % 75% 75% OMC

Drive Type Variable Speed Variable Speed OMC

Operating SAG Mill Bal l Load % TBC 12% ** OMC

Maximum SAG Mill Bal l Load % TBC 15% ** OMC

Ball Material (High Cr, Cast or Forged) High Cr High Cr SENET/Client

Mill Liner Material (Steel, Rubber, Polymet) Steel/Composite Steel/Composite Client

Ball Solids Density t/m³ 7.82 7.82 Supplier

Selected Bal l Size mm TBC 125 OMC

SAG Mill Discharge Screen

Screen Type Vibrating Vibrating Client

Screen Sol ids Flowrate t/hr 236 196 Calculated

Type of Screen Panels PU Slotted PU Slotted Industry Practice

Screen Aperture mm x mm 12 x 55 12 x 55 Industry Practice

Screen Draining Rate m³/m²/h 250 250 Supplier

Selected Screen Area m² 4.65 4.65 Supplier

Ball Milling

Number of Mills # 1.0 1.0 OMC

Wet or Dry Mill ing Wet Wet OMC

Open or Closed Circuit Closed Closed OMC

Overflow or Grate Overflow Overflow OMC

Feed Size (F80) mm 12 12 SENET

Product Size (P80) µm 75 75 Testwork/Client

Discharge Slurry % Solids % 55% 75% SENET

Selected Ball Mill Size:

Diameter (Inside Shell) m 4.50 4.50 OMC

Effective Grinding Length (EGL) m 6.75 6.75 OMC

Mill Speed as % of Critical Speed % 75% 75% OMC

Drive Type Fixed Speed Fixed Speed OMC

Operating Ball Mil l Ball Load % TBC 30% OMC

Maximum Ball Mil l Ball Load % TBC 35% OMC

Ball Material (High Cr, Cast or Forged) High Cr High Cr Industry Practice

Mill Liner Material (Steel, Rubber, Polymet) Composite Composite Client

Selected Bal l Size mm TBC 65 OMC

Ball Mill Trommel Screen

Screen Type Trommel Trommel SENT/Client

Screen Sol ids Flowrate t/hr 200 365 Calculated

Type of Screen Panels PU Slotted PU Slotted Industry Practice

Screen Aperture mm x mm 12 x 55 12 x 55 Industry Practice

Screen Draining Rate m³/m²/h 170 340 Handbook

Required Screen Area m² 1.37 1.26 Calculated

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17.1.4.3 Classification and Trash Handling

A cyclone cluster will classify the mill product to produce an overflow of 40% w/w solids

(saprolite) and 43% w/w solids (bedrock). Testwork results showed that viscosity issues will

not be encountered at a cyclone overflow density of 40% for saprolite. The overflow will be

screened and directed to the CIL circuit for recovery of gold.

The cyclone overflow will gravitate onto a vibrating trash screen where any trash material

present in the ore will be removed as screen overflow. The screen underflow will be

channeled into a leach feed splitter box.

Table 17.6: Design Criteria – Classification and Trash Handling

17.1.4.4 Gravity

Batch laboratory GRG tests that were conducted indicated both ore types contain free gold.

The testwork results were used to predict the plant GRG recoveries. The design criteria for

the gravity circuit are shown in the table below.

Table 17.7: Design Criteria – Gravity

CLASSIFICATION AND TRASH HANDLING Units Saprolite Bedrock Source

Mill Discharge Sump

Cyclone Feed Slurry Flowrate m³/hr 623 618 Calculated

Residence Time mins 2.0 2.0 Industry Practice

Selected Sump Volume m³ 20 20 Calculated

Classification

Type of Classification Cluster Cluster Industry Practice

Number of Clusters # 1.0 1.0 SENET

Circulating Load % 150% 250% Industry Practice

Cyclone Feed Slurry Flowrate m³/hr 623 618 Calculated

% Solids in Cyclones Overflow % 40% 43% Testwork/SENET

Cyclone Overflow Slurry Flowrate m³/hr 381 286 Calculated

% Solids in Cyclones Underflow % 70% 70% Industry Practice

Cyclone Underflow Slurry Flowrate m³/hr 242 333 Calculated

Overflow Product Size (P₈₀) µm 75 75 Industry Practice

Trash Screening

Type of Screen Vibrating Vibrating Client

Aperture Size µm 0.63 x 8.8 0.63 x 8.8 Industry Practice

GRAVITY Units Saprolite Bedrock Source

Scalping Screen

Type of Screen Vibrating Vibrating Industry Practice

Aperture Size mm 2.0 2.0 Industry Practice

Screen Dry Solids Feed Rate t/hr 120 66.6 Calculated

% Solids in Screen Feed % 70% 70% Industry Practice

Feed Slurry Flowrate m³/hr 94.4 52.1 Calculated

Dilution Flowrate m³/hr 39.9 33.0 Calculated

Gravity Concentrator

Type of Gravity Unit Centrifugal Centrifugal Industry Practice

Number of Units # 1.0 1.0 SENET

% Solids in Concentrator Feed % 55% 50% Calculated

Gravity Feed Grade g/t 2.10 2.40 Client

Gravity Recovery (% Au of Head grade) % 30% 20% Testwork/SENET

Concentrate Mass as % of Feed to Conc % 0.03% 0.03% Industry Practice

Gravity Concentrate Treatment Type Intense Leaching Intense Leaching SENET/Client

Concentrate Batches Treated per Day # 1.0 1.0 SENET

Gravity Electrowinning

Number of Gravity Electrowinning Cells # 1.0 1.0 SENET/Client

Number of Electrowinning Cycles per Month # 30 30 Industry Practice

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17.1.4.5 Carbon-in-Leach (CIL)

Pre-leaching ahead of the CIL was excluded because the testwork results indicated that

there is a significant presence of preg-robbers in both saprolite and bedrock ores.

A conditioning or oxidation stage has been included in the design for the introduction of

oxygen in the ore prior to leaching with sodium cyanide. This step helps increase the leach

kinetics. The pH adjustment is also fine tuned in the oxidation tank prior to leaching.

The cyanidation tests that were conducted resulted in a leach residence time of 24hrs for

saprolite and 32 hours for bedrock. As discussed in Section 13.1.5 these leach residence

times were selected as optimal. The design criteria for the CIL circuit are shown in the table

below.

Table 17.8: Design Criteria – CIL

17.1.4.6 Cyanide Detoxification

Batch laboratory tests conducted indicated that cyanide destruction could be achieved by

using the Inco process (i.e. using sodium metabisulphite, and copper sulphate as a catalyst).

A residence time of 2 hours (including redundancy) in this circuit was initially assumed prior

to conducting testwork. This resulted in a residence time of 1 hour per detox tank. The

testwork results revealed an optimum detox residence time of 45 minutes. Post study

options, each having the same effect, for optimizing the circuit are:

CARBON IN LEACH (CIL) Units Saprolite Bedrock Source

Leach

Solids Flow to CIL t/hr 205 170 Calculated

CIL Feed Grade - Design Only % 2.10 2.40 Client

% Solids in Leach Feed % 38.8% 41.3% Calculated

Feed Slurry Flowrate m³/hr 397 302 SENET

CIL Tank Residence Time hrs 24 32 Testwork

Air/O₂ Hold-up % 8.0% 8.0% Industry Practice

Number of Oxidation Tanks # 1.0 1.0 SENET/Client

Number of CIL Stages/Tanks # 7.0 7.0 SENET/Client

Total Number of Tanks # 8.0 8.0 Calculated

Leach Dissolution (Au) % 91.9% 90.4% Testwork/SENET

CIL Solution Tai ls Grade g/t 0.023 0.020 Calculated

CIL Solids Tai l Grade g/t 0.100 0.160 Testwork

Design Carbon Concentration in Slurry g/L 10.0 10.0 SENET

Carbon to Gold loading Ratio 1 000 1 000 Industry Practice

Design Incremental Carbon Loading g/t 1 253 1 564 Calculated

Carbon Batch Size t 6.0 6.0 SENET

Eluted Carbon Value g/t 100 100 Industry Practice

Loaded Carbon Value g/t 1 353 1 664 Calculated

Interstage Screens

Interstage Screen Type MPS (P) MPS (P) Industry Practice

Aperture Size µm 800 800 Industry Practice

Loaded Carbon Screen


Aperture Size mm 0.63 x 8.8 0.63 x 8.8 SENET

Carbon Safety Screen m³/hr Calculated


Aperture Size µm 0.8 x 8.8 0.8 x 8.8 SENET

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• Reduce tanks sizing by half;

• Remove one of the tanks

Lime slurry will be added into the circuit to neutralise sulphuric acid that is generated during

the detoxification process, thus maintaining the pH within the desired range of 8 – 10.

The residual WAD cyanide concentration in the post leach tails was assumed to be 100

ppm. The detoxification process has been designed to decrease the WAD cyanide

concentration in the final tailings to 10 ppm. These assumptions will be reviewed in the next

project phase. The design criteria for the cyanide detoxification circuit are shown in the two

tables below.

Table 17.9: Design Criteria – Cyanide Detoxification

17.1.4.7 Acid Wash, Elution and Carbon Regeneration

An acid wash column will be used and a batch tank will also be provided to facilitate the measurement of a loaded carbon batch to acid wash and elution. The Anglo American Research Laboratories (AARL) method has been selected as the elution method due to the high number of elution cycles per month predicted for a 6 t carbon batch elution size. A horizontal regeneration kiln will be used. Carbon transfer from acid wash column to elution column and elution column to regeneration

CYANIDE DETOXIFICATION Units Saprolite Bedrock Source

Detox method

Sodium

Metabisulphite &

Copper Sulphate

Sodium

Metabisulphite &

Copper Sulphate

Client

Required Residence Time hrs 2.0 2.0 Assumption

Residual Cyanide in Tail ings (before Detox) ppm 100 100 Assumption

Cyanide in Final Detoxified Tails ppm 10 10 Industry Practice

Solids Flow from CIL tph 205 170 Calculated

Solids Density in Detox Feed % 37.6% 39.8% Calculated

Feed Slurry Flowrate m3/hr 413 318 Calculated

Calculated Detox Volume m3 826 635 Calculated

Number of Tanks # 2.0 2.0 Industry Practice

Calculated Volume per Tank m3 413 318 Calculated

Air Hold-up % 10% 10% Industry Practice

Calculated Tank Volume (incl. Air Hold-up) m3 459 353 Calculated

Selected Volume per Tank m3 500 500 SENET

Detox Sump

Required Residence Time mins 4.0 4.0 Industry Practice

Calculated Sump Volume m3 27.5 21.2 Calculated

Selected Sump Volume m3 30 30 SENET

Detox Air Requirements

Volume of Air Required per TankNm3/h/1000m3

tank vol1000 1000 Industry Practice

Number of Tanks to be Aerated # 2.0 2.0 Industry Practice

Slurry Volume per Tank m3 500 500 Calculated

Total Volume of Air Required Nm3/h 1000 1000 Calculated

Excess Return Water Detox

Method H₂O₂/HCl H₂O₂/HCl Testwork

Catalyst Used CuSO₄ CuSO₄ Testwork

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facility will be via hydraulic means.

Due to the price of power in Cote d’Iviore being relatively low, it was considered

economically viable to use electric heaters for the elution and regeneration heating systems.

Table 17.10: Design Criteria – Acid Wash, Elution and Regeneration

17.1.4.8 Electrowinning and Smelting

Gold will be recovered from the pregnant eluate via conventional electrowinning in a single

20-cathode electrowinning cell. Stainless steel cathodes will be utilized. Gold sludge washed

from the cell will be dewatered using a pot filter and dried in a drying oven prior to smelting.

Fluxes will form a slag which removes various metal oxides from the gold sludge. A mixture

of molten gold and slag will be poured into cascade moulds where gold will be retained in the

initial moulds due to its higher density.

Gold bullion bars will be cleaned, assayed, labelled and readied for shipping. The design

criteria for electrowinning and smelting are shown in the table below.

ACID WASH, ELUTION & REGENERATION Units Saprolite Bedrock Source

Type of Acid Wash Vessel Column Column SENET

Material of Construction FRP FRP Industry Practice

Flow Through the Vessel BV/hr 2 2 Industry Practice

Acid Wash Tank

Minimum Tank Volume BV 1.0 1.0 Industry Practice

Minimum Tank Volume m³ 13.3 13.3 Industry Practice

Acid Wash Solution Strength % 3.0% 3.0% Industry Practice

Rinsing

Rinse Volume BV 2.0 2.0 Industry Practice

Rinse Time hrs 1.5 1.5 Industry Practice

Acid Wash Vessel Emptying Method Hydraulic Hydraulic Calculated

Elution

Elution Method AARL AARL Industry Practice

Operating Temperature 0C 130 130 Industry Practice

Operating Pressure kPa 350 350 Industry Practice

Carbon Transfer Method Hydraulic Hydraulic Industry Practice

Number of Elutions per Month - Design # 40 30 Industry Practice

Total Elution Cycle Time hrs 10 10 Calculated

Carbon Batch Size t 6.0 6.0 SENET

Design Barren Carbon Loading g/t 100 100 Industry Practice

Flow Through the Column 2.0 2.0 Industry Practice

Eluant Tank

Minimum Eluant Tank Volume BV 1.0 1.0 Industry Practice

Cyanide Strength in Eluant % 2.0% 2.0% Industry Practice

Caustic Strength in Eluant % 3.0% 3.0% Industry Practice

Elution Heating

Elution Heating Type Thermic Oil Thermic Oil Industry Practice

Elution Heaters Type Electric Electric Client

Kiln Feed Hopper

Dewatering Means Strainers Strainers Calculated

Hopper Capacity m³ 2 Strip Batches 2 Strip Batches Calculated

Regeneration Kiln

Type of Kiln Horizontal Horizontal Vendor

Type of Kiln Heating Electric Electric Client

Temperature Control Automatic Automatic Vendor

Regeneration Temperature deg C 750 750 Industry Practice

Regeneration Time per Day hrs 20 20 Industry Practice

Quench Tank Type Pan Pan Industry Practice

Type of Regenerated Carbon Storage Tank Eductor Eductor Industry Practice

Carbon Transfer Method Hydraulic Hydraulic Industry Practice

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Table 17.11: Design Criteria – Electrowinning and Smelting

17.1.4.9 Consumables

Consumptions derived from testwork were used as a basis for sizing make-up and dosing

facilities for reagents and consumables. The design criteria for the reagents are summarised

in the tables below. Major reagents and consumables considered are jaw crusher liners, mill

liners, grinding media, lime, sodium cyanide, caustic, sodium metabisulphite, copper

sulphate, hydrochloric acid, activated carbon, smelting fluxes, hydrogen peroxide and

flocculant. Various lubricant and minor consumables have been costed by factoring or as a

nominal sum only.

Lime The bulk of the lime required in the CIL will be provided by the screw conveyor and the ring

main will serve a fine tune control function for lime dosage. The ring main will be capable of

delivering the full CIL dosage. The required detox lime dosage will be supplied from the ring

main.

Table 17.12: Design Criteria – Lime

ELECTROWINNING & SMELTING Units Saprolite Bedrock Source

Electrowinning

Flash Tank Residence Time min 0.50 0.50 Industry Practice

Cell Type Atm Sludging Atm Sludging Industry Practice

Number of CIL Cells # 1.0 1.0 Client

Cell Solution Temperature deg C 85.0 85.0 Industry Practice

Type of Cathode S/Steel Mesh S/Steel Mesh Industry Practice

Type of Anode S/Steel S/Steel Industry Practice

Barren Solution Grade ppm 5.0 5.0 Industry Practice

Percentage Gold Recovery 98% 98% Industry Practice

Cathode Sludge RemovalHigh Pressure

Wash

High Pressure

WashIndustry Practice

Sludge Handling

Fi lter Type Pot Fi lter Pot Filter Client

Fi lter Cake Moisture % 20% 20% Assumption

Fi lter Capacity litres 60.0 60.0 Assumption

Drying & Smelting

Drying Oven Type Electric Electric SENET

Number of Drying Ovens # 1.0 1.0 SENET

Number of Trays # 6.0 6.0 Industry Practice

Smelting Furnace Type Induction Induction SENET

Bull ion Mould Capacity kg 25 25 Supplier

Type of Crucible TPX 400 TPX 400 Industry Practice

LIME Units Saprolite Bedrock Source

Lime Silo Sizing

Delivery Method Truck Truck Client

Type of Lime Hydrated Lime Hydrated Lime Client

Bulk Storage Silo Silo Industry Practice

Bulk Storage Capacity days 4.0 4.0 Client

Lime Make-up Tank

CIL Consumption as 100% kg/t head 3.23 1.50 Testwork

Detox Consumption as 100% kg/t head 0.30 0.13 Testwork

Combined Consumption as 100% kg/t CaO 3.53 1.63 Testwork

Lime Make-up System Lime Mixing Lime Mixing Industry Practice

Number of Make-ups per Day # 4.0 4.0 SENET

Make-up Tank Capacity hrs 8.0 8.0 SENET

Dosing Tank Capacity days 1.0 1.0 SENET

Lime Dosing Method Ring Main Ring Main Industry Practice

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Sodium Cyanide

Table 17.13: Design Criteria – Sodium Cyanide

Caustic Table 17.14 Design Criteria – Caustic

Sodium Metabisulphite

Table 17.15: Design Criteria – Sodium Metabisulphite

Copper Sulphate

Table 17.16: Design Criteria – Copper Sulphate

SODIUM CYANIDE (NaCN) Units Saprolite Bedrock Source

Delivery Method Bag in Box Bag in Box Supplier

Delivery Size kg/box 1000 1000 Supplier

Physical Form Briquettes Briquettes Supplier

Cyanide Consumption - ILR kg/t 0.01 0.01 Calculated

Cyanide Consumption - CIL kg/t 0.37 0.20 Calculated

Cyanide Consumption - Elution kg/t 0.02 0.02 Calculated

Total Cyanide Consumption kg/t 0.41 0.24 Testwork

Make-up & Dosing Tank Capacity days 3.0 3.0 Industry Practice

Number of Make-up & Dosing Tanks # 1.0 1.0 SENET

Cyanide Make-up Strength % m/m 25% 25% Industry Practice

Cyanide Dosing Method Ring Main Ring Main Industry Practice

CAUSTIC SODA (NaOH) Units Saprolite Bedrock Source

Delivery Method Bags Bags Supplier

Delivery Size kg 25.0 25.0 Supplier

Physical Form Pearls Pearls Supplier

Total Caustic Consumption kg/t 0.13 486 Calculated

Caustic Solution Storage Capacity days 3.0 3.0 SENET

Number of Caustic Tanks # 1.0 1.0 SENET

Caustic Solution Strength % m/m 20% 20% Industry Practice

CAUSTIC SODA (NaOH) Units Saprolite Bedrock Source


Delivery Size kg 25.0 25.0 Supplier

Physical Form Pearls Pearls Supplier

Total Caustic Consumption kg/t 0.13 486 Calculated

Caustic Solution Storage Capacity days 3.0 3.0 SENET

Number of Caustic Tanks # 1.0 1.0 SENET

Caustic Solution Strength % m/m 20% 20% Industry Practice

COPPER SULPHATE (CuSO4) Units Saprolite Bedrock Source


Delivery Size kg 25 25 Supplier

Physical Form Crystals Crystals Supplier

Usage Rate per Residual Cyanide g/g WAD 0.465 0.277 Testwork

Number of Make-up &Dosing taxTanks # 1.0 1.0 SENET

Make-up Strength % 15% 15% Industry Practice

Copper Sulphate Solution Storage Capacity days 4.0 4.0 SENET

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Hydrochloric Acid

Table 17.17: Design Criteria – Hydrochloric Acid

Activated Carbon

Table 17.18: Design Criteria – Activated Carbon

17.1.5 Air Services and Oxygen

Compressed Air and Instrument Air Two duty/standby compressors have been sized to provide compressed air requirements for

the following:

• Instrument air; • Plant air.

Oxygen Two 2.5 tpd oxygen plants will be installed to provide oxygen required for the leaching of

gold in the CIL Both oxygen plants will be running at the same time. Oxygen will be

produced at a purity of at least 96% v/v.

HYDROCHLORIC ACID (HCl) Units Saprolite Bedrock Source

Delivery Method IBC IBC Supplier

Delivery Size litres 290 290 Supplier

Physical Form Solution Solution Supplier

HCl Delivered Strength % 33% 33% Supplier

Average Acid Consumption kg/t 0.102 0.098 Calculated

Average Acid Consumption per Carbon Batch kg/batch C 341 364 Calculated

ACTIVATED CARBON Units Saprolite Bedrock Source


Delivery Size kg/bag 500 500 Supplier

Type of Carbon in Use Mesh 8x16 8x16 Supplier

Type of Carbon in Use mm 1.68 x 2.38 1.68 x 2.38 Supplier

Carbon Bulk Density t/m³ 0.45 0.45 Handbook

Carbon Dry SG 0.80 0.80 Industry Practice

Carbon Wet SG 1.37 1.37 Industry Practice

Carbon Consumption Rate g/t 25 25 SENET

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17.1.5.1 Water Services

Table 17.19: Design Criteria – Water Services

* The raw water demand during the dry season is based on the assumption that 40% of water to the tailings dam will be

returned to the plant.

WATER SERVICES Units Saprolite Bedrock Source

Plant Raw Water Supply (incl. Fire Water)

Type of Storage Tank Tank Client/SENET

Average Raw Water Demand - Commissioning m3/hr 65 - Calculated

Average Process Water Top-up - Commissioning m3/hr 278 - Calculated

Average Raw Water Demand - Dry Season m3/hr 65 * 64 * Calculated

Average Process Water Top-up - Dry Season m3/hr 129 * 118 * Calculated

Average Raw Water Demand - Wet Season m3/hr 65 63 Calculated

Average Process Water Top-up - Wet Season m3/hr 0 0 Calculated

Residence Time hr 3.2 3.2 Assumed

Selected Raw Water Tank Volume m3 1 500 1 500 SENET

Process Water

Type of Storage Pond Pond Client/SENET

Average Return Water Flowrate - Commissioning m³/hr 0 - Calculated

Average Return Water Flowrate - Dry Season m³/hr 149 116 Calculated

Average Return Water Flowrate - Wet Season m³/hr 278 234 Calculated

Total Instantaneous Process Water Flowrate m³/hr 315 271 Calculated

Residence Time hr 21 21 Client/SENET

Selected Process Water Tank Volume m³ 6 500 6 500 Calculated

Plant Potable Water

Average Potable Water Usage m3/hr 3.0 3.0 SENET

Storage Capacity hrs 48 48 SENET

Selected Tank Volume m³ 150 150 SENET

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17.2 PROCESS PLANT

17.2.1 Process Plant Description17.2.1.1 Introduction

The Agbaou gold process plant

dry saprolite ore or 1.34 million tonnes per annum of dry bedrock ore. It will comprise

crushing, ore stockpiling, grinding, classification, gravity gold

tails screening, cyanide detox, tailings handling and disposal, acid wash, elution,

electrowinning, carbon regeneration, gold room, consumables, oxygen, air services and

water services. Gold will be produced as bullion bar

to the overall process flow diagram shown below.

Figure 17.1: Process Flow D

The process plant general layout drawing is show


Technical Report

PLANT

Plant Description

The Agbaou gold process plant will be capable of treating 1.6 million tonnes per annum of

1.34 million tonnes per annum of dry bedrock ore. It will comprise

crushing, ore stockpiling, grinding, classification, gravity gold recovery, carbon in leach (CIL),



water services. Gold will be produced as bullion bars ready for shipment to a refinery.

to the overall process flow diagram shown below.

