ESAASE GOLD PROJECT

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT
1526-REP-003

Table of Contents

				Page
1.0	SUMMARY			1
1.1		Introduction		1
1.2		General Summary		1
1.3		Mineral Resource Estimate		2
1.4		Mining		3
1.5		Metallurgy		4
1.6		Operating Cost Estimage		5
1.7		Capital Cost Estimate		6
1.8		Economic Analysis		7
1.9		Project Development Schedule		9
1.10		Conclusions and Recommendations		11
2.0	INTRODUCTION AND TERMS OF REFERENCE			11
2.1		Sources of Information		11
2.2		Participants		12
2.3		List of Abbreviations		14
3.0	RELIANCE ON OTHER EXPERTS			15
4.0	PROPERTY DESCRIPTION AND LOCATION			15
4.1		Background Information on Ghana		15
4.2		Project Location		17
4.3		Property Description and Ownership		18
4.4		Ownership and Encumbrances		20
4.5		Environmental Liabilities		21
5.0	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY			21
5.1		Access		21
5.2		Physiography and Climate		21
5.3		Local Resources and Infrastructure		22
6.0	HISTORY			22
6.1		Ownership and Exploration History		22
6.2		Resource and Reserve History		22
7.0	GEOLOGICAL SETTING			23
7.1		Regional Geology		23
7.2		Property Geology		24
8.0	DEPOSIT TYPES			28
9.0	MINERALIZATION			29
10.0	EXPLORATION			30
10.1		Soil Sampling Program		30
10.2		Geophysical Programs		32
11.0	DRILLING			33
11.1		Introduction		33
11.2		Drilling Procedure		34
11.2.1			Accuracy of Drill hole Collar Locations	34
11.2.2			Downhole Surveying Procedures	34
11.2.3			Reverse Circulation Drilling Procedures	35
11.3		Diamond Drilling Procedures		35
11.4		RC and Core Sampling Procedures		35
11.5		Summary Results		35
11.6		Drilling Orientation		37
11.7		Topographical Control		37
12.0	SAMPLING METHOD AND APPROACH			39
12.1		RC Sampling and Logging		39
12.2		Diamond Core Sampling and Logging		40
12.3		Sample Recovery		40
12.4		Sample Quality		40
13.0	SAMPLE PREPARATION, ANALYSIS AND SECURITY			41
13.1		Sample Security		41
13.2		Analytical Laboratories		41
13.3		Sample Preparation and Analytical Procedure		42
13.3.1			Transworld Tarkwa	42
13.3.2			SGS Tarkwa	42
13.3.3			ALS Kumasi	42
13.4		Bulk Density Determinations		43
13.5		Adequacy of Procedures		43
14.0	DATA VERIFICATION			43
14.1		Quality Control Procedures		43
14.1.1			Keegan	43
14.1.2			SGS Tarkwa	44
14.1.3			Transworld Tarkwa	44
14.1.4			ALS Kumasi	45
14.2		Quality Control Analysis		45
14.2.1			Transworld Laboratory, Tarkwa	46
14.2.2			SGS Laboratory, Tarkwa	47
14.2.3			ALS Laboratory, Kumasi	48
14.2.4			Keegan QA/QC	55
14.3		QA/QC Conclusions		56
15.0	ADJACENT PROPERTIES			56
16.0	MINERAL PROCESSING AND METALLURGICAL TESTWORK			56
16.1		Metallurgical Testwork		56
16.1.1			Samples	56
16.1.2			Initial Program	57
16.1.3			Main Findings	58
16.1.4			Recent Testwork	61
16.1.5			Comminution Testwork	62
16.1.6			Coarse Ore Bottle Roll Testwork	62
16.1.7			Gravity / Leach Testwork	62
16.1.8			Leach Testwork Reagent Consumption	63
16.1.9			Interpretation of Results	64
16.1.10			Performance Predictions	65
16.1.11			Suitability of Available Testwork	65
16.2		Process Description		65
16.2.1			Summary Design Criteria	65
16.2.2			Process Overview	66
16.2.3			Primary Crushing, Ore Stockpile and Reclaim	66
16.2.4			Grinding and Classification	66
16.2.5			Gravity Concentration	67
16.2.6			Pre-leach Thickener	67
16.2.7			Carbon in Leach (CIL)	67
16.2.8			Elution and Gold Recovery	67
16.2.9			Cyanide Detoxification	68
16.2.10			Services and Water	68
17.0	MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES			68