Process Flow Diagram

The process plant general layout drawing is shown below.

Page 17-14

million tonnes per annum of

1.34 million tonnes per annum of dry bedrock ore. It will comprise

recovery, carbon in leach (CIL),



s ready for shipment to a refinery. Refer

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Figure 17.2: Plant General Layout

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17.2.1.2 Crushing and Ore Stockpiling

The crushing plant will treat both saprolite and bedrock ore at a rate of 290 tonnes per hour

to a product size of 100% passing 240mm through a primary jaw crusher. When treating the

bedrock ore the crushing circuit will operate for a period of 16 hours per day and 18 hours

per day when treating the saprolite ore. The balance of the hours in a day will be used for

maintenance.

ROM ore will be trucked and tipped directly into a 100m³ ROM bin equipped with an 800mm

x 800mm static grizzly to scalp off any oversize material. Ore will be reclaimed from the bin

by an apron feeder that will discharge onto a vibrating grizzly feeder with 150mm spacings.

The vibrating grizzly feeder oversize will be routed to the primary jaw crusher that will reduce

the ore size to 80% passing 148mm. The crusher product and the vibrating grizzly undersize

will be conveyed to a 5,000 tonne live capacity stockpile. The total stockpile capacity is

approximately 60,000 tonnes. Crushed ore will be reclaimed from the stockpile using two

apron feeders and transferred to the SAG mill. Each stockpile reclaim apron feeder will be

sized to reclaim 205tph.

Pebbles (with a top size of 60mm) recycled from the SAG mill discharge screen will report to

the pebble crusher which will have a closed side setting of 15mm. The pebble crusher

discharge will join the SAG mill feed stream as mill feed material.

17.2.1.3 Milling and Classification

The milling circuit will comprise a SAG mill, ball mill and pebble crusher that will be capable

of treating 205 tph of saprolite ore and 170tph of bedrock ore. The SAG mill will operate in

open circuit while the ball mill will run in closed circuit with a cyclone cluster.

Crushed ore reclaimed from the stockpile will be conveyed to the SAG mill. Dilution water

will be added to the SAG mill to achieve the required 55% w/w solids density for saprolite

and 75% for bedrock in the SAG mill discharge for optimum milling efficiency. The mill feed

dilution water will be added and controlled as a ratio of the SAG mill solids feed rate. The

ratio control constant can be adjusted from the relevant SCADA screen to account for

varying mill feed rate and ore moistures.

The 6.71m diameter x 3.54m long (inside liners) SAG mill will be fitted with a 2900kW motor

and will have a grate discharge. The mill is designed to operate at 75% critical speed but can

be varied to optimise operation for the different ore types. The ball load will be 12% by

volume, while the total load will be 25%.

Lime will be added to the SAG mill feed conveyor for pH adjustment in the downstream CIL

circuit. The SAG mill will discharge onto a single deck vibrating screen that will remove

pebbles as oversize (with a maximum lump size of 60mm). The screen will be equipped with

polyurethane panels with 12x55mm apertures. The dewatered pebbles will be recycled to

the pebble crusher, before returning to the SAG mill.

The SAG mill discharge screen undersize will gravitate into the mill discharge sump, where it

will combine with the ball mill discharge slurry. The combined slurry will be diluted with

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process water to a solids density of 54% for saprolite and 59% for bedrock, and pumped to a

cyclone cluster for classification. The cyclone cluster will classify the slurry to produce an

overflow of 80% passing 75µm at 40% for saprolite and 43% for bedrock that will gravitate

onto a vibrating trash screen for removal of any trash material.

A portion of the cyclone underflow, at a solids density of 70% for both saprolite and bedrock,

will be routed to the gravity concentration and intensive cyanidation circuit for recovery of

free gold. The remainder of the cyclone underflow will return to the ball mill, where further

grinding will take place. The 4.50m diameter x 6.75m long (EGL) ball mill with 2100kW of

installed power will operate at a ball charge of 30% and total charge of 35% by volume. It will

operate in closed circuit with the cyclone cluster. The ball mill will be equipped with a

trommel screen (fitted with 12x55mm polyurethane panels) to remove any grinding media

that may escape from the ball mill. The trommel screen undersize will report to the mill

discharge sump where it be mixed with the SAG mill discharge screen undersize as

mentioned above.

17.2.1.4 Gravity Gold Recovery

The feed to the gravity circuit will gravitate onto a gravity scalping screen to remove the

+2mm material. Dilution water will be added to the scalping screen feed box to decrease the

concentrator feed solids density to 55% for saprolite ore and 50% for bedrock ore. The

screen oversize will return to the ball mill feed while the screen undersize will be fed to a

gravity concentrator to recover the free gold. Concentrator tails will gravitate back to the mill

discharge sump.

The leaching of gravity concentrates will be a batch-wise process using an In-line Leach

Reactor (ILR150BA). The leach solution will be prepared by first adding caustic to water for

pH adjustment and then sodium cyanide to a 2% concentration. The leaching of gold will be

effected by circulating the 2% cyanide solution through the rotating reactor drum. Oxygen

will be added to the circulating electrolyte during the leach.

At the end of the leach cycle, which will range between 14 – 16 hrs, the drum will be stopped

and the solution in will gravitate to the ILR sump. The solution in the sump will be pumped

into the solution storage tank where it will be clarified by adding flocculant. The clarified

solution will be pumped to a gravity electrowinning tank. Wash water will be added to the

drum to wash entrained solution from the solids and allowed to clarify in the solution tank

before being pumped to the electrowinning tank.

The pregnant solution stored in gravity electrowinning tank will be pumped to a dedicated

electrowinning cell for the recovery of gold. Loaded cathodes will be periodically removed

from the gravity electrowinning cell to a wash tank, where the gold sludge will be washed off

the cathodes using high-pressure spray water. Gold sludge will be decanted filtered and

transferred to a calcining furnace.

17.2.1.5 Carbon-in-Leach (CIL)

The screened cyclone underflow will be sampled (for metallurgical accounting and process

control) prior to reporting to a leach feed splitter box in the CIL section. The slurry will

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overflow from the splitter box into the oxidation tank, where oxygen will be added through the

agitator shafts and dispersed at the bottom of the tanks for enhanced mass transfer. Oxygen

will be added throughout the CIL train. Lime slurry from a ring main will also be added to the

oxidation tank via the splitter box to fine tune the pH adjustment prior to the addition of

sodium cyanide solution in the CIL train.

The CIL circuit will consist of seven tanks, each with a live volume of 1500m³, resulting in a

total retention time of 24 hours for saprolite ore and 32 hours for bedrock ore. The oxidation

tank and seven CIL tanks will all be equipped with dual impeller agitators to keep the slurry

and carbon particles in suspension during the leach process.

Sodium cyanide solution from a ring main, will be automatically dosed into the first CIL tank.

Manual addition of cyanide solution into the first three CIL tanks will also be provided and

can be used in the event where the determined cyanide concentrations are low.

Each CIL tank will be fitted with one interstage wedge wire cylindrical screen which will

prevent the migration of activated carbon during slurry transfer. The pumping action of the

internal impeller mechanism of the interstage screens will drive the flow of slurry from one

CIL tank to another. In the case of an interstage screen becoming silted, it will be lifted with a

tower crane from the tank and placed onto a wash frame. A spare screen will be used to

replace the silted screens. High pressure low volume wash sprays will be used to clean

blocked screen holes.

Regenerated, eluted or virgin carbon will be added to the last CIL tank. Recessed impeller

vertical spindle pumps will be installed in the CIL tanks to periodically transfer carbon

upstream from CIL tank 7 all the way to CIL tank 1. Loaded carbon in CIL tank 1 will be

pumped onto a vibrating loaded carbon screen. The slurry will return to CIL tank 1. High

pressure water will be sprayed on the screen to ensure that clean carbon reports as screen

oversize to the acid wash batch tank. The tails slurry leaving the last CIL tank will be

pumped to a tails screen prior to cyanide detoxification.

17.2.1.6 Tails Screening and Cyanide Detox

The CIL tails slurry will be screened to recover any carbon that escapes from the CIL. The

screen oversize (carbon) will gravitate to a basket that will periodically be lifted using the CIL

tower crane and emptied into the last CIL tank. The screen undersize will gravitate to the

detox feed splitter box. The splitter box will feed the slurry to the first of two detox tanks

operating in series. If required, one tank can be online with the other bypassed. Each detox

tank will have a live volume of 500m³, thus resulting in a total residence time of 2 hours in

the detox circuit.

Sodium metabisulphite, copper sulphate and blower air will be added to the circuit to provide

sulphur dioxide, copper catalyst and oxygen, respectively. The free cyanide and/or weakly

bound metal cyanide complexes present in the tailings slurry will be oxidized to the less toxic

cyanate (OCN⁻) according to the reaction:

CN⁻WAD + SO₂ + O₂ + H₂O = OCN⁻ + H₂SO₄

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Lime slurry will be added into the circuit to neutralise the generated sulphuric acid, and thus

maintain the pH within a range of 8 – 10.

The detoxified slurry exiting the second detox tank will overflow into a final tails sump. An in-

line two stage sampler will take the tails slurry samples in the launder at regular intervals.

The final tails pumps will pump the detoxified tails slurry into the tailings dam.

17.2.1.7 Tailings Handling and Disposal

As mentioned above, the final tails pumps will pump the detoxified tails slurry into the tailings

dam. The return water pumps will pump supernatant water from the tailings dam to the

process water pond in the plant. This water will serve as process water to be used in the

plant.

A facility will be installed at the return water pond to destroy cyanide in the excess return

water prior to discharge into the environment. This will consist of a copper sulphate make-up

and dosing system, as well as storage and dosing systems for hydrochloric acid and

hydrogen peroxide.

Hydrogen peroxide will be used for the destruction of residual cyanide. It will be supplied at a

concentration of 60% in intermediate bulk containers (IBC’s), from where it will be pumped to

a 5m³ dosing tank. Copper sulphate will be used to as a catalyst. It will be supplied in 25kg

bags and made up to a solution of 15% concentration by weight. Hydrochloric acid will be

used for pH adjustment. It will be supplied at a concentration of 33% in 210L drums and

transferred to a 5m³ dosing tank. All these reagents will be dosed into the excess return

water pond prior to discharge of water into the environment.

17.2.1.8 Acid Wash

Loaded carbon will be discharged from the loaded carbon screen into an acid wash batch

tank. The batch tank will be sized to accommodate 12 tonnes of loaded carbon (i.e. 2

batches of 6 tonnes each). Loaded carbon will gravitate to the acid wash column.

During the acid wash cycle the dilute hydrochloric acid solution (3% HCl) will be circulated

from the acid wash tank through the acid wash column at a rate of 2 bed volumes per hour

for a period of 1 hour. The solution exiting the column will return to the acid wash tank via

the external strainers that will prevent any carryover of carbon from the column to the tank.

At the end of one hour of acid washing, the acid wash pump will be stopped. The contents of

the acid wash column will be rinsed with a volume of raw water equivalent to 2 bed volumes

in order to remove the residual acid from the loaded carbon. The chlorides may result in

reduced strip efficiencies and stainless steel corrosion, especially that of the elution column.

The rinse effluent exiting the acid wash column will be directed to the cyanide detox circuit

for disposal into the tailings dam. Rinsed carbon will be hydraulically transferred to the

elution column.

After about 4 acid wash cycles the dilute acid in the acid wash tank will be too contaminated

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for further use. Caustic solution will be added to the tank to neutralise the acid. The

neutralized solution will be pumped to the cyanide detox circuit for disposal into the tailings

dam.

A fresh dilute hydrochloric acid solution will be made up by adding approximately 1186 litres

of 33% hydrochloric acid to raw water make up dilute hydrochloric acid solution of 3%

strength.

17.2.1.9 Elution

The elution of gold from loaded carbon will be done using the Anglo American Research

Laboratories (AARL) method. The eluant (3% caustic, 2% cyanide solution) will be pumped

at a flowrate of 2 bed volumes per hour from the eluant tank via the recovery and primary

heat exchangers into the elution column. In the column gold adsorbed onto the carbon will

be removed by washing with a number of bed volumes of hot water. This will result in the

stripping of gold from the carbon into the eluate solution.

Heat for the elution circuit will be provided by two electrical elution heaters. The heater

elements will heat up thermic oil, which will be pumped into the primary shell and tube heat

exchanger. Heat will be transferred from the thermic oil to the eluant solution entering the

primary heat exchanger. The hot eluate leaving the elution column will be cooled down in the

recovery heat exchanger by contacting with fresh eluant solution that is going into the elution

column. Both the primary and recovery heat exchangers will be shell and tube type.

The eluate will flow out of the elution column as pregnant electrolyte and into the pregnant

electrolyte tank in the electrowinning circuit.

Each elution cycle will run for a period of 12 hours, and the heating cycle will be in operation

for approximately 10 hours of that 12-hour cycle. At the end of each elution cycle, eluted

carbon will be hydraulically transferred to the eluted carbon holding tank in the regeneration

circuit. Alternatively, if regeneration is not required eluted carbon will be transferred into the

last CIL tank.

17.2.1.10 Electrowinning

The eluate from the elution column will be directed to fill the pregnant tank. When the tank is

sufficiently full of eluate and the electrowinning circuit is available, the pregnant solution will

be pumped to the CIL electrowinning cell. Excess solution in the cell feed tank will overflow,

bypassing the cells, and join the cell outlet solution that returns to the selected pregnant

tank. The nominal electrolyte flow rate to the cell is 30m3/h.

Sludging type stainless steel mesh cathodes will be utilized to electrowin gold from the

eluate solution. Gold will be deposited as a loosely adhering fine sludge onto the pad of

stainless steel knit mesh cathodes contained in the baskets. The electrowinning cycle will

run for a period of approximately 10 hours. Once an electrowinning batch is complete i.e. the

gold grade of the solution has reduced to a stipulated level (5 ppm and below), the spent

electrolyte will be directed to the barren electrolyte tank, from where it will be pumped to the

eluant tank or the leach feed splitter box in the CIL circuit.

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The loaded cathodes will periodically be harvested and placed inside a wash tank, where

they will be washed with high pressure low volume sprays of raw water to remove the gold

sludge from the cathodes into the wash tank. The clean cathodes will be returned to the

cells. The resultant gold sludge will be collected in a sludge tank, and dewatered using a

batch type pot filter. Compressed air will be utilized in the pot filter to assist in dewatering the

sludge. The filter cake will be packed into trays and placed in the calcining oven for drying.

A fume extraction fan will be installed to collect potentially poisonous and explosive gases

that evolve from the cells during electrowinning. The fan will discharge through the gold

room roof into the atmosphere.

17.2.1.11 Carbon Regeneration

At the end of the elution cycle eluted carbon will be hydraulically transferred to a carbon

holding tank in the regeneration circuit. The tank, sized to handle two barren carbon batches,

(12 tonnes) will be fitted with strainers at the bottom to help drain off excess water. A

provision will be made for carbon to bypass the regeneration facility and be transferred

directly to the last CIL tank.

Eluted carbon will be withdrawn from the holding tank at a rate of 600kg/hr by a screw feeder

and discharged to an electric rotary kiln for thermal regeneration. This rate will allow for two

regeneration cycles per day. Excess water which might not have drained through the carbon

holding tank strainers will be drained through the wedge wire screen mounted on the screw

feeder. On entering the kiln, the wet carbon, which will still contain about 50% moisture; will

be dried and heated to the required regeneration temperature of 750°C. Regenerated

carbon exiting the kiln will be quenched and screened to remove carbon fines. The screen

oversize will gravitate into a carbon transfer vessel, from where it will be hydraulically

transferred to the last CIL tank.

17.2.1.12 Gold Room

The filter cake from the pot filter will be poured onto stainless steel trays and loaded in an

electric drying oven that will operate at 150 - 200°C. At the end of calcining the trays will be

removed from the oven and allowed to cool down.

When the dried sludge has cooled down, fluxes will be added in determined proportions. The

mixture of the sludge and fluxes will be poured into a smelting crucible which will, in turn, be

placed in an electric smelting furnace operated between 1200⁰C and 1400⁰C. The furnace

will be placed under a fume hood with an exhaust duct connected to an extraction fan.

During smelting, metal oxides will report to the slag. At the end of smelting the furnace

crucible contents will be poured into cascading moulds mounted on a trolley. The bullion will

collect in the bullion bar moulds while slag will flow and collect in slag moulds. The heavy

metallic phase will sink to the bottom of the moulds whilst the light slag phase will float on

top of the metallic phase. When both phases cool down and solidify the glassy slag phase

will be easily broken away from the metallic phase, leaving a relatively pure gold bar.

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When the gold bars have cooled down, they will be cleaned by hammering to remove any

traces of adhering slag. Then they will be sampled, labeled, weighed and stored in a safe

prior to dispatch to the refinery.

17.2.1.13 Consumables

Mill Balls Grinding media of sizes 125mm and 65mm will be used in the SAG and ball mills,

respectively. For ease of transportation, grinding media will be delivered in drums (200L) and

stored in separate concrete bunkers. Grinding media balls will be charged to the mills using

a bottom discharge kibble.

Hydrochloric Acid Hydrochloric acid will be delivered in 210L plastic drums at a concentration of 33% w/w.

During make-up the acid will be pumped from the drums into the acid wash tank. The weak

acid will be pumped to the acid wash column during acid wash cycle as described above.

Lime Hydrated lime will be delivered to the plant in trucks. It will be pneumatically offloaded into a

storage silo. The bottom cone of the silo will have a dual mechanical transfer arrangement

comprising two screw feeders. The first screw feeder will supply lime to a make-up tank,

where the slurry of 20% w/w will be made up. The second feeder will transfer lime to the

SAG mill feed conveyor. Both screw feeders will be sized to be able to supply the full

demand of hydrated lime in the plant.

The lime silo will also be equipped with a vibrator in order to enhance the discharge of lime

from the silo. The silo will also be fitted with a filter to render the top of the silo dust free

during periods when lime is being transferred.

After make-up, lime slurry will be transferred to a dosing tank. Lime dosing pumps will

distribute lime slurry to the CIL for fine tune pH control and cyanide detox circuits by means

of a ring main. Unused lime slurry will be returned to the dosing tank through the ring main

return line.

Caustic Soda (Sodium Hydroxide) Caustic will be delivered to site in 25kg bags packed into wooden pallets. During caustic

make-up, a hoist will lift the caustic pallet to a platform on top of the caustic tank. An

operator will manually lift one 25kg bag at a time from the platform to the bag breaker.

Caustic pearls will be discharged into the caustic tank to be mixed with raw water to form a

20% w/w caustic solution. The tank mixer will ensure complete dissolution of caustic.

The caustic dosing pumps will only be run for the time required to deliver the various

quantities of the caustic solution to various distribution points, including the cyanide make-

up, acid wash, elution, pregnant tanks and intense leach reactor.

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Sodium Cyanide Sodium cyanide will be delivered to site in 1 tonne bulk bags, packed into wooden crates to

minimise any possible cyanide spillages during transportation. The dissolution of sodium

cyanide briquettes will take place in a covered cyanide make-up tank. This tank will have two

compartments, with the mixer installed in the larger compartment. The larger compartment

will be used for the make-up and dosing of cyanide solution. The smaller compartment will

be used for the dosing of cyanide solution during periods when a batch of fresh cyanide

solution is being made up.

Prior to the addition of cyanide briquettes in the cyanide tank, the larger compartment will be

filled with water to a predetermined level. A reagents hoist will be used to lift the cyanide

bags onto a bag breaker, from where they will be discharged into the larger compartment of

the cyanide make-up tank to form a 25% cyanide solution. The tank mixer will ensure

complete dissolution of cyanide briquettes during the make-up process. The cyanide solution

will be distributed in a ring main to the intensive cyanidation circuit, CIL circuit and elution

circuit.

Sodium Metabisulphite Sodium metabisulphite powder will be delivered to site in 1 tonne bulk bags. The dissolution

of sodium metabisulphite will take place in a partitioned tank with two. The larger

compartment will be used for the make-up and dosing of sodium metabisulphite solution.

The smaller compartment will be used for the dosing of sodium metabisulphite solution

during periods when a batch of fresh solution is being made up.

When the bags have been delivered to the make-up area, they will be lifted onto a bag

breaker by means of a hoist. They will be discharged into the larger compartment of the

make-up tank and dissolved in water to form a 25% sodium metabisulphite solution. The

tank mixer, installed in the larger compartment, will ensure complete dissolution of the

sodium metabisulphite powder during the make-up process. The variable speed peristaltic

pumps will be used to transfer sodium metabisulphite solution to the cyanide detox circuit at

a controlled rate.

Copper Sulphate Copper sulphate crystals will be delivered to site in 1 tonne bulk bags. The dissolution of

crystals will take place in a partitioned tank with two compartments. The larger compartment

will be used for the make-up and dosing of copper sulphate solution. The smaller

compartment will be used for the dosing of copper sulphate solution during periods when a

batch of fresh solution is being made up.

When the bags have been delivered to the make-up area, they will be lifted onto a bag

breaker by means of a hoist. They will be discharged into the larger compartment of the

make-up tank and dissolved in water to form a 15% copper sulphate solution. The tank

mixer, installed in the larger compartment, will ensure complete dissolution of the copper

sulphate crystals during the make-up process. The variable speed peristaltic pumps will be

used to transfer copper sulphate solution to the cyanide detox circuit at a controlled rate.

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Hydrogen Peroxide Hydrogen Peroxide will be delivered to site in 1.2 tonne intermediate bulk containers at a

concentration of 60%. It will be pumped into the excess return water pond, together with

copper sulphate and hydrochloric acid, to aid in the destruction of residual cyanide prior to

discharge into the environment.

Activated Carbon Activated carbon will be delivered to site in 500kg bulk bags. A forklift will be used to

transport the bags from the storage area to the carbon attritioning area located in the carbon

regeneration circuit.

17.2.1.14 Oxygen and Air Services

Oxygen Two bought-out oxygen plants will be installed to supply a total oxygen demand of 5 tonnes

per day, as required for leaching of gold in the CIL circuit. Oxygen will be produced at a

pressure of 700kPa and a purity of least 96% v/v.

Air services Two air compressors, one duty and one standby, will deliver 1400Nm3/h of plant

compressed air and 300Nm3/h of instrument air at a pressure of 750kPa. Plant air will be

filtered and stored in a receiver, from where it will be distributed to mainly the cyanide detox

circuit and other areas in the plant for general usage. Instrument air will be distributed from

its receiver to all instruments throughout the plant.

17.2.1.15 Water Services

Water supply Rain run-off, a bore field and pit dewatering make up the raw water supply. Raw water will

be stored in a dam and pumped by two duty/standby submersible pumps, mounted on a raft,

to a pump station tank. Water will be transferred into a raw water tank, located in the plant.

Process Water Return water from the tailings storage facility and a top-up stream from the raw water tank

will constitute process water and will be stored in a pond located close to the processing

plant. A pair of duty/standby low pressure pumps will transfer process water from the pond to

the milling circuit and the gravity scalping screen for dilution. A pair of duty/standby high

pressure pumps will supply spray and hosing water throughout the plant.

Raw Water A 1500m³ raw water tank, situated in the plant, will be used to provide raw water to the plant.

The bottom part of the tank will be dedicated storage facility for the fire water. The fire water

system will consist of an electric jockey pump to maintain the desired pressure in the fire

reticulation system, an electric fire water supply pump and a diesel driven standby pump that

will start automatically when a power failure occurs.

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Other sets of pumps will distribute raw water throughout the plant for several uses, including

dust suppression in the crushing area, fluidizing water for the gravity concentrator, mill

cooling water, carbon transfer water, cathode wash in electrowinning, reagents make-up,

and gland water service. This water will be supplied from the top part of the raw water tank.

Potable Water Raw water will be treated in the potable water treatment plant to a drinking water quality and

will be stored in a potable water storage tank. Potable water will be pumped to a

hydrosphere, from where it will be distributed to the safety shower and potable water

headers. The hydrosphere will help maintain the required pressure in the respective

headers. In the event of a power failure the diesel driven potable water pump will fill up the

hydrosphere and thus ensure the availability of safety shower water and potable water.

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SECTION 18. PROJECT INFRASTRUCTURE

18.1 INTRODUCTION The selected Agbaou site is a greenfields site without any infrastructure except for the

existing Agbaou Exploration Camp. An existing access road currently passes through the

Agbaou village. The proposed infrastructure will support the mining, plant and construction

operations. The main infrastructure required for the development of the project will be:

• Raw water dam; • Raw water supply system; • Tailings storage facility; • New access road to the mining and plant site facilities, as well as the camp; • Site roads; • Camp accommodation and catering facilities, including laundry, potable water supply

system and sewage disposal unit; • Mining workshops, wash bay, refueling station and explosives storage; • Plant workshops; • Warehouses and lay down yards; • Plant administration buildings and medical facilities; • Assay laboratory; • Reagents storage building; • Change house; • Stand-by power plant; • Communications; • Security; • Sewage disposal; • Plant laundry.

This section details the facilities that are envisaged for the development.

The below site layout shows the position of the process plant, tailing storage and potential

water storage dams relative to each other and the surrounding topography.

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Figure 18.1: Site General Layout

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18.2 MINING FACILITIES The design of the mining facilities has been driven by the need to minimize costs, yet

provide substantial support systems for the life of mine.

The following will be allowed for in the mining facilities:

• Main workshop and repair facilities, to maintain the fleet of haul trucks and major mining equipment, complete with admin buildings alongside;

• Tyre repair and replacement bay; • Mining equipment re-fuelling centre; • Explosive storage, which will be located away from the main facilities; • Change house and security office; • Messing facility for mining personnel are provided in a separate building.

Raw materials, such as ammonium nitrate and/or explosive emulsions, primary explosives

used in the explosive manufacturing process will be brought to site by road, and stored until

required.

18.2.1 Mining Administrative Building The mining administration will have offices located within the mining workshop building. A

separate building will include the security offices as well as change house and all ablutions.

A separate building has been provided for a tyre bay office.

18.2.2 Mining Equipment Workshop A suitable workshop, which will have the capability of servicing the type and size of mining

equipment selected, will be constructed. A steel framed sheet clad building, 30m wide x

40m long x 7m high will be constructed with an initial configuration consisting of five front

bays:

• Two bays (8m wide x 10m long) for access of the trucks, one of the bays will be fitted with inspection pits;

• The remaining three bays will be utilized for heavy earthmoving equipment servicing such as the motor graders, wheel dozer, bulldozers and service trucks;

• A monorail electric overhead crane to traverse the 3 working bays and capable of lifting 5 tonne will be installed;

• Smaller bays to the rear of the building (8m wide x 5m long) will be utilized as a welding area, machine shop, drill rig repair and tyre repair services area;

• The mining maintenance offices will be placed beyond these; • A concrete wash down pad to ensure the equipment can be cleaned prior to

maintenance operations.

A smaller office has been included for drill rig repair and tyre repair services.

18.2.3 Mining Equipment Refueling Facility Fuel storage and refueling facility will be providing by an external contractor as well as the

necessary fire prevention system. The refueling bay will be supplied by a contracted fuel

supplier.

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18.2.4 Explosive Magazine The bulk explosives facility will be located adjacent to the mining facilities. The explosives

facility will consist of 2 used marine shipping containers as the store and surrounded by

earthen screening mounds and security fencing.

18.3 PROCESSING PLANT & ADMINISTRATION FACILITIES The following will be allowed for in the processing plant and administration facilities:

• In-plant access roads; • Plant administration buildings such as the security office, change house, workshop,

main administration offices, medical facility, assay laboratory and warehouses; • Sewerage treatment and disposal; • Water services inclusive of raw water abstraction, potable water and fire water; • Communications; • Security; • Vehicles to provide transport of staff from the accommodation areas; • Accommodation; • Power supply and distribution.

18.3.1 In-Plant Access Roads Plant roads will be stripped of organic material and compacted, allowing access to the

required plant areas. Drainage ditches and culverts will be placed in accordance with the site

drainage requirements.

18.3.2 Plant and Administration Buildings Buildings located in the plant area will consist of security offices, change house, plant mess,

process plant equipment workshop, control room, general administration offices, medical

facility, assay laboratory, reagent warehouses and spares warehouse.

18.3.2.1 Security Office and Change House

A security office and change house will be constructed at the access to the plant which will

have male and female change rooms fitted with lockers, showers and ablutions. There will

be a security search zone located next to the change house. The security search zone will

provide the security personnel with the means to conduct individual body searches, isolation

rooms and general scanning. The main access gate to the plant will have a low security

office for the control of vehicle access to the plant.

18.3.2.2 Plant Control Room

A dedicated plant control room is to be located on top of the CIL tanks. The control room will

house the SCADA system and provide operators with an elevated view of the entire plant.

The plant control room will be constructed using a modified used marine shipping container.

18.3.2.3 Plant Mess

A plant mess will be centrally located near the administration building and will be of a single-

story, prefabricated panel construction. This will be utilized by the process plant personnel.

Meals will be brought in from the camp kitchen.

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18.3.2.4 Process Plant Equipment Workshop

A suitable workshop will be established adjacent to the process plant to enable repair of

plant equipment. The workshop will be a steel frame building. Offices for warehouse,

maintenance and planning personnel will be provided in the form of a prefabricated building

housed within the steel framed building.

18.3.2.5 Administration Buildings

The administration building will be of a single-storey prefabricated panel construction. The

building will include general areas for engineering, administration personnel and offices for

the general manager, mining manager, plant superintendent, administration superintendent,

chief geologist, plant maintenance superintendent and chief security officer.

18.3.2.6 Assay Laboratory

SGS has submitted a proposal to manage the laboratory on ENDEAVOUR’s behalf. If

awarded the contract, SGS will provide laboratory design and engineering services. The

laboratory building will be of a single-storey prefabricated panel construction. The laboratory

will be fitted with equipment that will form part of SGS’s supply.

18.3.2.7 Medical Facilities

The equipped medical facility will allow for the treatment of any injuries during construction

and operation, as well as treatment of sick personnel. Medical services and administration

will be outsourced to a subcontractor.

18.3.3 Plant and Administration Warehousing A total of 3 warehousing facilities for storage of general spares, cyanide, reagents and

consumables will be provided.

18.3.3.1 General Plant Warehousing

This will be used to store all general items including mining and process plant spares. It will

consist of a 40m x 30m covered main store. Half the floor area will include heavy duty

racking evenly spaced across its width. Spacing between the individual racks has been

allowed such that a 5 ton forklift will be able to turn between them. The remainder of the

store floor area has been left open to store items of larger size and mass. The 30m x 40m

covered structure will include a personnel access door, prefabricated interior offices and

manually operated service door.

18.3.3.2 Plant Reagents & Consumables Warehouses

Due to the bulk nature of most reagents used in the plant, 2 dedicated warehouses will be

supplied to store these. Hydrated lime will be stored in the lime silo only.

Cyanide consumption was estimated to be 55 tonnes per month and due to the remote

location of Agbaou Project, independent covered warehouse facilities to store 3 months

cyanide stockholding will be provided.

The cyanide warehouse facility with a length of 13m x 13m width and will be located in the

plant area. The cyanide storage area will comprise of a concrete floor and will be located

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within the fenced, gated and locked high security area.

The balance of the reagents used in the process plant such as carbon, cyanide detoxification

reagents, hydrochloric acid, smelting fluxes, hydrogen peroxide, will be stored in a dedicated

warehouse with dimensions 15m x 28m.

18.3.4 Process Plant Site Drainage There are no permanent creeks or rivers on the site; however, during periods of rains, there

will be some surface runoff, open channel flow and standing water in low areas. Nominal

grading and ditching will be adequate to maintain a well-drained site. Culverts will be

installed where required, to ensure proper drainage and to maintain good road condition.

Surface runoff will be housed in sediment ponds.

18.3.5 Sewage Disposal The sewage plant is of the extended aeration activated sludge process type. Both the camp

and plant sites for the Agbaou Project will each have a totally independent sewage treatment

facility capable of handling the waterborne waste generated by about 200 persons per site.

Each of the sewerage treatment plants will be contained in a 15m x 15m fenced off area and

will be capable of producing final effluent discharge into the environment

18.4 ACCESS ROAD An allowance has been made for two access roads. The first road links the plant to the main

provincial road between the towns of Hire and Divo and is 9.17km whilst the second road

links the plant and the main camp and is 2.47km. Horizontal alignment of the roads was

based on the topographical survey. The route was plotted based on the shortest distances

from each of the respective destinations and points of departure while aiming to maintain

gradients of no more than 10% where possible. The roads were based on a 5.5m wide

gravel carriageway.

18.5 WATER SUPPLY

18.5.1 Water Supply Dam The design of this facility is based on meeting or exceeding agreed design criteria which

comply with World Bank and other international standards.

A water balance was developed and was used as the basis for sizing the water storage dam

and the raw water requirements. Raw water stored in the water supply dam will be pumped

to the process plant for make-up operations during the plant start-up and during periods

where the return water from the tailings storage facility is insufficient to meet plant

requirements. Groundwater abstracted from the open pits will also be discharged to the

water supply dam to augment the process water supply and avoid release to the

environment. The maximum inflow to the proposed open pits is expected to be

approximately 27L/s.

Knight Piésold identified five potential locations for the water supply dam, as indicated in

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Figure 18.2 below. The potential storage capacities, catchment areas and dam crest

elevations have been listed below in Table 18.1.

Table 18.1: Water Storage Dams – Catchment and Storage Capacities

Catchment Storage

Capacity

Dam Foundation

Elevation Dam Crest Elevation

(ha) (Mm3) (m) (m)

WSD Option 1 303 3.44 176 188

WSD Option 2 127 3.01 201 221

WSD Option 3 337 7.85 175 193

WSD Option 4 7605 3.68 139 161

WSD Option 5 332 3.55 155 175

Of the five options it was determined that 3 of the 5 locations were not suitable. Option 1 was

dismissed as it was located within a major drainage channel while options 4 and 5 were

considered too far from the open pit and plant site. With options 2 and 3 remaining, the dams

were modelled to identify fill quantities to be used in the cost estimate. The storage

capacities of the two viable options significantly exceed the storage requirements; therefore

the dams could be reduced to lower construction costs. Of the two viable options, Option 3 is

the preferred location as a smaller dam is required to store significantly larger volumes of

water. The storage capacity curves for the preferred option have been provided below in

Figure 2. From the information provided in Figure 18.3, the dam could be reduced to a crest

elevation of 183mamsl a dam height reduction of 10 meters. This will provide an estimated

storage of 1.4Mm3 with a 1m freeboard.

Two pumps mounted on pontoons will be utilized to recover water directly off the WSD,

pumping water to an intermediate pumping station and from there, into the process water

pond and raw water tank located in the plant. These pumps will be sized to cater for the

commissioning, dry and wet seasons where raw water demands vary.

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Figure 18.2: Water Storage Dams – Options 1 to 5

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Figure 18.3: Water Storage Dam Option 3 Capacity Curves

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18.5.2 Potable Water Distribution Raw water will be supplied to the potable water treatment plant from the plant raw water tank

and in turn through a pipe line from the raw water dam. There will be a separate water

treatment plant at the process plant to the one at the mine village. The plant is designed to

supply 5m3/h of potable water.

Potable water will be reticulated to all areas of the plant including a safety shower header

tank.

The total volume of groundwater available from dewatering of the open pits will not be

sufficient to meet the process water requirements of the operating plant. As a result,

additional sources of process water supply will be required.

Sufficient groundwater resources are available from dedicated potable water supply

boreholes in the area to provide potable water for human consumption. Current groundwater

quality indicates that the resource is suitable for potable use according to the WHO

Guidelines for Human Consumption.

Open pit dewatering during mine operation will have negative impact on Agbahou village

water system and augmentation of supply may be required.

18.5.3 Fire Water Distribution There will be an electric and diesel powered fire water pumping system. The electric

powered pump will be used in the event of a fire. The diesel pump will be used in the event

of a fire where electrical supply is unavailable. A jockey pump will be provided to maintain

the pressure in the fire water header during normal plant operation. An alarm will be

sounded at the plant site for low system pressure.

The fire water system will consist of a fire water distribution system with hydrants located

within the plant site and ancillary building areas. Hose cabinets will be placed at the fire

hydrant locations and the system supplemented with portable fire extinguishers placed within

the process facilities.

18.6 COMMUNICATIONS A satellite communication system will be provided for the initial phases of the project,

following which ENDEAVOUR will install a system according to their needs.

18.7 SECURITY The entire Agbaou plant site will be encompassed by a 2.5m x 75mm x 75mm diamond

mesh wire fence. Plant access will be via gates located on access roads to the site.

Additional fencing will be provided for further safety and security within process plant areas,

such as ponds, power plant, fuel storage, gold room area, transformers and substations, as

required.

Furthermore, the plant will be fitted with CCTV cameras installed at strategic locations to

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provide for monitored surveillance. The cameras will be integrated with the plant’s overall

network. Views from the cameras will be fed to the security control room and the security

manager’s office.

18.8 ACCOMMODATION (MAIN CAMP) A mine camp will be constructed to accommodate approximately 200 people.

A permanent camp will be constructed to house senior and junior staff members. Staff

accommodation will consist of the following pre-fabricated modular buildings:

• Six off 2 bedroom senior staff units; • Eighteen off 10 room units. Each unit will have its own bathroom.

A prefabricated panel construction will be utilized for all the structures. In addition to the

above mentioned facilities provision will also be made for the following:

• Kitchen and camp dining room; • An entertainment area consisting of a bar and TV room; • Laundry; • Potable water plant; • Sewage disposal plant.

18.9 POWER SUPPLY (GRID) The electrical power supply will be from the national grid.

Three options were considered. Making reference to the plant load list document number

SP462 1000 1E8 001, and document number AGB-AP001-PP (prepared by BEC for

ENDEAVOUR) the selected option meets the required level of electrical power estimated for

the plant at Agbaou in accordance to the grid operating criteria. For further details regarding

the BEC report and load list refer to Appendices 18-A and 18-B.

Selected option - Proposed power supply includes the following:

• New 90/33kV – 20MVA Transformer bay at the hire substation; • Dedicated 15km OHL feed to Agbaou; • Terminal substation with 33/11kV – 15MVA transformer; • Installation of reactive energy compensation 5 steps 2.4MVAR a maximum of

18.63MW can be achieved compared to the available 13.48MW; • All civil work in the delivery substation, overhead line foundations, transformer bay

foundation and cable trenching. The SENET battery limit is the incoming terminals of the incomer.

Final routing of overhead line routing is dependent on the topographic evaluation of the final

routing of the overhead line.

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18.9.1 Standby Power Plant Two options were considered:

Option 1: Provide a full estimate for the supply and installation of a power plant. The loads were

estimated using the peak power excluding standby. This option was not economically viable

and a second option was considered.

Option 2: 2.4MW of standby power was allocated to the plant critical infrastructure equipment. The

option of connecting the 400V generator sets to the 400V motor control centres (MCC’s) was

considered, i.e. suitable for emergency backup or assistance during brown out periods

On the operating principle, our proposal is based on automatic synchronization and load

transfer after the grid comes back in a stable mode. Final detail list of critical load will be

defined at execution phase of the project.

The fully containerized solution offered by Africa Power Systems is as follows:

• Two off 1200kW CAT SR5 Generators selected for the application.

This solution will be supplied without a day fuel tank, electrical distribution switchgear and

other requirements associated with such a plant. The generator sets will be designed to

supply a total of 2.4kW of prime power at PF of 0.8 for the shutdown of plant during a power

outage and to sustain all critical loads for the full duration of the power outage if required.

Final standby plant design will be confirmed at execution phase.

Design criteria are as follows:

• Generator cooling method - Radiator; • Power per engine(Prime) - 1200kW; • Generator voltage - 400V; • Frequency - 50Hz; • Power factor - 0.80; • Service voltage - 400V; • Fuel system - Electronic Unit Injection; • Lower heating value liquid fuel - 42,780KJ/kg.

18.9.2 Power Distribution An 11kV supply cable from the main intake substation at Agbaou will feed power to the main

11kV distribution panel (SRK Consulting, SENET, Knight Piésold Consulting, 2012 Appendix

18-C). Power distribution to the SAG mill motor and ball mill motor will be done at 11kV.

The process building and power system modules will generally include outdoor oil-filled

transformers, MCC’s, main distribution boards (MDB’s), local circuitry at 400V, distribution

panels and local control devices. All electrical distribution will be in cable trays using steel

wired armoured PVC sheet cables.

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The process plant, site ancillary facilities switchgear and electrical equipment will be installed

in a dedicated substation. Remote MCC’s will be containerised.

In non-process areas, such as the administration building, sewage treatment plant, fuel

storage facility, water tanks and workshop complex, a combination of armoured-type cable

and rigid galvanized steel conduit and wire system will be used in exposed areas.

MCC’s will be complete with motor management system, motor starters, contactors,

transformers, panels and circuit breakers.

Final routing of overhead line to the camp, mining workshop, tailing return water and raw

water is dependent on the topographic evaluation at execution phase.

18.9.3 Fuel Storage Design of the plant and camp fuel storage facility will be based on providing power for back-

up purposes only.

Plant power is based on a fuel consumption of 230 litres per hour for a 1,500kVA generator.

Thus assuming plant standby power of 2 generators we would have a full draw of 11kl/day.

Two options exist for refilling the fuel storage facility at the plant. Firstly fuel can be moved

by the diesel bowser from the diesel facility at the mining workshop/filling facility. Secondly

tankers can the bring fuel in from Abidjan. Based on this fuel provision of 5.5 days will be

allowed, with a further 12hrs being provided by the individual tanks integral to each

generator container.

Storage facility will include:

• 1 x 60m3 Diesel storage tank; • Bulk lubes storage; • All civils - concrete bund; • Distribution piping.

In addition to the plant fuel storage facility a similar 60m³ diesel storage tank will be installed

at the mining/construction camp. Based on the usage here the fuel will be able to run the full

camp for 10 days before requiring refilling.

Please note that fire prevention system in respect of diesel storage has not been provided.

The design and provision of such a system will need to done with consultation with

specialists in this field.

18.10 TAILINGS STORAGE FACILITY

18.10.1 Introduction and Design Criteria

The design of tailings storage facility (TSF) is based on meeting or exceeding agreed design

criteria which comply with World Bank Standards and other international standards.

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The design criteria are summarised in Table 18.1 below.

Table 18.2: Design Criteria – Tailings Storage Facility

Description Criteria Comment

Capacity 12.1 million tones

(8,963,000 m3)

Life of mine (LOM) 8.1 years

Average tailings throughput 1.2x106 tpa LOM average

Tailings throughput -

Saprolites 1.6x10

6 tpa

SENET material balance (for

operational years 0 to 4.35)

Tailings throughput -

Bedrock 1.34x10

6 tpa

SENET material balance (for

operational years 4.35 to 8.1)

In-Situ dry density 1.35 t/m3

Assumption based on numerous

gold tailings samples tested in

South Africa

Average calculated rate of

rise: 2.2 m per year

Based on available

topography and option 2

layout

In this case it is proposed that the

impoundment is downstream

constructed rockfill. Rate of rise

will not affect overall stability.

Downstream slopes 1V : 3H Facilitates rehabilitation and

meets stability requirements

Individual upstream slopes 1V : 2H Meets stability requirements

Overall upstream slopes 1V : 2.4H 4 meter bench at each phase

Area of footprint 1.0km2

(100 Ha)

Tailings Particle Size

Distribution 80%<75micron SENET mass balance

Slurry density (Saprolites) 1.296 t/m

3 (35.6 %solids by

mass) Slurry calculation

Slurry density (Bedrock) 1.315 t/m

3 (37.1 %solids by

mass) Slurry calculation

Particle SG (Saprolites) 2.79 SENET mass balance

Particle SG (Bedrock) 2.82 SENET mass balance

Slurry delivery rate 318 m3/hr to 404 m

3/hr MDME / Slurry calculation

Slurry distribution line 315mm OD HDPE, SDR 17

Based on approximate minimum

settling velocities. To be confirm

during detailed design

Maximum wall height 18.5m Volumetric and rate of rise

analysis

Water removal Pumping barge (2 pumps) Common practice – minimal &

maximum volumes monitored

Downstream construction –

rockfill / random fill Good practice

Waste rock available from open

pit overburden

Upstream construction –

select fill

Lateritic material

Select fill (Lateritic material)

available from borrow pits or

from open pit overburden

Minimum Factor of Safety :

1.5 (Static)

1.0 (With 1:1000 year

seismic event)

World Bank Standards /

Good practice

Analysis of embankment walls

required

Design flood

1 : 100 Year recurrence

interval 24 hr duration.

163mm

Adequate freeboard to be

provided by providing a minimum

1 metre freeboard above the

normal pool operating level at all

times. Additional water will be

removed via TSF pump barge.

Spillway Good practice

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18.10.2 Site Selection and Access to the Site Two TSF locations were identified for the EOS. Both sites differ from the original location

selected for the 2008 Bankable Feasibility Study. For the purpose of this report the sites

have been labelled as Option 1 and Option 2. Condemnation drilling was undertaken to

determine that both sites were free of mineralisation. Condemnation drilling results were

received in January 2012 for Option 1 and drilling for Option 2 was completed in mid-March

2012.

The selected TSF site (Option 2) is located directly north of the proposed plant site, where

the southernmost embankment wall is less than 300 metres from the nearest plant structure,

thereby presenting a high risk situation. The topography drains to the east and is enclosed

by hills located on the South West and North East corners of the proposed TSF. Access to

the TSF from the proposed plant site currently exists with the road passing by the South-

West corner of the TSF. Due to the close proximity of the TSF to the plant site a second

road would be recommended from the plant site to the South-East corner of the TSF as this

is the proposed route for the tailings slurry pipeline. This road would provide a pipe route for

the tailings delivery and return water pipes.

Site selection for the Option 1 TSF location was based on a site visit in October 2011. This

site selection took into account the following parameters:

• General topography; • Distance from the plant; • Surface geotechnical conditions in the footprint zone; • Watercourse locations; • Position in relation to other mine infrastructure; • Geology of the site.

The site selection for Option 2 was based on additional review of the topography and the

revised capacity requirements for the TSF. This option also considered the parameters

indicated for Option 1.

As the Agbaou project progresses to the next phase of development, it is highly

recommended that a new topographic survey be completed even though a site specific

ground survey was completed for Option 2. The new survey would highlight drainage paths

more thoroughly, correctly identify elevations and allow for the site selection to be improved.

As it stands, the preferred option based on the topographic survey and the recent ground

survey provided to Knight Piésold is the second option. Option 2 utilises the existing terrain

more efficiently, thereby significantly reducing the volume of fill required to develop the TSF.

18.10.3 TSF Construction and Wall Raising Procedure The TSF will be constructed in three phases. The first phase has been sized to provide

capacity for roughly two years of the saprolite ore feed while maintaining a supernatant pond

capacity of 150,000m3 after approximately 1.4 years and 1.5 meters of freeboard plus

spillway capacity. The succeeding phases (Phase 2 and Phase 3) will provide the capacity

for the remainder of the 8.1 years life of mine (LOM) when the ore feed is from bedrock.

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Table 18.3 identifies the intended ore feed, duration, tonnage, expected volume and Phase.

Table 18.3: TSF Available Capacity per Ore Feed Type

Ore Feed Duration

(years)

Expected Tonnage

(tonnes)

Equivalent Volume

(m3)

Phase

Saprolite 2.05 3,292,850 2,464,800 1

2.30 3,703,000 2,683,300 2

Bedrock 3.75 5,146,250 3,722,200 2 & 3

Totals 8.1 12,096,250 8,890,200 Final

The Phase 1 starter embankment will be constructed to elevation 249.5mamsl starting from

a low point of 243mamsl. The Phase 1 embankment will provide capacity allowance of

3.29Mm3 (2.44Mt) of tailings storage, 150,000m3 of supernatant pond and 589,000m3 of

freeboard for various storm events plus 1.5m. The large capacity of the freeboard is to

ensure multiple large event storms can be stored within the TSF without discharge to the

environment. It is expected that the 150,000m3 capacity of the pond is the preferred

operating volume and that this volume should be maintained to ensure sufficient distance is

maintained between the pond and the embankment walls. Additional volume will be present

at certain times. This will typically occur during the rainy seasons; therefore the pond

volume will potentially be reduced during the dry months. The minimum that must be

maintained will be dependent on the tailings beach development, pond location and the

return water pump requirements. Construction sequencing is provided in Table 18.4 below.

Table 18.4: TSF Construction Sequencing

TSF

Phase

Construction

Start(2)

Durati

on

(years)

Crest

Elevation (m)

TSF

Raise(1)

(m)

Maximum

Capacity (tonnes)

Phase 1 Year: -1 2.05 249.5 6.5 3,292,850

Phase 2 Year: 1.0 3.45 255 5.5 5,288,600

Phase 3 Year: 4.5 2.6 258 3.0 3,514,800

LOM 8.1 211.5 18.5 12,096,250 (1) The TSF is developed by constructing 4 perimeter dams; the TSF raise is for largest dam only as each dam starts at a

different elevation.

(2) Assume construction for each phase requires a maximum of one year to complete.

To reduce CAPEX costs it is intended to construct Phase 1 using material from localised

borrow pits that reduce the haul distance. Once the mine is operating under OPEX

conditions, the intent is to construct the TSF using overburden material from the open pits.

The difficulty will be to provide sufficient lateritic material for the 3 metre wide strip located on

the upstream slope of all phases. The geotechnical investigation during the design phase will

identify and confirm suitable material. Alterations to the preliminary design would be required

if certain materials are not available.

The preliminary design has assumed suitable material will be available from local borrow pits

inside the TSF basin for Phase 1 and that the overburden will be available for use during the

later phases. If additional overburden material (random fill of non-lateritic type) is available,

construction of the Phase 2 and 3 random fill sections can commence earlier than currently

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proposed.

The impounding embankments will be comprised of either waste rock overburden from the

open pit (Phase 2 & 3), with a zone of selected, finer, well graded, low permeable material

on the upstream face of the embankment. The crest will vary from 14 m wide for Phase 1 to

10m wide for the later phases. The downstream slope will be 1 vertical: 3 horizontal. The

upstream face of the embankment will be constructed at a slope of 1:1.5 with 4 m wide

intermediate bench between phases with individual slopes between benches at 1:2.5

The area covered by tailings at the end of phase 1 is shown in Figure 18.4 along with the

supernatant pond location. Similar figures for phase 2 and 3 can be found in the EOS (SRK

Consulting, SENET, Knight Piésold Consulting, 2012 Appendix 18-D). Stage /capacity and

Rate of Rise Curves can also be found in the EOS (SRK Consulting, SENET, Knight Piésold

Consulting, 2012 Appendix 18-E).

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Figure 18.4: TSF Phase 1

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18.10.4 Stability Analysis The computer program Slide, version 5.029, from Rocscience was used to determine the

stability of the TSF. The Bishop Simplified method was used to determine the factor of

safety (FOS) for a circular slip surface. The current information suggests that the area of the

TSF is relatively flat; therefore no wedge type failures were investigated at this time. The

wedge type failure plane will be investigated once more topographical information and more

accurate material properties are available. A further analysis was conducted to investigate

the effect of including pore water pressure by means of the B-bar method. The pore water

pressure was obtained from the seepage analyses and converted to B-bar factors and

added into the analyses. In general it is expected that adding in a B-bar factor will lower the

factor of safety.

Table 18.5 presents the material properties that were used during the analysis. An

additional analysis was conducted with a horizontal and vertical seismic loading of 0.1g,

which was expanded to investigate the influence of only a horizontal seismic loading. Table

18.6 presents the summary of the different scenarios that were investigated and Table 18.7

presents the results that were obtained from the analyses. The analyses were based on the

current dam design where the dam will be constructed in three phases.

Table 18.5: Material Properties

Material Cohesion c

(kPa)

Friction Angle

(Degrees)

Unit Weight

(kN/m3)

Tailings 0 31 13.5

Random Fill 0 36 18.0

Silty Gravel 0 31 15.0

Laterite 16 30 16.0

Select Fill 16 30 16.0

Table 18.6: Definitions of the Analyses Scenarios Scenario Definition

Rev 1 Downstream slope, no additional seismic loadings (tailings at final height for that

particular phase, 1m freeboard)

Rev 2a Downstream slope, horizontal and vertical seismic loading (tailings at final height

for that particular phase, 1m freeboard)

Rev 2b Downstream slope, horizontal seismic loading (tailings at final height for that

particular phase, 1 m freeboard)

Rev 3 Upstream slope, no additional seismic loadings (tailings at final height for previous

phase, )

Rev 4a Upstream slope, horizontal and vertical seismic loading (tailings at final height for

previous phase)

Rev 4b Upstream slope, horizontal seismic loading (tailings at final height for previous

phase)

Rev 5 Downstream slope, no additional seismic loadings (tailings at final height for

previous phase)

Rev 6a Downstream slope, horizontal and vertical seismic loading (tailings at final height

for previous phase)

Rev 6b Downstream slope, horizontal seismic loading (tailings at final height for previous

phase)

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Table 18.7: Results of Analysis

Scenario Factor of Safety

Phase 1

Factor of Safety

Phase 2

Factor of Safety

Phase 3

Rev 1 2.23 2.04 2.27

Rev 1 (B-bar) 2.29 2.02 2.07

Rev 2a 1.68 1.57 1.73

Rev 2b 1.60 1.49 1.68

Rev 2b (B-bar) 1.66 1.46 1.47

Rev 3 1.69 5.02 5.51

Rev 4a 1.37 3.13 3.39

Rev 4b 1.35 2.93 3.14

Rev 5 2.31 2.21 2.27

Rev 6a 1.75 1.69 1.72

Rev 6b 1.71 1.63 1.67

The factors of safety that were obtained during the analysis are all above 1.3 which is the

minimum FOS over the short term (during operation) and above 1.5 which is the minimum

for the long-term or after closure. The minimum FOS for seismic loading is 1. The minimum

FOS that was obtained for the analyses was 1.69 for Phase 1 Rev 3 (Upstream face storing

no tailings) and 1.35 for Phase 1 Rev 4b (upstream face storing no tailings and adding only a

horizontal seismic load). On average the horizontal seismic loading returned lower FOSs

than for the combination of horizontal and vertical loadings.

The FOSs obtained from the two scenarios that were analysed utilising the B-bar method are

very similar to the FOSs obtained from the standard analyses. Usually one would expect the

inclusion of pore pressures to result in lower FOSs but in this case the location of the

phreatic surface played a role in obtaining similar FOS values. The standard analyses were

conducted using very high phreatic levels to simulate an event where the precautions to

lower the phreatic surface proved inadequate. The phreatic level utilising the B-bar method

was located at the level obtained from the seepage analyses which were considerably lower

than for the standard analyses. This implies that the TSF is very sensitive to the phreatic

level and therefore care must be taken to implement and sustain good water management

practices.

18.10.5 Tailings Distribution and Deposition The average slurry flow rate from the plant for saprolite ore feed will be approximately

404m3/h at a slurry mass concentration of 35.6% once full production is achieved. The

average slurry flow rate from the plant for bedrock ore feed is expected to be slightly less at

318m3/h at a slurry mass concentration of 37.1%. Both rates are based on an average

yearly operation of 360 days per year – 24 hours per day. The pumping distance from the

plant to the valve slab at the South corner of the TSF is less than 1000m dependant on the

final plant arrangement and location in relation to the TSF. There is very little (less than 5 m)

to no static head to pump against initially with the plant situated between 245mamsl and

250mamsl and the Phase 1 crest at 249.5mamsl, but this will increase to a maximum of 13m

when the TSF embankment is raised to its final elevation of 258mamsl. The pipeline length

from the valve slab to the furthest point on the distribution ring mains is 2000m. The

elevation difference between the valve slab and the crest of the impounding embankment

(Phase 3) is approximately between 9m and 14m.

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18.10.6 Tailings Disposal Method The method of tailings discharge will be sub-aerial deposition. This will allow the tailings to

form a beach above the water level, sloping gently towards the supernatant pond. As the

tailings discharge onto the beach, small streams will develop that allow the tailings to settle

and segregate from the water. The general practice for subaerial deposition is to have

multiple discharge points that break the flow energy and promote gravitational settling.

Maintaining a lower discharge velocity will be beneficial to the water quality and tailings as

the low velocity will expose beached tailings to oxygen and water thereby helping in the

degradation of cyanide compounds that will be used in the ore extraction process. Multiple

spigots and a well-managed rotational deposition cycle around the dam perimeter will help

control the geometry of the embankments and location of the supernatant pond. This can

also reduce the risk of loss of freeboard by preventing the pond from encroaching on the

embankments. Controlling the location of discharge can also be used to position the pond to

the required location. Multiple spigots around the facility will also allow the deposition of

tailings to be rotated between different locations, which will allow the newly discharged

tailings to bleed, dry and consolidate while tailings slurry continues to be discharged to other

zones. The frequency of discharge point rotation and the number of zones will depend on

the how the tailings reacts to the deposition cycle, the season (wet season versus dry

season), the tailings production rate, tailings drying characteristics (saprolite versus bedrock

ore feed) and the tailings storage facility’s overall shape.

The general process that will be used for multiple sub-aerial deposition of tailings slurry is as

follows:

• Tailings will be discharged onto a portion of the gently sloping beach creating low energy braided flows towards the supernatant pond;

• Once a proper beach has formed, the discharge point(s) are moved to another location of the facility on a managed rotational cycle to re-establish the beach and to maintain the pond location. By moving the discharge locations the newly established beach is allowed to drain, bleed and dry;

• The upstream under-drainage systems will intercept and convey the seepage out of the tailings and reintroduce it back into the facility to the pond for use in the plant. This process will promote faster drying of the tailings by removing water from covered tailings in the zone near the outer wall;

• The cycle is repeated with new tailings deposited on the dried tailings beach. It will be important to identify a deposition plan early on to develop the TSF appropriately;

• The supernatant ponds location will be established by the tailings deposition and the size (volume) will be adjusted and maintained by the pump barge that feeds the plant processes.

18.10.7 Hydrology and Water Management The mean annual precipitation (MAP) at Agbaou is 1259mm, and the average annual

evaporation is 1363mm. The TSF has been designed to contain the net storm water runoff

from the tailings storage facility arising from the 1:100 year storm (163mm in 24 hrs, as

derived from the rainfall records).

Rain water will be retained on the TSF, which by virtue of the rock fill impoundment will have

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adequate freeboard to store multiple storm events and the required volume based on proper

operation of the TSF. The preliminary design includes a spillway on the northern end of the

dam that is located on the South-West side of the TSF. The spillway has been included in

the preliminary design as it is considered a good engineering practice. Additionally the

spillway will be constructed in in-situ soil rather than fill, which is also generally accepted

practice.

Diversion canals will divert any clean runoff from the catchment upstream of the facility

around the TSF to either the South-East or South-West depending on the location.

18.10.8 Geotechnical Conditions at the TSF Site The TSF occupies a low lying area north of the proposed plant with high hillsides located on

all corners of the facility. A geotechnical investigation was not completed for the EOS due to

time constraints. Therefore the study has developed a preliminary design based on a grid of

30 test pits that were excavated by hand over an area identified in the 2008 Feasibility

Study. The 2008 TSF location is approximately 1500m east of the TSF location proposed

within this study. The test pits showed a reasonable degree of uniformity across the

designated area, and comprise typically of either a silty sand or silty gravel (in a few cases

silty clay) overlying hard laterite at depths between 1m and 3m. In most cases refusal was in

the laterite. A full geotechnical investigation will be required during the design phase.

Where the test pits were able to penetrate through the laterite, a low plasticity clay was

encountered below it.

18.10.9 Water Balance A monthly water balance for the tailings circuit was carried out using tailings behaviour

characteristics typical of South African gold tailings since data was not available locally. The

water balance was calculated for the situation of average rainfall years at the end of phases

1, 2 and 3 of development of the facility. The detailed water balance can be found in the

EOS (SRK Consulting, SENET, Knight Piésold Consulting, 2012 Appendix 18-F).

18.10.10 Return Water System A return water pump station located on a barge inside the TSF with both duty and standby

pumps will be capable of pumping water at a rate of 400m3/h to meet the plant requirements.

The return water pipeline will have a cross connection with the tailings slurry pipeline to be

used for flushing and cleaning the slurry line.

18.10.11 Safety Classification A safety classification in accordance with SANS 10286 Clause 7.4.1 has rated the TSF as a

“Moderate Hazard” because the Zone of Influence of the TSF could affect nearby residents,

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18.11 COUNTRY INFRASTRUCTURE Figure 18.5: Map of Country

18.11.1 Road National road conditions between Abidjan and Agbahou village are described with

photographic representation in the road survey by Bollore Africa Logistics (SRK Consulting,

SENET, Knight Piésold Consulting, 2012 Appendix 18-G). Bridges found on route as well as

clearance dimensions are also included in the Bollore Africa logistics report.

18.11.2 Port Facilities Port facilities are as follows:

• The Port of Abidjan is the main port in Côte d’Ivoire. San Pedro also has a port

facility;

• With a total of six kilometers of quay, the Port of Abidjan has 34 berths including

berths dedicated for timber, cereals, fruits, petroleum products, and containers.

• The Port of Abidjan can accommodate vessels up to 260m long;

• The port contains 407.6 thousand square meters of open storage and 143.5m2 of

covered warehouses and sheds;

• Three berths specialize in container-handling, and one berth is devoted to roll-on/roll-

off cargoes. All of the port’s wharves are connected to the rail network.

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Figure 18.6: Abidjan Port Map

Labels to figure above are: 1 – North Quay; 2 – West Quay; 3 – South Quay; 4 – Container Terminal; 5 – Container Terminal; 6

– Fishing Quay; 7 – Banana Carrier Wharf; 8 – Fertilizer Factory “Sieving” Quay; 9 – Ro/Ro Quay; 10 – Mineral Berth; 12 –

Fertilizer Factory “Sieving” Quay; 13 – Oil Berths.

Table 18.8: Abidjan Port Facilities / Storage Capacities Item Capacity

Bonded warehouses 134,614 m2

Container storage area 70,000 m2

General cargo storage yard 250,000 m2

Plugs for reefer containers 350

Free storage zone for cargo in transit to Burkina Faso 3,204 m2

18.11.3 Rail There is no rail infrastructure that the project can benefit from.

18.11.4 Air The Abidjan International Airport (the Port Bouet or Félix Houphouët-Boigny airport) is

located in Abidjan and is operated by the civil government. Although the airport does not

offer flights to and from the United States, it does offer flights throughout Africa, Europe and

the Middle East. Airlines serving the airport are: Air Algerie; Air Burkina; Air France; Air

Ivoire; Air Mauritanie; Air Senegal International; Brussels Airlines; Cargolux; Egyptair;

Ethiopian Airlines; GIB; MEA; Panaf; Royal Air Maroc; South African Airways, and Tunisair.

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18.12 LOGISTICS AND TRANSPORT

18.12.1 Logistics Transit Time With a large portion of cargo being shipped out of South Africa the below time line was used

for project scheduling purposes.

Figure 18.7: Logistics Transit Time Summary

18.12.2 Routing Abidjan to Agbaou Village A route survey was conducted by Bolloré Africa Logistics (SRK Consulting, SENET, Knight

Piésold Consulting, 2012 Appendix 18-G).

Two different routes were surveyed between Abidjan and Agbaou (as shown in Figure 18-5),

these being:

• Route 1 – Abidjan to Agbaou (D) via Toumodi (B) with a distance of 305.5km; • Route 2 – Abidjan to Agbaou (D) via Divo (E) with a distance of 216.2km.

Figure 18.8: Map Showing Travel Route Options to Agbaou

Key: B – Toumodi

C – Oumé

D – Agbahou

E – Divo

F – Tiassale

G – Abidjan

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Route 1: From Abidjan to Agbaou via Toumodi (points G, B, C, D) there is a double highway for

141km, then 55.5km of main road all in good condition. The maximum size of freight for this

section is: 4.90m (height) x 7.00m (width) x 250tonnes (weight).

For the 52.3km section from Toumodi to Oumé (points B – C) the road has sections in a

poor state with missing asphalt. Poorly installed electrical and phone cable crossings in

urban areas would require specialized services for the dismantling when transporting

larger loads. The maximum size of freight: 4.90m (height) x 7.00m (width) x 250tonnes

(weight).

From Oumé to Agbaou, which is 56.7km, the road deteriorates further with repairs required

to sections of laterite where the asphalt is missing.

Route 2: From Abidjan to Agbaou via Divo (points G, F, E, D) there is a double highway for 109km,

then 85km of main road to Divo, all in good condition, bar a few potholes.

The 16.1km section from Divo to Didoko the road deteriorates further with a particularly bad

section in Didoko. Approximately 5km outside Didoko is the turnoff to Agbaou. Both routes

would require an escort for any freight with a width greater than 3m.

18.12.3 Documentation In order to ensure effective management of logistics and for all parties’ expectations of the

project to be met, a written guide and plan will be a requisite to enable better understanding

of how the logistics will be performed. This logistics execution plan will outline the

responsibilities of all stakeholders (contractor / company / suppliers / other interested

parties) and will indicate how cargo management and control from time of receipt by the

contractor to time of delivery is to be achieved.

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SECTION 19. MARKET STUDIES AND CONTRACTS This section is not applicable to this technical document.

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SECTION 20. ENVIRONMENTAL STUDIES, PERMITTING AND

SOCIAL OR COMMUNITY IMPACT

20.1 INTRODUCTION An Environmental Impact Assessment (EIA) and an Environmental Management Plan (EMP)

was conducted for the Agbaou Project by African Mining Consultants (AMC) in 2008 (SRK

Consulting, SENET, Knight Piésold Consulting, 2012 Appendix 20-A). The report was

compiled in line with applicable international standards and in accordance with Côte d’Ivoire

legislation.

The EIA covered the social and local environment prior to project development and the

potential impacts, positive and negative, the project would have on the environment.

Some of the potential positive impacts include, but are not limited to:

• Creation of approximately 300 jobs; • Local infrastructure development and increased standard of living; • Economic growth in local areas and Côte d’Ivoire through the service construction

and manufacturing sectors; • Increased national income through taxes, royalties and fees; • Training and essential skills to increase the contingent of local employees; • Social development projects.

Some of the potential negative impacts include, but are not limited to:

• Land clearing of approximately 410ha and the resultant loss of natural habitats of the affected flora and fauna;

• Changes to natural drainage patterns; • Contamination of soils, surface water and ground water; • Localized lowering of the ground water levels; • Degradation of local air quality through dust generation and vehicle fumes; • Increased noise in the local; • Increased traffic volumes and the potential safety implications of this on local

communities; • Generation of acid mine drainage; • Increase in HIV and AIDS prevalence; • Increased immigration into the area could lead to conflicts resulting from competition

for jobs and resources; • Loss of agricultural land and the subsequent loss of financial security for the affected

farmers.

The EIA identifies the necessary management measures required to mitigate the identified

environmental and social impacts. These form part of the EMP. In addition a Relocation

Action Plant (RAP) was developed along with the budget to cover the relocation costs of the

affected communities.

The detailed report is as per the previous NI 43-101 submission compiled by MDM for the

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Agbaou Project on behalf of Etruscan Resources, submitted in September 2009.

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SECTION 21. CAPITAL AND OPERATING COSTS

21.1 PROJECT REQUIREMENTS

21.1.1 Introduction The aim of the capital and operating cost estimates is to provide costs to an accuracy level

in the opinion of those listed in 21.1.3, of ± 10% that can be utilized to evaluate the

economics of the Agbaou Project when treating 1.6Mtpa. All costs are presented in United

States Dollars (US$).

21.1.2 Scope of the Estimate The capital and operating cost estimates were developed for a conventional open pit mining,

carbon-in-leach process plant and support infrastructure for an operation capable of treating

1.6 million tonnes of ore per annum.

21.1.3 Responsibilities The responsibilities for the estimate are listed below, but in broad terms the following

applied:

• Plant and infrastructure – SENET; • Mining capital and operating cost – SRK; • Tailings management facility – Knight Piésold; • Raw water storage facility – Knight Piésold; • Camp water supply Knight Piésold.

21.1.4 Estimate Accuracy The level of accuracy of the capital cost estimate (excluding the raw water storage facility) is

in the opinion of the responsible companies listed in 21.1.3 within the ±10% of the overall

project costs as of the 1st Quarter 2012. It does not include any escalation factors.

Since a preliminary level design was not completed on the raw water storage facility the level

of accuracy cannot be considered within +10%.

21.1.5 Exclusions The following were not included in this estimate:

• Scope changes; • Escalation beyond the 1st Quarter of 2012; • Financing costs; • Schedule delays such as those caused by:

o Scope changes; o Unidentified ground conditions; o Labour disputes; o Environmental permitting activities; o Receipt of information beyond the control of the EPCM contractors.

• Permits;

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• Sunk costs; • Force majeure.

21.1.6 Escalation No escalation has been allowed for in the estimate. The EPCM contractor’s rates reflect

rates expected in 1st Quarter of 2012.

21.1.7 Exchange Rates The exchange rates used are shown in the table below:

Table 21.1: Exchange Rates Exchange Rates Rate of Exchange

EUR : ZAR 10.50

GBP : ZAR 12.00

USD : ZAR 7.50

CA$ : ZAR 7.50

AU$ : ZAR 7.50

CFA : US$ 515.00

AU$ : US$ 1.03

21.1.8 Taxes, Duties and Fees This section describes the applicable taxes and duties in Côte d’Ivoire and how they would

be applied to the Agbaou Project.

21.1.8.1 Tax and Duty Rates for Importation of Equipment and Materials into Côte d’Ivoire

All mining, process plant and construction equipment, materials and spares will be

exonerated from import duties during the construction phase prior to production as per

document “Law No.95-553”, dated 18th July 1995 pertaining to the mining code that is

administered by the Ministry of Mines and Energy. A provision is also made for VAT

exoneration; however this does not extend to any goods purchased locally. An import tax

(regional statistical tax and two community / solidarity taxes) of 2.5% does apply to all

imported items and was taken into account during initial capital calculations.

21.2 CAPITAL COST ESTIMATE

21.2.1 Capital Cost Summary The total estimated cost of bringing the Project into production is US$158,939,551 and is

inclusive of US$11,777,985 contingency and US$5,039,749 working capital. Table below

gives a summary of capital requirements.

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Table 21.2: Capital Cost Summary US$ % US$

Process Plant Direct Costs

Machinery & Equipment 17 732 002 5% 18 618 603

Civils & Earthworks 11 556 402 12% 12 887 449

Structural Steel & Platework 5 803 972 10% 6 384 369

Piping & Valves 1 433 928 10% 1 577 321

Electrical & Instrumentation 5 094 739 12% 5 686 298

Transportation 5 554 000 15% 6 387 100

Subtotal 47 175 044 51 541 140

Infrastructure Costs

Tailings (Start-up only) 5 622 127 10% 6 184 340

Standby Power Plant 1 283 884 5% 1 348 078

OHL Grid Power to Plant 6 500 000 0% 6 500 000

Access Roads 2 467 329 20% 2 960 795

Main Camp 4 951 394 5% 5 198 964

Onsite Infrastructure Buildings

etc 2 649 174 15% 3 046 550

Raw Water Supply 1 565 642 10% 1 722 207

Mining Facilities 727 079 10% 799 787

Communications 53 989 5% 56 689

Vehicles 1 564 400 5% 1 642 620

Mobile Plant 2 308 820 10% 2 539 702

Subtotal 29 693 838 31 999 730

Plant Pre-Production

Plant First Fill 899 952 5% 944 950

Spares 3 712 670 5% 3 898 303

Subtotal 4 612 622 4 843 253

Mining Capital Costs

Mining Contractor

Mobilisation 10 372 000 0% 10 372 000

Mining Pre-Strip 4 384 749 0% 4 384 749

Mining Management - Pre-Strip

Period 1 500 000 0% 1 500 000

Subtotal 16 256 749 16 256 749

Other

Insurances 1 511 760 10% 1 662 936

Relocation Cost 5 600 000 7% 6 000 000

Import Tax 1 146 763 0% 1 146 763

Vendor Services 756 553 10% 832 208

Subtotal 9 015 076 9 641 907

Management Costs

Project Management 15 500 000 15% 17 825 000

Owner's Preproduction Costs 11 345 774 5% 11 913 063

Working Capital 5 039 749 10% 5 543 724

Steel, Plate, Mech & Piping

Construction 3 763 419 10% 4 139 761

Electrical Construction 1 757 895 10% 1 933 685

Instrumentation Construction 282 373 10% 310 611

Infrastructure Construction 200 201 10% 220 221

Construction Equipment Hire &

Power 2 518 825 10% 2 770 708

Subtotal 40 408 237 44 656 772

GRAND TOTAL 147 161 567 158 939 551

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21.2.2 Mining Capital Estimate The cost of mining infrastructure has been accounted for under Section 21.2.3.3 below. With

the philosophy of using a mining contractor, the mining equipment has been included as an

operating cost.

Mining capital estimate include the following capital cost items:

• Mobilization costs of the contractor – US$10,372,000;

• Three months of pre-strip and stock-piling prior to the first gold pour – US$4,384,749;

• Mining management cost during pre-strip and stockpiling period – US$1,500,000.

21.2.3 Process Plant and Infrastructure The initial process plant and infrastructure capital costs required for construction are

summarised in the sections below.

21.2.3.1 Basis of Estimate and Assumptions

The capital cost estimate was based on the following technical documents:

• Process plant design information; • General layouts of the process plant; • Process flow diagrams; • Process plant equipment lists; • Piping and Instrument diagrams; • Line lists; • Valve lists; • Instrument lists; • Various discipline material take-offs; • Electrical single line diagrams; • Quotations from vendors on major mechanical process equipment; • EPCM schedule; • In-house historical database.

The following assumptions were made in the preparation of this estimate:

• Executing the work as a single EPCM contract; • Site work being continuous and not constrained by the Owner, war, riots, terrorism or

political interference; • The construction schedule running approximately 19 months; • The chosen site being suitable for shallow foundations and that there are no specific

problems arising from excessive precipitation or ground water, dramatic changes in site geology across the plant area and excessive settlement;

• As a result of the above, no allowances have been made for geotechnical soil improvements of any kind;

• No substantial river crossings were to be encountered for the access road from site to the camp and from the plant to the main road from Hire & Divo. As a consequence only minimum of 5 culverts were allowed for in the road drainage works;

• No blasting operations will be required anywhere during excavations for all earthworks;

• The area does not fall in a seismic zone and an acceleration of 0.02g has been assumed.

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21.2.3.2 Plant Direct Costs

Mechanical Equipment Equipment required was derived from the equipment lists and process flow diagrams. Three

competitive quotes were obtained for major equipment and these were adjudicated on the

technical and commercial aspects. The total cost for mechanical equipment is

US$17,732,002 (SRK Consulting, SENET, Knight Piésold Consulting, 2012 Appendix 21-A).

Earthworks and Civils Earthworks and civil costs were determined by using material take-offs from general

arrangement drawings. Certain items in the bill of quantities (BOQ) were considered as

bought out materials and thus would have to be supplied to the Contractor by the Engineer.

Costs for these items were obtained from the suppliers which were adjudicated commercially

and the resulting costs then populated into the civils and earthworks BOQ’s.

Separate bills of quantities for both the civils and earthworks were prepared based on the

South African National Standards (SANS 1200) in order to obtain rates-based fixed price

“contract values”.

The table below provides a summary of the concrete quantities obtained from the civils

material take-offs for the process plant.

Table 21.3: Summarised Construction Quantities Area Unit Concrete Quantities

Camp m3 1940

Light Industrial Area m3 689

Plant Building & Other Plant Infrastructure m3 1716

Process Plant Structural Concrete m3 4389

Totals m3 8734

The total cost for earthworks is US$1,754,062 and civil works US$9,802,341. Detailed cost

estimate inclusive of BOQ’s are included in the EOS (SRK Consulting, SENET, Knight

Piésold Consulting, 2012 Appendix 21-B).

Structural Steel and Platework Quantities were based on material take-offs (MTO’s) obtained from General Arrangement

Drawings. Quantities from similar jobs were used as a further guide. Unit rates for platework,

tanks and structural steel were obtained from quotes from fabricators. The CIL tanks were

quoted on independently.

The total cost for structural steel and platework is US$5,803,972. The detailed cost estimate

is included in the EOS (SRK Consulting, SENET, Knight Piésold Consulting, 2012 Appendix

21-C).

A summary of the structural steel and platework masses is shown below.

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Table 21.4: Structural Steel and Platework Masses

Description Mass

(Tonnes)

Plate work 256

Tanks 661

Structural Steel 791

Total 1708

Piping and Valves Quantities were based on MTO’s that were obtained from Piping & Instrumentation

Diagrams (P&ID’s) and General Arrangement Drawings (GA). Quantities from similar jobs

were used as a further guide. Fabricators quoted for each material spec (e.g. carbon steel,

stainless steel, HDPE) using the piping quantities and specifications supplied.

A valve list that was compiled from P&ID’s and was submitted to vendors for quotations on

each valve type.

The total cost for piping and valves is US$1,433,928 (SRK Consulting, SENET, Knight

Piésold Consulting, 2012 Appendix 21-D).

Electrical and Instrumentation Electrical equipment quantities were calculated based on PFD’s, P&ID’s, equipment lists,

plant layout drawings and single line diagrams. Quotes were obtained for the design and

supply of the motor control centres (MCC’s), transformers, cable racking, cabling, lighting,

etc.

The PLC and SCADA costs were based on a typical plant configuration with required plant

control from a central control room. The instrumentation costs were based on instrument and

actuated valve lists that were derived from P&ID’s.

Electrical and instrumentation costs are US$5,094,739. The detailed cost estimate is in the

EOS (SRK Consulting, SENET, Knight Piésold Consulting, 2012 Appendix 21-E).

21.2.3.3 Infrastructure costs

Camp and Site Access Roads An allowance has been made for two access roads. One links the plant to the main

provincial road between the towns of Hire and Divo and the other links the process plant and

the main camp. The roads were based on a 5.5m wide gravel carriageway of 9.17km and

2.47km respectively.

The total cost for the access roads is US$2,467,329. The detailed cost estimate is in the

EOS (SRK Consulting, SENET, Knight Piésold Consulting, 2012 Appendix 21-F).

Transport The cost of the total project transport was determined by taking the following into account:

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• The weight/volume restrictions associated with the roads and bridges both in country of manufacture and in Côte d’Ivoire;

• Size, mass and volume of cargo that will be moved. To obtain optimum ratio of weight to volume the cargo would be packed as a mixture of high volume and high weight which will optimize the number of containers to be shipped and hence lower the transport costs;

• Transport costs associated with moving cargo by road from Abidjan Port to site; • Method of containment to be applied in order to reduce the number of containers to

be moved while adhering to the weight restrictions. In coming up with project cargo schedules, it was noted that due to the weight/volume ratio

the most cost effective form of transport would be to combine the transportation of civils with

mechanical equipment that has a high volume ratio. The same concept applies for the

balance of disciplines.

The majority of the cargo will be shipped in 40 foot general purpose containers and 40 foot

open top containers. The mills and the crusher will require breakbulk shipment from country

of manufacture. A charter vessel will be required and indicative global freight rates have

been used for the purpose of the EOS.

Typical shipping costs per container type are given in table below:

Table 21.5: Typical Shipping Costs of Container Types

JHB-FCA

Durban

(US$)

Durban-

Abidjan

(US$)

Port Clearance &

Handling (US$)

Abidjan to

Site (US$)

Abidjan

Customs

(US$)

Total

(US$)

20 foot 1,617.06 1,614.31 2,193.37 1,150.80 179.47 6,755.01

40 foot 1,915.89 2,784.31 2,879.74 1,986.50 179.47 9,745.91

The total project transport costs amounted to US$5,554,000.

An estimate by type of equipment moved is included in the EOS (SRK Consulting, SENET,

Knight Piésold Consulting, 2012 Appendix 21-G).

Tailings Capital Costs

The capital cost estimates included in this report have been prepared on the following basis:

Phase 1

• Construction of two earthfill starter walls to crest level 249.5 mamsl;

• Construction of a permanent return water system;

• Construction of filters on the upstream face of the earthfill embankment, and

installation of housing pipes for submersible pumps;

• Excavation of diversion channels to avoid rain water from entering the TSF;

• Installation of a filter drain located within the embankment for removing seepage

through the dams;

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• Supply and installation of toe drains to remove seepage collected in the filter drains;

• Supply and installation of collection sumps downstream of TSF embankments;

• Supply and installation of sump pumps and piping to discharge back into TSF;

• Supply and installation of tailings distribution pipeline with spigot offtakes around the

TSF;

• TSF perimeter roads and fencing.

The start-up cost for the TSF is US$5,622,127. This includes an allowance for contractors

P&Gs, Engineering and Supervision fees. The cost breakdown is summarised in the table

below. The detailed cost estimate is in the EOS (SRK Consulting, SENET, Knight Piésold

Consulting, 2012 Appendix 21-H).

Table 21.6 : Cost Breakdown Summary

Section Description Amount (US$)

3 Main Impoundment 3,000,159.50

4 Underdrain System Sump Sections & Filter Drain 430,498.00

5 Slurry Distribution Pipework 573,355.75

SUB TOTAL MEASURED ITEMS 4,004,013.25

1 Preliminary & General 1,119,700.00

2 Engineering and Design Costs 498,414.12

Total 5,622,127.37

Water Supply Knight Piésold was asked to identify potential locations to store water to be used during plant

start-up when the tailings pond cannot provide return water back to the plant. The water

storage dams will also be used to store dewatering from the open pit.

The coast of drilling, pump installation and aquifer testing for two water supply boreholes at

the Agbaou camp is given in the table below. This includes the cost of drilling two water

supply boreholes and equipping them with suitable submersible electric borehole pumps.

Table 21.7: Cost of Drilling, Pump Installation and Aquifer Testing at the Agbaou Camp Item Sub-Total, US$

Drilling and aquifer testing 40,000

Supply and install submersible pumps 11,300

Total 51,300

For abstraction from the water supply dam, allowance was made for two pumps mounted on

pontoons, an intermediate pumping station and the 4.3km pipeline to the plant. In addition to

this an access road and an overhead line was included in the costing.

A summary of the water supply costs is included in the table below.

The cost breakdown of water supply from the preferred dam location is in the EOS (SRK

Consulting, SENET, Knight Piésold Consulting, 2012 Appendix 21-I).

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Table 21.8: Water Supply and Storage Capital Cost Summary DESCRIPTION - WATER SUPPLY COST (US$)

Pipe Line 506,646

Pumps and Tank 161,684

OHL, Transformer and Switchgear Cost 314,933

Access Road Cost 148,120

Water Storage Dam 434,257

Total 1,565,642

Power Supply Power will be provided from the Côte d’Ivoire national grid. The cost of the overhead line

installation was provided by ENDEAVOUR, and is US$6,500,000 and is inclusive of

contingency.

Two off 1.5MVA diesel generators have been allowed for as standby power in the cost

estimate this at a cost of US$1,283,884.

Plant Administration Buildings and Warehousing Buildings located in the plant area will consist of security offices, change house, plant

messing, process plant equipment workshop, control room, general administration offices,

medical facility, substation, assay laboratory, reagent warehouses and spares warehouse.

A total of 3 warehousing facilities for storage of general spares, cyanide, reagents and

consumables will be provided.

The total cost for the above in plant infrastructure including all civil costs amounts to

US$2,649,174. The detailed cost estimate is attached in the EOS (SRK Consulting, SENET,

Knight Piésold Consulting, 2012 Appendix 21-J).

Main Camp The main camp was sized to accommodate 200 people. Quotes were obtained from

suppliers who have provided similar camp materials in Africa for prefabricated type housing.

The capital costs of the main camp include:

• Six (6), 2 bed roomed manager’s houses; • Eighteen (18), 10 bed housing units each with its own bathroom; • A laundry unit; • A guard-house; • A centrally located camp kitchen & dining room; • An entertainment/bar area; • Potable water, sewage disposal units and piping reticulation; • Associated civils; • All furniture, catering and laundry equipment.

The capital cost for the main camp is US$4,951,394 and includes civils and earthworks. The

detailed estimate for the main camp is in the EOS (SRK Consulting, SENET, Knight Piésold

Consulting, 2012 Appendix 21-K).

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There is no capital expenditure associated with accommodation in nearby villages.

Mining Infrastructure The following will be allowed for in the mining facilities:

• Main workshop and repair facilities, to maintain the fleet of haul trucks and major mining equipment, complete with admin buildings housed within;

• Tyre repair and replacement bay; • Mining equipment re-fuelling centre; • Explosive storage, which will be located away from the main facilities; • Change house and security office; • Messing facility for the mining personnel.

Capital cost for mining infrastructure is US$727,079. The breakdown for mining

infrastructure is in the EOS (SRK Consulting, SENET, Knight Piésold Consulting, 2012

Appendix 21-L).

Communications, Security, Networking and Radios Capital costs for the communication system, networking (VSAT) and radios were allowed for.

The CCTV cost is included in the instrumentation costs.

The capital cost for communications is US$53,989.

Vehicles and Mobile Equipment A vehicle and mobile equipment list was drawn up and was submitted to suppliers for

quotations that were then used in the estimate. The capital cost for vehicles is

US$1,564,400 and US$2,308,820 for mobile plant. The breakdown for vehicles and mobile

plant equipment are in the EOS (SRK Consulting, SENET, Knight Piésold Consulting, 2012

Appendix 21-M).

21.2.3.4 Plant Pre-Production Costs

First Fills The first fill costs were defined as those costs incurred prior to commissioning in preparing

the circuit to accept ore. These costs included the initial charge of steel balls, at various

sizes to the SAG and Ball Mills, carbon in the CIL circuit and thermic oil for the elution

heaters.

The quantities were developed from first principles. The capital cost for first fills is

US$899,952. Detail regarding the build-up of the first fill costs, is in the EOS (SRK

Consulting, SENET, Knight Piésold Consulting, 2012 Appendix 21-N).

Spares Spares costs cover commissioning, strategic and two year spares. Spares quantities and

costs were supplied by the respective vendors.

The capital cost of spares required is US$3,712,670. Detailed costs are in the EOS (SRK

Consulting, SENET, Knight Piésold Consulting, 2012 Appendix 21-O).

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21.2.3.5 Other Costs

Insurances Insurances were estimated to cover contractor’s all risk, public liability, professional

indemnity, marine insurance, motor vehicle third party insurance and employee’s liability.

The insurance capital cost is US$1,511,760.

Relocation An allowance of US$5,600,000 was made and covers relocation of local farmers and

villagers in the vicinity of the mine.

Import Taxes An import tax (regional statistical tax and two community / solidarity taxes) of 2.5% applies to

all imported items. This cost is US$1,146,763.

Vendor Services The cost for vendor services includes items where the equipment is of sufficient complexity

to require the presence of the vendor in order for equipment to be installed. The costs are

based on quotes obtained from the respective vendors. The capital cost for vendor services

is US$756,553. The detailed cost breakdown of vendor services cost is in the EOS (SRK

Consulting, SENET, Knight Piésold Consulting, 2012 Appendix 21-P).

21.2.3.6 Management Costs

Engineering, Procurement and Construction Management (EPCM) EPCM cost was estimated as 10% of total project costs. This total will cover design,

procurement, project management, construction management and disbursement costs. The

estimate for EPCM is US$15,500,000.

Construction Labour Structural, mechanical, platework, piping, electrical & instrumentation installation and

erection will be performed by contractors. Civils and earthworks construction labour costs

including P&G’s are contained in the civil and earthworks capital costs. Pricing was based on

rates based quotations where all work and P&G’s (including their own camp) was supplied

as part of the pricing. The independent contractors would work under the supervision of the

EPCM contractor.

The cost for construction labour is US$6,003,888. The detailed construction labour costs are

in the EOS (SRK Consulting, SENET, Knight Piésold Consulting, 2012 Appendix 21-Q).

Construction Equipment and Associated Costs Construction equipment and its associated costs include construction power, construction

fuel, vehicles, rental of crane and lifting equipment hire. The cost is US$2,518,825 (SRK

Consulting, SENET, Knight Piésold Consulting, 2012 Appendix 21-R).

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Owner’s Preproduction Costs Owner’s preproduction cost is the cost incurred prior to the commencement of the project.

This cost includes owner’s construction team costs, all labour costs inclusive of salaries,

recruiting fees, accommodation and messing, training, assay and offsite offices.

This also allows for the gradual ramp-up of the production staff during the construction

phase in preparation for plant production.

This cost excludes study and exploration costs. The owner’s cost is US$11,345,774 (SRK

Consulting, SENET, Knight Piésold Consulting, 2012 Appendix 21-S).

Working Capital Working capital is the costs incurred from commissioning until the revenue from gold sales

can pay for the total operating costs. The operating costs include mining, plant, G&A, assay

and refining. Certain assumptions were made i.e., tonnage ramp up, gold inventory ramp-

up, bullion shipment, price of gold, reagents and consumables stockholding period.

Using these assumptions, the calculated working capital is US$5,039,749.

21.2.3.7 Contingency

An effective contingency of 8% has been included to cover items included in the scope of

work, but cannot be adequately defined at this time due to lack of accurate detailed design

information. This cost is US$11,777,985.

21.3 SUSTAINING, REHABILITATION AND CLOSURE COSTS

21.3.1 Mining Rehabilitation, Mining Contractor Demobilization and Closure Costs

The basis of the costing is the assumption that mine equipment (as opposed to third party

equipment) will be utilised to undertake the earth moving activities required to place the

cover material. Therefore the plant rates that are used in the liability assessment are those

derived by SRK during the feasibility assessment for the Project.

No local information on vegetation, fertiliser and mulch costs was available at the time of

preparing this assessment. Therefore the rates that have been used in this assessment are

an average of rates that SRK have collected in other African states (excluding South Africa).

A provision has been made to undertake post closure management and monitoring for a

period of three years after closure on all of the facilities. This is required to repair any erosion

that may occur as vegetation becomes established as well as to monitor the effectiveness of

the rehabilitation measures.

An amount of US$249,300 has been allowed for demobilization of the mining contractor.

The closure costs for the three waste rock dumps (WRD) are presented in the table below.

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Table 21.9: Mining Rehabilitation and Closure Costs DESCRIPTION OF PROPOSED WORK TO BE UNDERTAKEN ON WRD 1, 2 AN 3:

• Load and haul soils

• Place soils

• Mulch, fertilise and vegetate

WASTE ROCK DUMP 1 – BILL OF QUANTITIES

Item

No Unit

Rate

(US$/unit) Quantity

Total

(US$)

1 Load and haul soil m3 1.18 196,000 231,280

2 Placement of cover material m3 0.86 196,000 168,560

3 Fertilise and mulch Ha 4,105 65 266,825

4 Seed Ha 900 65 58,500

Maintenance

4 Maintain vegetated areas m2 0.05 650,000 32,500

CAPITAL COST TO REHABILITATE (carried to summary) 725,165

MAINTENANCE COSTS 32,500

TOTAL COST 757,665


Item

No Unit

Rate

(US$/unit) Quantity

Total

(US$)




4 Seed Ha 900 35 31,500

Maintenance




TOTAL COST 406,875


Item

No Unit

Rate

(US$/unit) Quantity

Total

(US$)




4 Seed Ha 900 42 37,800

Maintenance




TOTAL COST 391,610

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21.3.2 Tailings Sustaining, Rehabilitation and Closure Costs

21.3.2.1 Sustaining Capital

Sustaining capital is US$10,837,622 and includes the following TMF costs in years 3 and 6:

• Main impoundment; • Under drain system sump sections and filter drain; • Slurry distribution pipework; • P&G’s; • Engineering and design costs.

21.3.2.2 Rehabilitation and Closure Costs

Rehabilitation and closure works that will take place concurrently with the construction of the

facility will include stripping and stockpiling of topsoil from the site for use in the rehabilitation

and closure process. Being a downstream constructed rockfill embankment, no rehabilitation

work will be able to be carried out on the outer slopes of the facility until the Phase 3 wall

raise has been completed. The overall outer slope will be 1V:3H. The slope surface will be

top soiled and vegetated. It is expected that decommissioning, rehabilitation and closure of

the facility at the cessation of operations will include the following:

• Removal of the slurry delivery and return water pumps and pipelines and all associated works;

• Ridge ploughing for dust control; • Construction of dividing paddocks into 100 x 100m grids on the top surface of the

TSF, so as to hold rainwater evenly over the entire footprint of the TSF, thereby enhancing water evaporation and reducing the volume of water collecting in the central pool zone.

Grassing of the area will lead to a significant reduction in surface erosion.

Monitoring of surface and groundwater quality in the area is likely to be required to continue

for a period of up to 30 years. This will be achieved by installing collection chambers and

sampling points in order to monitor water quality both for underground seepage and surface

water drainage.

Annual rehabilitation costs and closure costs have been estimated and are shown in Table

21.10 below.

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Table 21.10: Rehabilitation and Closures Costs

Year no.

Concurrent Closure

Costs

(US$)

Total

(US$) Rehabilitation

(US$)

0

1 140,625 140,625

2 140,625 140,625

3 140,625 140,625

4 140,625 140,625

5 140,625 140,625

6 140,625 140,625

7 140,625 140,625

8 70,313 70,313

9 4,855,305 4,855,305

Total

(US$) 1,054,688 4,855,305 5,909,993

21.3.3 General Sustaining Capital An allowance of US$250,000 per year has been included to cover minor sustaining capital

items.

21.4 OPERATING COST ESTIMATE

21.4.1 Introduction The annual operating costs for the life of the mine (LOM) were estimated for mining,

processing, general and administration (G&A), royalties and refining charges and are

summarized in the table below and described in detail in the sections that follow.

The operating costs (which are exclusive of duties and taxation) were developed by:

• Mining – SRK; • Process plant – SENET; • Tailings storage facility – Knight Piésold; • G&A – SENET and ENDEAVOUR; • Assay laboratory – SENET (SGS).

A summary of the overall operating costs are presented in the table below.

Table 21.11: Overall Operating LOM Costs

Mining Operating Costs In conjunction with ENDEAVOUR a mining contractor was selected. The annual operating

costs estimated by this contractor as well as mining management costs supplied by SRK are

shown in Table 21.12.

Unit Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 LOM

Mining US$/t 21.79 27.33 29.87 29.70 48.18 35.60 28.66 29.54 30.94

Plant US$/t 9.26 9.37 9.35 9.59 9.76 9.61 9.69 9.61 9.51

Tailings Storage Facility (TSF) US$/t 0.14 0.15 0.15 0.17 0.15 0.15 0.15 0.15 0.15

G&A US$/t 4.91 5.19 4.97 5.62 6.16 5.72 5.83 5.76 5.46

Assay US$/t 1.01 1.06 1.06 1.20 1.31 0.79 0.81 1.55 1.06

Refining US$/t 0.26 0.35 0.43 0.45 0.32 0.39 0.37 0.48 0.37

Government Royalties US$/t 1.97 2.58 3.22 3.36 2.38 2.88 2.75 3.58 2.77

Total Operating Costs US$/t 39.33 46.03 49.05 50.08 68.27 55.14 48.25 50.67 50.26For

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Table 21.12: Overall Mining LOM Operating Cost

Year Haul & Load

(US$)

Drilling &

Blasting (US$)

Management

Contractor (US$)

Rehandling

(US$)

Mine Management

ENDEAVOUR (US$) Sum (US$)

1 25 378 055 0 8 855 840 186 284 3 322 000 37 742 179

2 30 954 209 0 9 888 969 542 062 3 322 000 44 707 240

3 35 267 273 685 116 9 302 246 509 901 3 334 000 49 098 536

4 31 145 298 1 118 825 7 526 010 345 949 3 059 000 43 195 082

5 40 047 726 9 576 723 10 978 453 442 197 2 874 000 63 919 099

6 30 558 918 6 879 607 7 449 684 216 593 2 906 000 48 010 804

7 22 054 292 7 306 678 5 865 157 239 341 2 459 000 37 924 467

8 9 042 510 3 475 961 3 529 388 149 460 1 835 000 18 032 319

Total (US$) 224 448 281 29 042 911 63 395 747 2 631 787 23 111 000 342 629 725

The mine management costs in the table above include ENDEAVOUR management costs,

administration costs estimated to be US$0.25/t ore and grade control costs estimated to be

US$1.2/t ore.

21.4.2 Processing Plant Operating Costs

21.4.2.1 Summary

The annual process plant operating costs for the life of the mine of US$9.66/t of ore

processed for the Agbaou Project are presented in the table below.

Table 21.13: Overall Process Plant LOM Operating Costs

Unit Yr. 1 Yr. 2 Yr. 3 Yr. 4 Yr. 5 Yr. 6 Yr. 7 Yr. 8 LOM Plant & Maintenance Labour

US$/t 1.43 1.51 1.51 1.70 1.87 1.84 1.87 1.85 1.67

Consumables & Reagents US$/t 4.83 4.83 4.80 4.64 4.42 3.92 4.05 3.84 4.49

Maintenance Supplies US$/t 0.51 0.54 0.54 0.61 0.67 0.66 0.67 0.66 0.60

Power US$/t 2.48 2.49 2.51 2.63 2.81 3.20 3.09 3.26 2.75

TSF US$/t 0.14 0.15 0.15 0.17 0.15 0.15 0.15 0.15 0.15

Total Process Plant Costs

US$/t 9.40 9.52 9.50 9.75 9.91 9.76 9.84 9.76 9.66

21.4.2.2 Basis of Estimate

The plant operating costs were compiled from a variety of sources, namely:

• ENDEAVOUR’s database; • First principles; • Supplier quotations for reagents and consumables; • SENET’s experience on similar sized plants.

The major cost centres of the plant that were covered in this estimate were:

• Process plant operating labour; • Plant maintenance labour; • Reagents and consumables; • Maintenance supplies; • Power.

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21.4.2.4 Plant Operating and Maintenance Labour

The combined plant and maintenance labour cost, is estimated to be US$1.67/t for the life of

the mine. These costs are inclusive of:

• Plant operating labour; • Plant maintenance labour; • Expatriate labour travel and accommodation; • Safety supplies; • Training.

A detailed cost breakdown of each cost associated with plant operating and maintenance

labour are in the EOS (SRK Consulting, SENET, Knight Piésold Consulting, 2012 Appendix

21-T).

The plant operating and maintenance labour costs were derived from first principles where

the actual labour complement was drawn up in conjunction with ENDEAVOUR. The

expatriate labour is broken down into two categories; Expat 1 and Expat 2. It was assumed

that Expat 1 are the expats from outside West Africa and Expat 2 are from West Africa.

The labour schedule has been drawn up assuming a roster of 8 weeks on and 3 weeks off

for the expatriates and three by 8 hour shifts for the local labour.

The labour schedule was prepared with estimated costs associated with each position

including the employer remittance to the government of 12% for expatriate labour and 2.8%

for local labour (SRK Consulting, SENET, Knight Piésold Consulting, 2012 Appendix 21-T).

The travel and accommodation costs for expatriates were determined by using the roster

mentioned above, number of expatriates, the number of flights per annum, costs per flight

and food costs.

The annual cost of the plant personnel’s safety supplies was determined using a safety gear

list and its costs and the number of employees in the plant.

An allowance of 2% of the total plant salaries has been allocated to cater for any external

training consultants that might be hired for ongoing training of labour.

21.4.2.5 Process Plant Reagents and Consumables

The estimated LOM plant reagents and consumable costs equated to US$4.49/t. A

combination of projected reagent consumptions and unit costs were used to obtain these

costs. The costs were then weighted based on the combination of types of ores being

treated and are summarized in the table below (SRK Consulting, SENET, Knight Piésold

Consulting, 2012 Appendix 21-U).

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Table 21.14: LOM Reagents and Consumables Costs

Unit Cost

(US$/unit)

LOM Weighted

Consumption

(kg/t)

LOM Cost

(US$)

LOM Cost

(US$/t)

Tonnage Processed - - - 11 074 927

Crusher Liners set 14 202 - 793 333 0.072

Mill Liners set 687 542 - 5 468 444 0.494

SAG Mill Grinding

Media t 1 388 0.33 5 146 308 0.465

Ball Mill Grinding Media t 1 165 0.40 5 161 131 0.466

Cyanide t 2 282 0.35 8 810 706 0.796

Lime t 330 4.24 15 517 144 1.401

Caustic t 845 0.13 1 233 972 0.111

Hydrochloric Acid (33 %

HCl) t 476 0.10 528 416 0.048

Carbon t 3 592 0.03 994 593 0.090

Sodium Metabisulphite t 617 0.59 4 014 285 0.362

Copper Sulphate t 2 726 0.07 1 964 907 0.177

Hydrogen Peroxide t 717 0.01 114 356 0.010

Smelting Reagents t 871 0.001 11 913 0.001

Total 49 759 508 4.49

21.4.2.6 Plant Maintenance Costs and Supplies

Maintenance costs refer to the costs of operating spares and lubricants for the plant. It has

been assumed that the plant will experience a moderate amount of wear. A provision has

been made to allow maintenance costs at an annual cost of 5% of the mechanical

equipment cost. This equates to a LOM cost of US$0.60/t.

21.4.2.7 Power

The power usage for the LOM calculated to 380,530,572kWh and at a power cost of

US$0.08/kWh, a LOM power cost of US$2.75/t was determined. A summary of the LOM

power costs are shown in the table below (SRK Consulting, SENET, Knight Piésold

Consulting, 2012 Appendix 21-V).

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Table 21.15: LOM Power Costs Unit Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 LOM

Tonnage Processed

Saprolite

Tonnage tpa 1,732,010 1,623,809 1,569,528 1,073,867 650,225 18,437 282,685 0 6,950,560

Transition

Tonnage tpa 0 11,936 62,069 209,236 261,705 187,026 12,366 0 744,338

Bedrock

Tonnage tpa 0 0 12,406 171,127 414,639 1,143,225 1,028,285 610,347 3,380,029

Combined

Tonnage

Processed

tpa 1,732,010 1,635,744 1,644,002 1,454,229 1,326,569 1,348,689 1,323,336 610,347 11,074,927

Variable Power Usage

Saprolite

Tonnage kWh/t 18.16 18.16 18.16 18.16 18.16 18.16 18.16 18.16 18.16

Transition

Tonnage kWh/t 21.73 21.73 21.73 21.73 21.73 21.73 21.73 21.73 21.73

Bedrock

Tonnage kWh/t 25.30 25.30 25.30 25.30 25.30 25.30 25.30 25.30 25.30

Weighted

Power kWh/t 18.16 18.19 18.35 19.51 21.10 24.71 23.74 25.30 20.58

Total Variable

Power kWh/a 31,453,305 29,747,729 30,165,244 28,377,622 27,985,308 33,322,498 31,417,880 15,441,789 227,911,374

Fixed Power Usage

Saprolite

Tonnage kWh/t 12.90 12.90 12.90 12.90 12.90 12.90 12.90 12.90 12.90

Transition

Tonnage kWh/t 14.20 14.20 14.20 14.20 14.20 14.20 14.20 14.20 14.20

Bedrock

Tonnage kWh/t 15.50 15.50 15.50 15.50 15.50 15.50 15.50 15.50 15.50

Weighted

Power kWh/t 12.90 12.91 12.97 13.39 13.97 15.28 14.93 15.50 13.78

Total Fixed

Power kWh/a 22,344,667 21,118,236 21,322,056 19,476,909 18,530,288 20,610,277 19,758,062 9,458,687 152,619,180

Total Power Usage

Total Power kWh/t 31.06 31.10 31.32 32.91 35.06 39.99 38.67 40.80 34.36

Total Power kWh/a 53,797,971 50,865,964 51,487,300 47,854,531 46,515,596 53,932,775 51,175,941 24,900,476 380,530,554

Power Costs

Grid Power

Costs US$/kWh 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08

Annual Power

Costs US$/a 4,303,838 4,069,277 4,118,984 3,828,362 3,721,248 4,314,622 4,094,075 1,992,038 30,442,444

Power Costs US$/t 2.48 2.49 2.51 2.63 2.81 3.20 3.09 3.26 2.75

21.4.2.8 Tailings Operating Costs

Current operating costs on similar sized tailings dams in South Africa are approximately

US$0.15/t of solids deposited. The estimate of operating costs allows for the operation of

the facility by a recognised tailings facility operator on a rate per ton basis and will include for

management, supervision, labour, equipment and materials required for:

• Operation of the slurry delivery system; • Operation of the pump barge return water system; • Routine inspections and maintenance of the facility and associated infrastructure,

including general housekeeping functions such as grass cutting, changing spigot discharge locations, etc.

The table below shows the estimated LOM tailing operating costs.

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Table 21.16: LOM Tailings Operating Costs Year no. Operating Operating

Cost (US$) Costs (US$/t)

1 241,500 0.14

2 241,500 0.15

3 241,500 0.15

4 241,500 0.17

5 201,000 0.15

6 201,000 0.15

7 201,000 0.15

8 201,000 0.15

Ave LOM 221,250 0.15

21.4.3 General and Administration Costs G&A LOM costs were determined to be US$5.46/t, with input from SENET and

ENDEAVOUR. The major G&A cost items that were included are as follows:

• Labour (includes salaries, security and training & recruiting); • Camp food & local messing; • Maintenance; • Offsite offices and travel; • Environmental; • Other costs.

A summary of the G&A costs is shown in the table and sub-sections below and details of

each cost item are included in the EOS (SRK Consulting, SENET, Knight Piésold

Consulting, 2012 Appendix 21-W).

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Table 21.17: LOM G&A Costs

21.4.3.1 General and Administration Labour Costs

The G&A labour costs were derived from first principles where the actual labour complement

was drawn up in conjunction with ENDEAVOUR. The expatriate labour is broken down into

two categories namely Expat 1 and Expat 2. It was assumed that Expat 1 are the expats

from outside West Africa and Expat 2 are from West Africa.

The labour schedule has been drawn up assuming a roster of 8 weeks on and 3 weeks off

for the expatriates and three by 8 hour shifts for the local labour.

A labour schedule was prepared with estimated costs associated with each position

including the employer remittance to the government of 12% for expatriate labour and 2.8%

for local labour (SRK Consulting, SENET, Knight Piésold Consulting, 2012 Appendix 21-W).

An allowance of 2% of the total G&A salaries has been allocated to cater for any external

Item Unit Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 LOM

Salaries & Wages

General & Administration Salaries Expat US$pa 1 799 680 1 799 680 1 669 319 1 669 319 1 669 319 1 669 319 1 669 319 760 188 12 706 143

General & Administration Salaries Local US$pa 857 719 857 719 857 719 857 719 857 719 857 719 857 719 390 595 6 394 626

General Maintenance Salaries Expat US$pa 708 729 708 729 515 498 515 498 515 498 515 498 515 498 234 751 4 229 697

General Maintenance Salaries Local US$pa 744 038 744 038 744 038 744 038 744 038 744 038 744 038 338 826 5 547 089

Total Salaries & Wages US$pa 4 110 165 4 110 165 3 786 573 3 786 573 3 786 573 3 786 573 3 786 573 1 724 360 28 877 555

Security

Staff US$pa 217 135 217 135 217 135 217 135 217 135 217 135 217 135 98 880 1 618 823

Vehicles US$pa 149 748 149 748 149 748 149 748 149 748 149 748 149 748 68 193 1 116 429

Total Security US$pa 366 883 366 883 366 883 366 883 366 883 366 883 366 883 167 074 2 735 252

Training and Recruiting

Training US$pa 82 203 82 203 82 203 82 203 82 203 82 203 82 203 37 434 612 858

Recruiting US$pa 36 033 36 033 36 033 36 033 36 033 36 033 36 033 16 409 268 642

Total Training & Recruiting US$pa 118 237 118 237 118 237 118 237 118 237 118 237 118 237 53 844 881 500

Camp Food and Local Messing

Camp Food US$pa 508 638 508 638 508 638 508 638 508 638 508 638 508 638 231 628 3 792 096

Local Messing US$pa 310 004 310 004 310 004 310 004 310 004 310 004 310 004 141 172 2 311 200

Total Camp Food & Local Messing US$pa 818 642 818 642 818 642 818 642 818 642 818 642 818 642 372 800 6 103 296

Maintenance

Main Camp and In-plant Roads US$pa 200 000 200 000 200 000 200 000 200 000 200 000 200 000 91 078 1 491 078

Infrastructure Plant Buildings US$pa 200 000 200 000 200 000 200 000 200 000 200 000 200 000 91 078 1 491 078

Local Village US$pa 100 000 100 000 100 000 100 000 100 000 100 000 100 000 45 539 745 539

Vehicle Running Costs US$pa 489 290 489 290 489 290 489 290 489 290 489 290 489 290 222 817 3 647 845

Total Maintenance US$pa 989 290 989 290 989 290 989 290 989 290 989 290 989 290 450 511 7 375 539

Offsite Offices & Travel

Abidjan Off ice US$pa 80 000 80 000 80 000 80 000 80 000 80 000 80 000 36 431 596 431

Expatriates Flight Costs - Admin US$pa 278 348 278 348 278 348 278 348 278 348 278 348 278 348 126 756 2 075 192

Additional camp visitors cost US$pa 10 770 10 770 10 770 10 770 10 770 10 770 10 770 4 905 80 295

Total Offsite Office & Travel US$pa 369 118 369 118 369 118 369 118 369 118 369 118 369 118 168 092 2 751 918

Environmental

Consultants US$pa 150 000 150 000 150 000 150 000 150 000 150 000 150 000 68 308 1 118 308

Consumables US$pa 50 000 50 000 50 000 50 000 50 000 50 000 50 000 22 769 372 769

Field Supplies US$pa 20 000 20 000 20 000 20 000 20 000 20 000 20 000 9 108 149 108

Waste Disposal US$pa 100 000 100 000 100 000 100 000 100 000 100 000 100 000 45 539 745 539

Total Environmental US$pa 320 000 320 000 320 000 320 000 320 000 320 000 320 000 145 724 2 385 724

Other Admin Costs

Amortisation of Fuel Farm US$pa 456 000 456 000 456 000 456 000 456 000 0 0 0 2 280 000

Office Consumables US$pa 60 000 60 000 60 000 60 000 60 000 60 000 60 000 27 323 447 323

Community Affairs US$pa 100 000 100 000 100 000 100 000 100 000 100 000 100 000 45 539 745 539

Insurances US$pa 600 000 600 000 600 000 600 000 600 000 600 000 600 000 273 233 4 473 233

License Fees for Softw are US$pa 20 000 20 000 20 000 20 000 20 000 20 000 20 000 9 108 149 108

Computer Hardw are Update US$pa 20 000 20 000 20 000 20 000 20 000 20 000 20 000 9 108 149 108

Annual Surface Fee US$pa 47 369 47 369 47 369 47 369 47 369 47 369 47 369 21 571 353 154

Accounting Tax Audit & Legal US$pa 100 000 100 000 100 000 100 000 100 000 100 000 100 000 45 539 745 539

Total Other Admin Costs US$pa 1 403 369 1 403 369 1 403 369 1 403 369 1 403 369 947 369 947 369 431 420 9 343 003

Total G &A US$pa 8 495 704 8 495 704 8 172 111 8 172 111 8 172 111 7 716 111 7 716 111 3 513 824 60 453 788

Tonnage Treated tpa 1 732 010 1 635 744 1 644 002 1 454 229 1 326 569 1 348 689 1 323 336 610 347 11 074 927

G & A Cost US$/t 4.91 5.19 4.97 5.62 6.16 5.72 5.83 5.76 5.46

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training consultants that might be hired for ongoing training of labour. Recruitment cost was

determined by assuming a 10% staff turnover and a recruitment fee of 12.5% of annual

salary.

21.4.3.2 Camp Food and Local Messing

All expatriates will be expected to be on an 11 week roster (8 weeks on and 3 weeks off

duty). During the 8-week period on duty, the expatriates will be accommodated in the main

camp where food, cleaning and laundering services will be provided at a cost of US$27.00

per person per day.

A detailed roster for the LOM was developed by taking into account the number of days on

and off for each expat from which the estimated camp food and catering costs were

developed.

Local labour will be expected to be at work for about 240 days per annum. Catering costs of

US$5.00/person/day for local staff on was used together with the number of local staff on

duty and the days per annum to determine the messing costs.

21.4.3.3 Maintenance

G&A maintenance costs will cover the following main items:

• Main camp and in-plant roads; • Infrastructure plant buildings; • Local village; • Vehicle running costs.

Allowances were made for all the costs except the vehicle running costs. This was estimated

by taking into account the running costs per hour per vehicle and the estimated annual

running time for each type of vehicle (SRK Consulting, SENET, Knight Piésold Consulting,

2012 Appendix 21-W).

21.4.3.4 Offsite Offices & Travel

A provision made to cover the Abidjan office costs and additional camp visitors. The travel

costs for expatriates were determined by using the expat labour roster (8 weeks on and 3

weeks off), number of expatriates, the number of flights per annum, costs per flight and food

costs.

21.4.3.5 Environmental

Estimated annual allowances were made for the following environmental cost items:

• Consultants; • Consumables; • Field Supplies; • Waste Disposal.

21.4.3.6 Other Administration Costs

Estimated annual allowances were made for the following other admin costs:

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• Communication; • Insurances; • Licence fees for software; • Computer hardware update; • Accounting tax audit and legal; • Annual surface fee.

21.4.4 Assay Laboratory Costs The LOM assay costs summarized in table below are based on the assumption that:

• SGS will provide equipment and manage the on-site laboratory for the first 5 years during which assay costs will be paid to SGS;

• ENDEAVOUR will own and operate the equipment and laboratory modules after Year 5.

Table 21.18: LOM Assay Cost Summary

During the first 5 years SGS will provide a total compliment of lab staff that will run and

manage the laboratory and all costs associated with this will be part of SGS fixed costs.

From year 6 onwards, the laboratory will be run by the mine operations and ENDEAVOUR

will take over the staff compliment from SGS.

A schedule of the LOM flights for each expat was developed along the same principles as

process plant expatriate schedule and will be allowed for from Year 5. Travel costs

associated with expatriates during years 1 to 5 from country of origin to site have been

assumed to be included in the SGS quote.

The annual cost of the plant personnel’s safety supplies was determined using a safety gear

list and its costs and the number of employees in the plant and varied from year 5 onward.

Training has been allowed at 2% of the total labour costs from Year 5.

SGS will be paid all assay costs during the first 5 years of operation in order to enable them

to supply the laboratory equipment and modules. They will manage a laboratory capable of

assaying 26,240 sample equivalents per month.

From Year 6 ENDEAVOUR will operate the laboratory and the following will be the

associated costs:

• Labour costs; • Safety supplies; • Training; • Expatriate flights; • Lab assay costs;

Units Years 1 Years 2 Years 3 Years 4 Years 5 Year 6 Year 7 Year 8 LOMLabour Costs US$ $215 627 $215 627 $125 722 $556 977

Flights and Accommodation US$ $22 948 $22 948 $9 057 $54 954

Safety US$ $1 456 $1 456 $1 456 $1 456 $1 456 $7 642 $7 642 $3 480 $26 041

Training US$ $0 $0 $0 $0 $0 $4 313 $4 313 $1 145 $9 770

Fixed Costs US$ $944 184 $944 184 $944 184 $944 184 $944 184 $0 $0 $0 $4 720 920

Variable Costs US$ $795 055 $795 055 $795 055 $795 055 $795 055 $795 055 $795 055 $795 055 $6 360 439

Spares US$ $0 $0 $0 $0 $0 $20 365 $20 365 $9 274 $50 004

Total Lab Costs US$ $1 740 694 $1 740 694 $1 740 694 $1 740 694 $1 740 694 $1 065 950 $1 065 950 $943 733 $11 779 104Lab Costs US$/t $1.01 $1.06 $1.06 $1.20 $1.31 $0.79 $0.81 $1.55 $1.06

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• Spares.

21.4.5 Refining & Royalties The refining process will be completed within 3 working days after receipt of bullion. The

refiner will charge “refining costs” to cover for melting, assaying and refining.

Generally 99.8% to 99.95% (for the purposes of this study this was assumed to be 99.90%)

of the gold contained in the bullion will be payable to Agbaou and the sale price will be fixed

on the day of the refinery outrun with a settlement of 2 business days.

Bullion transport, insurance and refining costs used were assumed at US$5/oz.

The Mining Code provides for a 3% royalty payable to the State of Côte d’Ivoire on all gold

production. The royalty is calculated less refining and bullion transportation costs. The cost

breakdown of refining and royalties is shown in the table below.

Table 21.19: Refining and Royalty Cost Summary Unit Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 LOM

Combined Production Combined Ore Processed (t) 1 732 010 1 635 744 1 644 002 1 454 229 1 326 569 1 348 689 1 323 336 610 347 11 074 927

Combined Plant Feed Grade (g/t) 1.76 2.31 2.89 3.02 2.14 2.62 2.50 3.27 2.50

Combined Recovery % 93.17% 93.17% 93.13% 92.84% 92.44% 91.54% 91.78% 91.39% 92.57%

Contained Gold (kg) 3 056 3 774 4 744 4 387 2 845 3 538 3 302 1 996 27 644

Total Gold Produced (kg) 2 848 3 516 4 418 4 072 2 630 3 239 3 031 1 824 25 579

Total Gold Produced oz 91 557 113 054 142 038 130 929 84 565 104 135 97 448 58 641 822 367

Total Payable Gold (99.90%) oz 91 465 112 941 141 896 130 798 84 480 104 031 97 350 58 582 821 544

Gold Price US$/oz 1 250 1 250 1 250 1 250 1 250 1 250 1 250 1 250 1 250

Gross Revenue US$ M 114.3 141.2 177.4 163.5 105.6 130.0 121.7 73.2 1 026.9

Refining Charges

Refining Charges US$/oz 5.00 5.00 5.00 5.00 5.00 5.00 5.00 5.00 5.00

Refining Costs US$ 457 327 564 706 709 479 653 992 422 401 520 153 486 752 292 912 4 107 722

Refining Costs US$/t ore 0.26 0.35 0.43 0.45 0.32 0.39 0.37 0.48 0.37

Government Royalties

Government Royalties % 3.00% 3.00% 3.00% 3.00% 3.00% 3.00% 3.00% 3.00% 3.00%

Government Royalties US$ 3 416 232 4 218 356 5 299 805 4 885 319 3 155 333 3 885 545 3 636 040 2 188 053 30 684 683

Government Royalties US$/t ore 1.97 2.58 3.22 3.36 2.38 2.88 2.75 3.58 2.77

Government Royalties US$/oz produced 37.31 37.31 37.31 37.31 37.31 37.31 37.31 37.31 37.31

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SECTION 22. ECONOMIC ANALYSIS

22.1 MARKETING The gold bullion bars typically no greater than 30kg each, produced at the Agbaou Project

site may be sent to any of the active gold refiners in the world for toll refining. ENDEAVOUR

is already operating two gold mines and thus gold produced from Agbaou Project will also be

shipped to the same refiners.

22.1.1 Refining Charges, Gold Pricing & Revenue The refining process will be completed within 3 working days after receipt of bullion. The

refiner will charge “refining costs” to cover for melting, assaying and refining.

Generally 99.8% to 99.95% (for the purposes of this study this was assumed to be 99.90%)

of the gold contained in the bullion will be returned to Agbaou Mine and the sale price will be

fixed on the day of the refinery outrun with a settlement of 2 business days.

Bullion transport, insurance and refining costs used were assumed at US$5/oz.

22.2 FINANCIAL ANALYSIS The Agbaou Project economic assessment has been prepared with the input from

ENDEAVOUR, SRK Consulting, (mining and geology), SENET (processing plant, general

administration and infrastructure) and Knight Piésold (tailings management facility & water

storage dam). SENET also undertook the economic evaluation and report compilation.

22.2.1 Evaluation Method Agbaou Project economics has been evaluated using the discounted cashflow method, by

taking into account year on year milled tonnages and grades for the ore and the associated

recoveries, gold price (revenue), operating costs, bullion transport and refining charges,

government royalties and capital expenditure (both initial and sustaining). The project has

been evaluated as stand-alone and 100% equity financed, with no debt financing and the

detailed model.

22.2.2 Assumptions The assumptions used in the financial analysis were provided by ENDEAVOUR and are

summarized in the Table 22.1 and explained below.

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Table 22.1: Assumptions Used in the Financial Evaluation Parameter Units Assumption

Gold Price US$/oz 1,250

Discount Rate % 5.0

Government Royalty % 3

Tax Rate after 5 years % 25

Fixed Equipment Residual Value % 10

Depreciation % 12.5

Gold Price: A gold price of US$1,250 per ounce was assumed.

Construction Capital Expenditure: Capital expenditure payments have been assumed to

be 60% and 40% in Project Years 0 and 1 respectively that is 60% in the first 12 months and

the balance in the following 7 months for the 19 month project duration.

Government Royalties: The Mining Code provides for a 3% royalty payable to the State of

Côte d’Ivoire on all gold production. The royalty is calculated less refining and bullion

transportation costs.

Tax: Corporate Tax of 25% of gross profit will only be applicable after 5 years of production

Depreciation: A straight line method over 8 years was used as the basis for depreciation

from the time production commences for all mechanical equipment.

Inflation: In line with the practice in the mineral industry, no inflation was applied to the

cashflow analysis.

Currency: Cashflow analysis is reported in United States Dollars and where applicable the

exchange rates shown in Table 21.1 were used.

Working Capital: This has been returned at the end of mine life.

Equipment Resale Value: The plant equipment resale value has been assumed at 10% of

the plant equipment costs.

Refining Costs: Bullion transport, insurance and refining costs used were assumed at

US$5/oz

Payable Gold: Only 99.90% of the gold contained in the bullion will be seen as payable gold

(due to non-gold elements)

22.2.3 Financial Analysis Results Using the assumptions shown in Table 22.1 and the production schedules, capital and

operating costs summarized in Table 22.2, a financial analysis was carried out. Sunk costs

were excluded from this analysis.

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Table 22.2: Summary of LOM Production, Capital and Operating Costs

22.2.3.1 Summary of Financial Analysis

Financial analysis results are summarized in Table 22.3.

Table 22.3: Summary of Financial Analysis Results

Financial Summary Units 2nd

Q 2012

LOM Tonnage Ore Processed t 11 074,927

LOM Feed Grade Processed g/t 2.50

LOM Gold Recovery % 92.5%

LOM Strip Ratio

7.87

LOM Gold Production oz 821,728

Production Period years 8.00

Gold Annual Production- LOM oz 102,716

LOM Direct Operating Costs US$/oz 635

LOM Total Cash Operating Costs US$/oz 677

LOM Total Cash Operating Costs US$/t 50.3

Total Capital Costs US$/oz 217.8

Total Production Costs US$/oz 895

Post Tax NPV US$ million 184.3

IRR % 27.7%


Project Net Cash Flow after Tax and Capex US$ million 273.8

Major highlights of the financial analysis are as follows:

Gold Production: Life of mine average production of 102,716 ounces per annum.

Direct Operating Costs: Life of mine average cash operating costs of US$635 per troy

ounce is inclusive of mining, processing, assay, general and administration.

Total Cash Operating Costs: Life of mine total costs of US$677 per troy ounce is inclusive

of direct operating cost, refining and royalty charges. Total Project Costs: Total project costs of US$895 per troy ounce are inclusive of

construction capital, life of mine operating costs and sustaining capital. This cost represents

the break even gold price for the project from commencement of production until after

payback is achieved; thereafter the breakeven price drops to US$677 per troy ounce.

Unit Construction Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 LOM

Tonnes Milled tpa 1 732 010 1 635 744 1 644 002 1 454 229 1 326 569 1 348 689 1 323 336 610 347 11 074 927

Plant Feed Grade g/t 1.76 2.31 2.89 3.02 2.14 2.62 2.50 3.27 2.50

Overall Plant Recovery % 93.17% 93.16% 93.08% 92.68% 92.29% 91.50% 91.60% 91.39% 92.46%

Recovered Gold oz 91 558 113 051 141 978 130 718 84 428 104 092 97 261 58 642 821 728

Mining Costs US$/t 21.79 27.33 29.87 29.70 48.18 35.60 28.66 29.54 30.94

Processing Costs US$/t 9.40 9.52 9.50 9.75 9.91 9.76 9.84 9.94 9.67

G&A and Assay Costs US$/t 5.91 6.26 6.03 6.82 7.47 6.51 6.64 7.30 6.52

Refining Costs US$/oz 5.00 5.00 5.00 5.00 5.00 5.00 5.00 5.00 5.00

Government Royalties US$/oz 37.31 37.31 37.31 37.31 37.31 37.31 37.31 37.31 37.31

Capital Costs (Including Sustaining) US$ 000 158 940 3 436 391 391 8 183 391 391 141 6 731 178 993

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Net Present Value (NPV): The project will realize a post-tax NPV of US$184 million on a

discount rate of 5% and a gold price of US$1,250 per troy ounce.

Nett Cashflow: Net cashflow of US$274 million will be realized at a gold price of US$1,250

per troy ounce.

Internal Rate of Return (IRR) & Payback Period: Project IRR of 27.7% with a 2.53 year

payback period will be realized for the assumed production and capital expenditure.

22.2.3.2 Project Life Cashflow

The project LOM cashflow's by year are given in Table 22.4.

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Table 22.4: LOM Project Cashflow LOM Totals Preproduction 1 2 3 4 5 6 7 8

Gold Production & Revenue

Tonnes Treated 11 074 927 - 1 732 010 1 635 744 1 644 002 1 454 229 1 326 569 1 348 689 1 323 336 610 347

Grade 2.50 - 1.76 2.31 2.89 3.02 2.14 2.62 2.50 3.27

Recovery 0.92 - 0.93 0.93 0.93 0.93 0.92 0.92 0.92 0.91

Gold Production oz 821 728 - 91 558 113 051 141 978 130 718 84 428 104 092 97 261 58 642

Payable Gold % 99.90% 99.90% 99.90% 99.90% 99.90% 99.90% 99.90% 99.90% 99.90%

Gold Price US$/oz 1 250 - 1250 1250 1250 1250 1250 1250 1250 1250

Gross Revenue US$ 000 1 026 133 - 114 334 141 173 177 295 163 234 105 429 129 985 121 454 73 229

Operating Costs

Mining US$ 000 342 630 - 37 742 44 707 49 099 43 195 63 919 48 011 37 924 18 032

Plant US$ 000 107 048 - 16 274 15 569 15 621 14 183 13 151 13 163 13 022 6 066

G & A and Assaying US$ 000 72 233 - 10 236 10 236 9 913 9 913 9 913 8 782 8 782 4 458

Total Opex (Direct) US$ 000 521 911 - 64 253 70 512 74 632 67 291 86 982 69 956 59 728 28 556

Royalties & Refining Charges

Government Royalties US$ 000 30 661 - 3 416 4 218 5 298 4 877 3 150 3 884 3 629 2 188

Refining Charges US$/oz 5.0 - 5.00 5.00 5.00 5.00 5.00 5.00 5.00 5.00

Refining Charges US$ 000 4 109 - 458 565 710 654 422 520 486 293

Total Cash Costs US$ 000 556 680 - 68 127 75 296 80 640 72 822 90 555 74 361 63 843 31 037

Total Cash Costs US$/oz 677 - 744 666 568 557 1 073 714 656 529

After Tax Cashflow

National Reconstruction Tax US$ 000 0 0 0 0 0 0 0 0 0

Operating Benefit US$ 000 469 453 46 207 65 877 96 655 90 412 14 874 55 624 57 611 42 192

Initial Capital Expenditure (Excl WC) US$ 000 (153 900) (153 900)

Deferred Capital US$ 000 0

Sustaining Capital US$ 000 (20 053) (3 436) (391) (391) (8 183) (391) (391) (141) (6 731)

Working Capital US$ 000 0 0 0 0 0 0 0 0 0

Return on WC 0 0 0 0 0 0 0 5 040

Plant Residual Value-Mining US$ 000 0 0 0 0 0 0 0 0 0

Process Plant Residual Value-Process US$ 000 2 426 0 0 0 0 0 0 0 2 426

Profit Before Tax US$ 000 302 966 (153 900) 42 771 65 486 96 265 82 229 14 483 55 233 57 470 42 927

Depreciation US$ 000 (153 900) (19 237) (19 237) (19 237) (19 237) (19 237) (19 237) (19 237) (19 237)

Taxable Income US$ 000 98 434 0 0 0 0 0 35 996 38 233 24 205

Tax Rate % 0% 0% 0% 0% 0% 25% 25% 25%

Tax US$ 000 (24 608) 0 0 0 0 0 (8 999) (9 558) (6 051)

After Tax Profit US$ 000 273 833 (158 940) 42 771 65 486 96 265 82 229 14 483 46 234 47 912 37 391

Undiscounted Cumulative Cashflow After Tax (US$'000) (158 940) (116 169) (50 682) 45 583 127 812 142 295 188 530 236 442 273 833

Discount Rate % 5%

Discounted Cashflow after Tax 184 284 (154 903) 38 795 56 570 79 197 64 429 10 808 32 858 32 429 24 103

Cumulative Discounted Cashflow after Tax (154 903) (116 108) (59 539) 19 658 84 087 94 895 127 753 160 182 184 284

Discount Rate % 5%

Post Tax NPV (US$'000) 184 284

Post Tax IRR (%) 27.7%


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22.2.3.3 Sensitivity Analysis

A sensitivity analysis was performed on the after tax profits by varying the major key

variables (gold price, capex and opex) to ± 30% of the base case cashflow and each

sensitivity was performed independent of the other.

The results of the sensitivities are summarized in Table 22.5, Table 22.6 and

Table 22.7, which show gold price varying from US$875 to US$1,625 per troy ounce, capex

and opex variations.

Figure 22.1 and Figure 22.2 show the detailed sensitivity analysis of changing the key

variables to ± 30%.

Table 22.5: Gold Price Sensitivity

Gold Price

(USD/oz) IRR (%)

NPV (USD 000)

0% 5% 10%

875 -0.2% -1 609 -32 669 -54 116

1063 15.4% 136 340 75 970 33 176

1250 27.7% 273 833 184 284 120 234

1438 38.3% 411 326 292 599 207 291

1625 47.7% 548 818 400 913 294 348

Table 22.6: Capex Sensitivity

CAPEX Change (%) IRR (%) NPV (USD 000)

0% 5% 10%

-30% 41.7% 322 658 232 027 166 891

-15% 33.8% 298 245 208 156 143 562

0% 27.7% 273 833 184 284 120 234

+15% 22.9% 249 421 160 413 96 905

+30% 19.0% 225 008 136 542 73 576

Table 22.7: Operating Costs Sensitivity OPEX Change (%) IRR (%) NPV (USD 000)

0% 5% 10%

-30% 38.8% 418 538 298 175 211 685

-15% 33.4% 346 186 241 230 165 959

0% 27.7% 273 833 184 284 120 234

+15% 21.6% 201 480 127 339 74 508

+30% 14.7% 129 128 70 394 28 782

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Figure 22.1: NPV Sensitivity at 5% Discount Rate

Figure 22.2: IRR Sensitivity at 5% Discount Rate

22.3 DISCUSSION When ranked, the sensitivity analysis indicates that the project is most sensitive to gold price

followed by operating costs and then capital expenditure.

-$100 000

-$50 000

$0

$50 000

$100 000

$150 000

$200 000

$250 000

$300 000

$350 000

$400 000

$450 000

-40% -30% -20% -10% 0% 10% 20% 30% 40%

NP

V U

S$

(0

00

)

Percentage Change

NPV SensitivitiesNPV Gold Price NPV CAPEX NPV OPEX

-10%

0%

10%

20%

30%

40%

50%

60%

-40% -30% -20% -10% 0% 10% 20% 30% 40%

IRR

Pe

rce

nta

ge

Percentage Change

IRR SensistivitiesGold Price Sensitivity Capex Sensitivity Opex Sensitivity

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SECTION 23. ADJACENT PROPERTIES

ENDEAVOUR has no adjacent properties to the Agbaou Exploration Permit area.

Newcrest Mining Limited’s Bonikro Mine is located approximately 30km north-west of the

Agbaou Project (Newcrest Mining Limited, 2006). Commercial production at Bonikro

commenced in August 2008.

Newcrest holds 85% of the shares in Equigold Mines CI SA which in turn holds a 100%

interest in the Bonikro gold project. The Côte d'Ivoire Government holds a 10% interest in

Equigold Mines CI SA, with the remaining 5% held by minority shareholders.

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SECTION 24. OTHER RELEVANT DATA AND INFORMATION

24.1 IMPLEMENTATION

24.1.1 Introduction The Agbaou Project implementation schedule is based on the requirement of producing first

gold by the end of the 4th Quarter 2013. The project execution schedule reflects the work

required from detailed engineering, construction through to commissioning. The schedule

assumes that there is a seamless advancement of the project between the various project

phases.

The project major milestones include the following:

• EPCM contract award – May 2012; • Commencement of construction (earthworks) - August 2012; • The first gold pour will be 19 months after the start of detailed engineering -

December 2013.

It is envisaged that ENDEAVOUR will appoint an independent engineering company to

execute the project on an EPCM basis.

24.1.2 ENDEAVOUR Execution Strategy The strategy for the execution of ENDEAVOUR’s Agbaou Project includes the major roles

and responsibilities mentioned below.

24.1.2.1 Project Manager

Overall responsibility for the execution of the project will fall under the control of the

ENDEAVOUR project manager and his team, to whom the appointed EPCM contractor’s

project manager will report to.

24.1.2.2 Owner’s Team

The project manager will draw on the specialist technical and project execution expertise of

an owner’s team that will typically comprise (and summarized in organogram below):

• Specialists in engineering disciplines essential to the successful execution of the Agbaou project, including a design and technical manager with a team consisting of a project engineer and a mining engineer. The technical specialists will be responsible for ensuring that appropriate standards are applied to all engineering work on the project and approving design and engineering work undertaken by the EPCM consultant;

• An in-country project manager with a team consisting of electrical, construction and earthworks supervisors, surveyors, procurement officers and an administrative team;

• A manager responsible for all matters relating to the environmental, health and safety aspects of the project execution;

• A manager responsible for all community relations pertaining to the execution of the project;

• An accounts manager;

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• A government relations manager responsible for dealing and communicating with government officials on all permitting and legislative matters;

• Legal counsel dealing with the Ivorian mining and corporate legislation.

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Figure 24.1: Owner’s Project Implementation Team

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The design and technical team would be based in Perth, Australia, whilst the rest of the team

would be based in Abidjan or on site, in Côte d’Ivoire.

24.1.2.3 EPCM Consultant

An EPCM consultant will be appointed to manage the engineering, procurement, contract

management and construction management of the Agbaou Project on behalf of

ENDEAVOUR. The EPCM contractor’s project manager will report to the ENDEAVOUR

project manager.

24.1.3 EPCM

24.1.3.1 Engineering

The EPCM Contractor will establish a project specific team to produce the following to

execute the project:

• Project specifications; • Design criteria; • Plant and infrastructure designs; • Data sheets; • Drawings; • Technical reports; • Engineering schedules (equipment lists, motor lists, etc.).

The EPCM contractor will establish design and engineering control functions through a

system of review and approval by the owner’s engineering team.

The EPCM contractor’s standard specifications, design criteria and basis for design will

generally be adopted for the Project. Changes to the standard specifications and design

criteria will be approved by ENDEAVOUR.

24.1.3.2 Procurement

The EPCM contractor will provide all necessary procurement services for the timely

completion of the Project, including:

• Generation of enquiry documents; • Obtaining and control of quotations; • Preparation of commercial adjudications; • Bid clarification and negotiations; • Preparation of orders and contract documentation and any modifications; • Expediting; • Delivery; • Quality assurance and inspections; • Freight forwarding control; • Preparation of control documents and schedules.

24.1.3.3 Construction Management

The construction team will implement and control all aspects of the site work within their

scope, interfacing with ENDEAVOUR as required to satisfy the required scope of work within

the required duration.

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The construction manager will be responsible for ensuring that the work proceeds according

to specifications and project schedules, and that quality and workmanship standards are

maintained. He will also be responsible for the day-to-day site activities, including

maintenance of occupational health, safety and environmental procedures.

24.1.4 Construction The EPCM contractor will ensure that all contractors are mobilized to site in such a time

frame to facilitate the start of their respective responsibilities in line with the execution

schedule. The construction management team will oversee all contractors within their scope.

Construction will also include the following responsibilities:

• Site material control; • Construction planning; • Human resource functions; • SHEQ; • Contract management; • Progress reporting.

24.1.5 Commissioning The commissioning approach will be based on four basic phases as described in a typical

enquiry document:

• Pre-commissioning; • Cold commissioning; • Hot commissioning; • Training and operational assistance period.

The completion of each phase will be signified by appropriate documentation and will be

completed before progressing to the next phase.

The EPCM Contractor will prepare a detailed commissioning manual identifying the

commissioning procedure, all test data and record sheets to be used, the commissioning

team and program.

24.1.6 Project Schedule In the development off the implementation schedule, the following assumptions were made:

• Commencement of detailed engineering, procurement and contracts administration in June 2012;

• Placing orders for long lead delivery items will be in month 1 from start of detailed engineering;

• Bulk Earthworks will commence within 4 months from start of detailed engineering. • The First Gold Pour will be 19 months after the start of Detailed Engineering

(December 2013).

The summarized project schedule is provided in Figure 24.2.

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Figure 24.2: Project Execution Schedule

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24.2 DESIGN CRITERIA

24.2.1 Engineering Design Criteria The proposed engineering design criteria consist of the following sections:

• Civil work; • Structural; • Platework; • Mechanical equipment; • Piping; • Valves; • Painting; • Vendor equipment; • Hand, knee railing, stanchions and stair treads; • Rubber lining.

The following engineering design criteria shown above will apply to the various engineering

disciplines. However, the over-riding basis is that all engineering, designs and construction

will be based on sound engineering principles and good practical applications.

Safety must be considered as a prime objective and all equipment supply and installation

shall comply with recognised international practice.

In all instances, equipment manufacturers’ installation, maintenance and operation

recommendation will be complied with.

The design criteria will at all times take into account ease of maintenance and the capital

expenditure will not be reduced at the expense of operating cost.

24.2.1.1 Civil Work

• Civil design and construction will conform to the relevant sections of SABS 0100. Design will be approved by a registered professional engineer.

• Concrete strengths at 28 days will be: o 25MPa for foundations, slabs and walls; o 10MPa for mass concrete and blinding.

• Concrete finishes will be wood float on all floors. • Walls and bunds will receive smooth off-shutter finish above ground, tolerance class

II. • Slump and cube tests will be performed as required to SABS 0100. • All wet area floor slabs will slope at 1:500 at the valley to a central sump area.

24.2.1.2 Structural

• Structural steel will be supplied in accordance with SABS 1431 or equivalent. • Where installed, roof and side sheeting will be IBR, hot-dip galvanised and 0.6mm

thick. • Flooring will be 30mm x 4.5mm bearer bars and 7.6mm diameter round transversal

bars. Bearer bars 43mm centres and transverse at 50mm centres, bitumen open grid type.

• Where applicable, wind loading will be allowed in accordance with the prevailing

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winds in the area. 24.2.1.3 Platework

• Platework will be predominantly mild steel and painted on all non-process contact surfaces.

• Liners will be installed in wear areas as follows: o Creusabro or similar suitable wear resistant / impact resistant steel, minimum

6mm thick. All steel liners shall be man-handable and fixed with countersunk bolts.

24.2.1.4 Mechanical Equipment

The broader mechanical equipment scope is as specified in the equipment lists provided in

this document.

Conveyors will be fitted with the following:

• Belt slip detectors; • Belt cleaning equipment; • Belt alignment switches, start sirens; • Blocked chute detectors, trip wires and pull switches; • Belt speed sensors; • Sequence switches; • Interlocking devices.

24.2.1.5 Piping

Piping will be supplied and installed to the following specifications:

• Up to 150mm NB: SABS 62 or equivalent, “medium” grade, flanged, screwed or welded connections;

• 200mm NB and larger: SABS 719 or equivalent, grade B, wall thickness 4.5mm minimum or thicker to suit operating duty, flanged or welded connections;

• Flanges: SABS 1123, 600/3 or 1000/3 etc. depending on line operating pressure. Conversion of the design and supply of flanges to ANSI – 16.5 ASA could be provided with minimal material and limited design cost increase;

• Materials of construction: mild steel for all applications, unless otherwise specified; • Connections in piping of diameter 40mm NB and less will be screwed connections; • Piping with a nominal bore exceeding 40mm will be screwed or flanged or welded,

depending on the physical installation, provided that all lines, regardless of size, have adequate flanged connections to ensure ease of removal / maintenance / unblocking.

24.2.1.6 Valves

• Process water, raw water and other clear solution isolation and control valves to be butterfly type.

• Slurry isolation valves to be either knife gate or diaphragm Saunders type “KB”. • All slurry pump suction lines will be fitted with a manual Saunders type KB dump

valve. • Diaphragm Saunders type “A” valves will be used on water hosing and compressed

air points. • All compressed air points will be equipped with an Atlas Copco, or equivalent, type

25 mm quick release coupling.

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24.2.1.7 Painting

The following painting and / or corrosion protection will be applied as a minimum to structural

steel, platework and aboveground piping:

• Abrasive blast clean to Sa 2 ½ (Blast Profile 45-70 µm); • Primer and final coat in fabricators / supplied works is one coat high build zinc

phosphate polyurethane; • Final total D.F.T to be 120µm; • All touch-up work on site and paint repair work will be carried out in accordance with

SABS 630; • All floor grating will be thoroughly cleaned, rust and/or scale and/or grease removed

and bitumen dipped prior to delivery to site. 24.2.1.8 Vendor Equipment

Vendor equipment will be supplied to vendor standard paint specifications and colour,

provided that the above specification as a minimum is met.

24.2.1.9 Hand, Knee Railing, Stanchions and Stair Treads

Tubular hand, knee railing and stanchions will be supplied to site painted as per with

vendor’s zinc chromate finish. A final coat will be supplied once hand railing has been

installed.

24.2.1.10 Rubber lining

• Abrasive blast to clean the mild steel surface in accordance with ISO 8501-01 Grade SA2½ and such that the profile depth is 40-70 microns.

• All rubber lining will be 6 mm thick, apart from flanges where the lining is to be 3mm thick.

• Rubber lining will be done in accordance with SABS 1198 or equivalent. • Rubber lining will be natural rubber to 40 - 50 Shore. • Mill discharge launders & mill sump high wear areas will be lined with either SKEGA,

Solidur low friction liner or ceramic tiles or equivalents.

24.2.2 Electrical Design Criteria

24.2.2.1 Safety

Equipment design, specification and selection have been chosen such that the highest

attention is paid to personnel and equipment safety. For this reason equipment is

conservatively rated and selected on the basis of proven design. Equipment has been

adequately selected to perform in accordance with the manufacturer's ratings where

applicable.

Equipment was selected and specified so as to provide safe enclosures for the environment

that it will be required to operate under. In general all equipment enclosures provide at least

IP41 degree of protection when closed and IP2x classification when normal access doors

are open.

Where enclosures are likely to be subjected to direct hosing down, the enclosure

classification suitable for the duty was chosen as IP65 degree of protection.

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24.2.2.2 Applicable Standards, Regulations and Code of Practice

The following standards and codes have been considered in the design criteria:

General

• The South African Institute of Electrical Engineers; • South African Bureau of Standards (in particular SANS 10142-1); • British Standards Institute; • International Electro technical Commission (IEC); • ESKOM Code of Practice.

Obligatory Standards and Regulations

The applicable obligatory standards and regulations as practised in the Côte d’Ivoire,

Western Africa will be considered during the execution of the project and the specifications

will be updated should they deviate from these requirements.

24.2.2.3 Power Supplies and Power Distribution Philosophy

Equipment Standard Voltages Equipment is specified for the following standard voltages:

• System Voltage Equipment Voltage BIL; • 11kV 17.5kV 95kV (Medium Voltage); • 400V and less 460V 2,5kV (Low Voltage).

Lighting and Small Power Installations Lighting and small power distribution (normal) 400 volts, 3 phase 4 wire system - neutral

solidly earthed. Emergency lighting will be available from an uninterruptable power source.

Control Voltages The following standards will apply:

• 110 Volts AC single phase (Low voltage); • 110 Volts DC derived from a battery tripping unit (Medium voltage); • PLC Input and outputs, 24 volts DC.

Quality of Power Supply The following criteria have been taken into consideration for the electrical equipment and

plant:

All MV and LV equipment will operate within a ±5% voltage drop of the operating voltage,

frequency of 50Hz within a ±2% tolerance and harmonic distortion limited to 3% THD.

Low Voltage Distribution Suitably rated transformers to feed each of the MCC’s will be provided.

Lighting and Small Power Lighting and small power distribution for the plant will be from a dedicated essential and non

– essential lighting distribution board.

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Instrumentation and Control Power Supplies Control and instrumentation voltages will be a combination of 110V, 1-phase, 50Hz and 24V DC 24.2.2.4 Equipment Specification

Standby Plant The power requirements for the proposed standby plant will be 2.4MW at a proposed low

voltage of 400V and frequency of 50Hz (SRK Consulting, SENET, Knight Piésold

Consulting, 2012 Appendix 18-C Refer to drawing number: SP462 9000 E0101 overall single

line diagram). The standby plant will provide power to the critical loads of the plant. Critical

loads to be finalised during execution phase of the project..

The generators will be powered with diesel.

Medium Voltage Switchgear One medium voltage switchgear panel is proposed at 11kV (SRK Consulting, SENET,

Knight Piésold Consulting, 2012 Appendix 18-C).

Suitably rated indoor switchgear equipped with overcurrent earth fault protection schemes

for the incomers and feeders will be provided. Details of protection schemes to be provided

during execution phase of the project.

Transformers Distribution and power transformers will be oil insulated (ONAN), double wound, 3 phase,

50Hz and Dyn11 complying with SANS 780.

Transformers will be designed using the following temperature indicators:

• Oil temperature: o (80-82oC) – Alarm; o (94-95oC) – Trip.

Transformer installations in the vicinity of plant structures or other transformers will be

provided with secure pens and firewalls between transformers and other structures. Cable

routing will run external to transformer pens or as short as possible within the pen.

Single transformers remote from structures will be installed on a slab with a fence around the

transformer.

Provision is made for a number of pole mounted transformers as well. This will be applicable

for all the overhead line supply areas with transformers not exceeding 400kVA (500kVA

transformers will be placed on the ground).

For more detail regarding the layout, refer to the overall single line diagram: SP462 9000

E0101. Transformer sizing to finalised during execution phase of the project.

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24.2.2.5 Motor Control Centres (MCC’s) and Distribution Boards (DB’s)

Eight MCC’s is proposed for the Agbaou Gold Mine plant infrastructure (SRK Consulting,

SENET, Knight Piésold Consulting, 2012 Appendix 18-A describes load list for starter

allocation SP462 1000 1E8 001 and MCC line). Three of which will be containerised due to

remote location and large cables sizes distribution.

MCC’s will be of the compartmentalised non-withdrawable type with moulded case circuit

breakers, magnetic contactors and earth bus, and will comply with the relevant SANS

specification on MCC’s.

The MCC’s will have the following features:

• All enclosures, MCCB’s, busbars and control transformers will be suitably rated for the application and load requirements;

• Front and bottom entry with separate cable ways; • Form 3b design.

24.2.2.6 Electric Motors Low Voltage

Electrical motors have been specified under the mechanical portion of the document.

The control of the motors can be summarised with the following:

• START-STOP or STOP push-button stations will be located at each motor location; • Stations located in wet process areas or outdoors will be of watertight construction; • STOP push-buttons will be of the stay-put type.

24.2.2.7 Electric Cables (HT, LT and Instrumentation)

All cables supplied will be suitably rated for their application considering the environment

and de-rated accordingly. All cables supplied comply to the SANS specification. Suitably

rated joint and termination kits will be provided accordingly.

Cable Racking Cables racking will be used were cables are running on structures, indoors or where cable

support is required. All cable racking have been suitably specified for their application.

Variable speed drives, rectifiers and soft starters VSD’s above 110kW will be housed in their own panel enclosures with an incoming isolator

and door mount handle.

Soft starters are incorporated in the MCC’s designs for motor above 110kW.

Overhead Lines Overhead lines are used for remote loads external to plant areas like the tailing return water,

raw water supply, camp and mining workshop. The overhead lines will be designed and

constructed in accordance with ESKOM code of practice and Côte d’Ivoire code of practice.

Materials, methods of construction and requirements stated here are for 11kV overhead line.

Conductor sizing to be finalised once the project is in the execution phase and layout have

been frozen. Conductors will be large enough to limit voltage drop to 5% of nominal.

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Earthing and Site Lighting Provision has been made for earthing of all electrical equipment and buildings where

applicable. Earthing readings will be conducted to finalise the design during the execution

phase.

Lighting of all buildings has been provided for. Site lighting will make use of pole mounted

lamps and a number of 15m high masts. Storage and workshops will make use of high bay

and low bay type fittings.

Instrumentation and PLC system Provision has been made all the instrumentation as per the process flow diagrams (PFD’s).

The programmable logic controllers (PLC’s) will be housed in the MCC panels and will

communicate on a fibre optic network.

Provision has been made for a supervisory, control and data acquisition (SCADA) system as

well that will be housed in the plant control room.

All drives will be dependent on the PLC for control due to the number of interlocks and

parameters that need to be monitored for safe operation.

The control of the plant will be defined in the process control specification document which

will form the basis for the functional specification.

24.2.2.8 Substation Design

• 400V MCC’s will be installed in block brick buildings with two doors, steel roof trusses, insulation and IBR sheeted.

• Sub-station buildings will be air-conditioned.

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SECTION 25. INTERPRETATION AND CONCLUSIONS

25.1 MINERAL RESOURCE The currently defined Mineral Resources coincides with the areas within the original

Exploration License area where the best mineralization was detected in soil samples.

However, recent sterilization drilling has indicated that the potential exists to discover more

mineralization outside the currently modeled areas, and also in those areas not explored

during the soil surveys.

In the opinion of SRK the sampling preparation, security and analytical procedures used by

ENDEAVOUR are consistent with generally accepted industry best practices and are

therefore adequate.

The apparent course nature of the gold results in a high variability in the field duplicate set.

The BV laboratory results of the blind CRM’s inserted by ENDEAVOUR in the field returned

very good results. Similarly blank material retuned acceptable results and SRK accepts the

BV Laboratory results.

The relatively poor results of the duplicate pulp samples submitted to SGS Laboratory in

Ghana (Referee Laboratory 1) and to the accredited SGS Laboratory in Canada (Referee

Laboratory 2) indicate poor but acceptable replication. It has to be concluded that the scatter

observed although less in the second referee laboratory, is partly caused by the particle

distribution of the gold, although the same should have been evident in the pulp duplicates

analyzed by the BV Laboratory if this was the only cause. The results of the second set of

replicate pulp analyses (Referee Laboratory 2) are however very similar to that of the

independent field coarse duplicates and a slight improvement on the results from the

unaccredited laboratory in Ghana (Referee Laboratory 1).


and accuracy of the sampling and assaying, is adequate and the BV Laboratory returned


The resource evaluation methodology involved the following procedures:

• Database compilation and verification; • Construction of wireframe models for the boundaries of the gold mineralization; • Definition of resource domains; • Data conditioning (compositing and capping) for geostatistical analysis and

variography; • Block modeling and grade interpolation; • Recoverable resource estimation; • Resource classification and validation; • Assessment of “reasonable prospects for economic extraction” and selection of

appropriate cut-off grades; • Preparation of the Mineral Resource Statement.

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The database supplied to SRK was in a Microsoft Access database, and the SRK validations

returned a small number of errors which were corrected through communication with the

ENDEAVOUR database manager. The small number of errors indicates that the database is

well maintained, and the quality of the data is considered to be good.

The 2012 mineral resource estimate is an update of the mineral resources estimated by

Coffey Mining in 2008, following the completion of an additional phase of drilling on the

project. The mineralization has been re-modeled to honor the new intersections, but the

Coffey interpretation was also revised as follows:

• Additional zones were modeled parallel to the main mineralized zones where sufficient continuity along strike and down dip was observed;

• The major zones were extended along strike where additional drilling had intersected the mineralization;

• The SRK interpretation attempted to reduce the low grade dilution within the mineralized envelopes, and this resulted in a trend towards decreased tonnes, at a higher grade, with some isolated high grade samples no longer included in the mineralized envelopes.

A total of ten quartz vein mineralized zones were modeled in the Main, Central and Southern

areas and three zones in the Western area. In addition the mineralized areas in the laterite

cap were modeled as a separate overlying zone. In contrast to the Coffey interpretation,

SRK excluded all laterite material from the quartz vein zones, where Coffey extended the

quartz vein zones into the laterite where the mineralization was elevated. SRK consider the

potential for lateral re-distribution of the gold in the laterite is significant, considering that

there is evidence of lateral transport of the lateritic material in places, and that this material is

better modeled as a single unit.

The weathering zones were modeled in a similar manner to the Coffey interpretation,

defining laterite, saprolite, transitional and fresh domains which formed the basis of the

density assignment.

Following a composite optimization exercise SRK composited all samples within the

mineralized envelopes to 2m lengths, forcing all samples to be included in the composites by

adjusting the composite length at the base of the mineralisation. Nine of the fourteen zones

required capping of high grade outlier values before estimation to limit the impact of the

unusually high values.

SRK was able to generate robust semi-variograms for the better informed zones, however

for some of the smaller modelled zones there was insufficient data available to generate

reliable experimental semi-variograms. The zones for which it was not possible to generate

robust semi-variograms were grouped together with better informed zones with robust semi-

variograms, and borrowed the semi-variograms from the well informed zones. For the well

informed zones the semi-variograms show robust strictures with long ranges between 50m

and 100m, which is in excess of the drill hole spacing in the densely drilled areas.

The QKNA block size optimization exercise indicated an optimal block size of 10m (X) x 20m

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(Y) x 10m (Z) considering both the dimensions of the ore body, the typical intersection

spacing, composite length, and the Kriging statistics for the block sizes tested.

A similar exercise completed to optimize the search neighborhood resulted in the selection

of an optimal number of samples of between 20 and 25 dependent on the zone. Four

separate search orientations were defined for the Main, Central and Southern area to best

match the variable ore body orientation is parts of the ore body. In the West area three

search domains were used. For the laterite estimate, where the mineralized envelope follows

the topography, and has a variable orientation, the search ellipse orientation was

dynamically rotated, based on the orientation of the base of the laterite, to optimally match

the orientation of the mineralized envelope.

All zones were estimated using Ordinary Kriging using the optimized search parameters,

using a three phase search strategy. The first pass estimates used the semi-variogram

ranges to define the search. The second and third passes used expanded search ranges

were designed to ensure the best estimates possible beyond the semi-variogram range, but

also to ensure that all blocks were estimated. This was required as in some areas the

interpreted mineralization showed some abrupt changes in orientation, possibly related to

smaller scale faulting.

In addition to the Ordinary Kriged estimates, for the zones with sufficient information and

reliable semi-variograms SRK generated a UC and MIK estimate. These zones are zones 1,

2, 3, 4, 7, and 9, which (excluding the laterite) make up 90% of the volume of the

North/Central/South area. The UC and MIK estimates were validated against the Ordinary

Kriged estimates at 0g/t cut off, and against each other at a range of cut off values. As the

UC estimate is based on the Ordinary Kriged estimate, the estimates correlate exactly, while

the MIK estimates show a higher grade without applying a cut off. The MIK estimates also

typically show a higher grade at cut-off values with the recoverable grades converging at

higher cut-off values. The MIK estimate therefore estimates that a greater degree of

selectivity will be achievable, than is estimated by the UC method. The MIK estimate was

selected for the final resource reporting to be consistent with previous resource estimates,

however there is a risk that the selectivity modeled using MIK is over estimated, and that the

recoverable tonnes above the selected cut off may be greater than estimated, at a lower

grade.

A number of validations were done on the Ordinary Kriged, UC and MIK estimates which

indicated relatively good agreement with the source data, and the estimates. The mineral

resources are reported above a 0.3g/t cut off value which was calculated using the operation

parameters used in the Reserve calculations, and an optimistic gold price of US$2,000/oz.

25.2 MINING SRK have not seen any updated studies regarding the geohydrology and geotechnical

aspects of the Agbaou deposits. Assuming that these parameters have not changed

adversely there are no reasons that the Agbaou deposits cannot be mined successfully.

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25.3 PROCESS PLANT AND ECONOMIC EVALUATION Based on the results of the previously conducted testwork and the results of the testwork

conducted as part of the EOS, SENET produced a process plant design capable of treating

1.6Mtpa saprolite ore or 1.34Mtpa bedrock ore. The plant design is based on proven

technologies and processes and incorporates a conventional gravity and CIL circuit design.

The key project features are:

• Processed tonnage 11.1 million tonnes • Grade 2.5 g/t • Recovery 92.5% • Life of Mine (LOM) 8 years • Annual gold production 102,716 oz per annum • Total cash costs US$677 per oz. • Capital cost US$159 million

It is in the opinion of SENET that the capital costs provided are within the 10% accuracy

required for this level of study. The operating costs were derived from accepted industry

standards, first principle calculations, SENET in-house database and input from

ENDEAVOUR.

The results of the financial analysis indicate the viability of the Agbaou Gold Project. The

expected NPV, IRR and payback period based on a gold price of US$1,250 per ounce are

indicated below:

• NPV at 0% discount rate is US$273.8 million; • NPV at 5% discount rate is US$184.3 million; • IRR 27.7%; • Payback period 2.53 years.

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SECTION 26. RECOMMENDATIONS

26.1 MINERAL RESOURCE The assay results obtained from the BV laboratory in Abidjan is of a good quality but due to

the poor repeatability in the results obtained from both the SGS laboratories in Ghana and in

Canada it is recommended that an alternative accredited commercial laboratory is selected

for the umpire laboratory function in future.

The current mineral resource estimates contain sufficient material to support the first five

years of mining in the measured and indicated categories. No additional exploration is

required to support the mining plan at this time. Peripheral to the reported resources,

additional potential mineralization targets have been intersected in condemnation drilling

holes. ENDEAVOUR is currently conducting a drilling program to evaluate and better define

these targets.

26.2 MINING The following are mining recommendations for Agbaou (in no particular order):

• More attention should be given to locating waste dumps (based on sterilization drilling) as this will effect mining operating costs;

• More accurate Mining contractors quotes should be obtained; • The comparison between Mining Contractors versus Owner Operated should be

investigated further.

26.3 METALLURGICAL TESTWORK Due to saprolite ore samples being too fine to conduct comminution testwork and EGRG

testwork, it is recommended that this testwork be conducted post EOS. In addition to this it is

recommended that the following testwork be carried out post EOS:

• Recovery testwork on saprolite ore; • Comminution testwork on saprolite ore; • Viscosity modifier testwork on thickener underflow product; • Variability viscosity testwork.

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SECTION 27. REFERENCES

• African Mining Consultants, 2008. Agbaou Gold Process PR177 Environmental Impact Assessment. African Mining Consultants: Kitwe, Zambia.

• Bessoles, B., (1977). Geologie de l'Afrique. Le Craton Ouest Africain; Mem. Bur.

Rech. Geol. Min. Paris, p. 88.

• Blenkinsop, B. Cahen, L. Snelling, N.J. Delhal, J. & Vail, J.R., (1984). The

Geochronology and Evolution of Africa. Oxford University Press: Toronto. pp. 512.

• Canty and Associates LLC, 2012. Weatherbase.[online] Available at:

<http://www.weatherbase.com> [Accessed 2012].

• Coffey Mining Pty Ltd., 2008. Agbaou Gold Deposit Resource Estimation. Coffey

Mining Pty Ltd.: Perth, WA, Australia. Prepared for Etruscan Resources Inc., March

2008 Project Number PAGB05.

• Deijs, W. A. H., 2012. Bulk Solids Flow Report Agbaou Gold Project. Bulk Solids Flow SA cc.: Kempton Park, South Africa. Report no. BSFSA-535.

• Eisenlohr, B.N. and Hirdes, W., 1992. The Structural Development of the Early

Proterozoic Birimian and Tarkwaian Rocks of Southwest Ghana, West Africa. Journal

of African Earth Sciences and the Middle East, pp. 313-325.

• Eisenlohr, B.N., 1989. The Structural Geology of Birimian and Tarkwaian Rocks of

Southwest Ghana. BGR rep. No. 106 448. (unpublished).

• Garland, B., 2012. Agbaou Project Tailings Storage Facility Engineering

Optimisation Study. Knight Piésold Consulting: Johannesburg, South Africa. Report

No. 301-00167/04. Rev 0 (April 2012)

• Gossage, B., Warries, H., 1998. Agbaou Project-Summary-Ore Body Modelling and

Pit Optimisation. Resource Service Group:

• Hillcoat, H., 1999. Agbaou Modelling, Resource Estimation and Pit Optimisation.

Resource Service Group:

• Junner, N.R. and Hirst, T., 1946. The Geology and Hydrology of the Voltaian Basin.

Mem. Geol. Survey Gold Coast, No. 7.

• Knelson Concentrators Africa, 2012. Gravity Recovery Testwork on Samples from SENET (Agbaou Project) Saprolite and Bedrock. Knelson Concentrators Africa: North Riding, South Africa. Report no. SENET 01/12.

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• Knight Piésold Ltd., 2012. Agbaou Gold Project – Gap Analysis for Pit Slope. Knight Piésold Ltd.: Vancouver, BC Canada. File no.: RI301-167/5 .01.

• Newcrest Mining Limited, 2006. Weatherbase.[online] Available at:

<http://www.newcrest.com.au/operations.asp> [Accessed April 2012].

• Orway Mineral Consultants, 2011. Agbaou Gold Project Comminution Testwork Circuit Design Review. Orway Mineral Consultants: Perth. Australia. Report no. 8819.30.

• Orway Mineral Consultants, 2011. Agbaou Gold Project Comminution Testwork Circuit Design Review. Orway Mineral Consultants: Perth. Australia. Report no. 8857.30.

• Patterson & Cooke Consulting Engineers, Pty. Ltd., 2012. Bench-top Thickening Test Report. Patterson & Cooke Consulting Engineers, Pty. Ltd.: Woodlands, South Africa. Report no. SEN-AGB 8284 RO3, Rev 1.

• Patterson & Cooke Consulting Engineers, Pty. Ltd., 2012. Slurry Test Report. Patterson & Cooke Consulting Engineers, Pty. Ltd.: Cape Town, South Africa. Report no. SEN-AGB 8284 RO2, Rev 1.

• Patterson & Cooke Consulting Engineers, Pty. Ltd., 2011. Endeavour Corporation’s Agbaou Project Slurry Test Report. Patterson & Cooke Consulting Engineers, Pty. Ltd.: Cape Town, South Africa. Report no. SEN-AGB 8284 RO1. Rev 1.

• SRK Consulting, SENET, Knight Piésold Consulting, 2012. Agbaou Gold Mine, Cote d’Ivoire, Engineering Optimization Study. SRK Consulting South Africa (Pty) Ltd., SENET, Knight Piésold (Pty) Ltd.: Johannesburg, South Africa. Report prepared for Endeavour Mining.

• Van der Merwe, J., Lotz P., 2012. Leachox and Cyanide Detox Test on Gold Bearing Ore Samples from Agbaou. Maelgwyn Mineral Services Africa: Roodepoort, South Africa. Report no. MMSA Report no. 11/93, Rev 3.

• Vann, J. Jackson, S. & Bertoll, O., 2003. Quantitative Kriging Neighbourhood

Analysis for the Mining Geologist. Australasian Institute of Mining and Metallurgy: Vic

3053 Australia. pp 215-223.

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Technical Report

Appendix 3-A:

Opinion from Theodore Heogan and Michel Ette Advocates Associes

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Appendix 4-A: Decree and Exploitation License

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