17.1		Database Validation		68
17.2		Geological Interpretation and Modelling		69
17.2.1			Mineralisation Interpretation	69
17.2.2			Weathering Interpretation	71
17.3		Statistical Analysis		71
17.4		Variography		78
17.4.1			Introduction	78
17.5		Esaase Deposit Variography		78
17.5.1			Zone 100	79
17.5.2			Zone 150	81
17.5.3			Zone 200	62
17.6		Block Modelling		84
17.6.1			Introduction	84
17.6.2			Block Construction Parameters	84
17.7		Grade Estimation		85
17.7.1			Introduction	85
17.7.2			The Multiple Indicator Kriging Method	85
17.8		Multiple Indicator Kriging Parameters		88
17.9		Resource Classification		90
17.10		Resource Reporting		91
17.11		Reserves		91
18.0	OTHER RELEVANT DATA AND INFORMATION			91
19.0	ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES			92
19.1		Preamble		92
19.2		Open Pit Geotechnical Assessment		92
19.3		Proposed Mining Operations		93
19.3.1			Mining Method and Equipment	93
19.3.2			Open Pit Design	93
19.3.3			Waste Storage Area	94
19.3.4			Grade Control	95
19.3.5			Production Schedules and Blending	95
19.4		Hydrology and Hydrogeology		96
19.4.1			Open Pit Dewatering	96
19.4.2			Water Supply	98
19.5		Tailings Storage Facility Designs and Management		98
19.6		Infrastructure and Services		98
19.6.1			Introduction	98
19.6.2			Power Supply	99
19.6.3			Site Buildings and Roads	99
19.6.4			Miscellaneous Facilities	100
19.7		Environmental / Permitting Process		101
19.7.1			Ghanaian Legislation and Guidelines	101
19.7.2			Permitting Process	106
19.8		Project Development Schedule		108
19.9		Project Economics		110
19.9.1			Capital Cost Estimate	110
19.9.2			Operating Cost Estimate	19.3
19.9.3			Economic Analysis	19.5
19.9.4			Sensitivity Analysis	19.9
19.9.1			Sensitivity Response and Range of Variation	19.10
20.0	INTERPRETATIONS AND CONCLUSIONS			20.11
21.0	RECOMMENDATIONS			21.11
21.1		Proposed Work Program to Advance Project to Pre-feasibility Level		21.12
22.0	REFERENCES			22.12
23.0	DATE AND SIGNATURE PAGE			23.13
				PAGE
TABLES
Table 1.1		Study Contributors		2
Table 1.2		Resource Estimate (February 2009)		3
Table 1.3		Summary Operating Cost Estimate (US$, 4Q09)		6
Table 1.4		Summary Capital Estimate (US$, 4Q09, +30% / -15%)		7
Table 1.5		Project Production Summary		8
Table 1.6		Project Cash Flow Summary		8
Table 1.7		Project Financial Measures Summary		9
Table 2.1		List of Abbreviations		14
Table 11.1		Esaase Gold Project Summary Drilling Statistics (Resource Estimate Database)		34
Table 11.2		Esaase Gold Project Summary Drilling Statistics to Current Date		34
Table 11.2		Esaase Project Drilling and Sampling Statistics		35
Table 11.4		Esaase Project Drilling Post Resource Estimate: Significant Intercepts		36
Table 14.1		Transworld Laboratory Tarkwa Laboratory Submitted Blanks and Standards		49
Table 14.2		Transworld Laboratory Tarkwa Field Submitted Blanks and Standards		50
Table 14.3		SGS Laboratory Tarkwa Laboratory Submitted Blanks and Standards		51
Table 14.4		SGS Laboratory Tarkwa Field Submitted Blanks and Standards		52
Table 14.5		ALS Laboratory Tarkwa Laboratory Submitted Blanks and Standards		53
Table 14.6		ALS Laboratory Tarkwa Field Submitted Blanks and Standards		54
Table 16.1		Testwork Samples		57
Table 16.2		Bond Index Determination		60
Table 16.3		SMC Test Results		60
Table 16.4		VFC Composite		61
Table 16.5		VOC Composite		61
Table 16.6		Bond Index Determination		62
Table 16.7		Coarse Bottle Roll Testwork		62
Table 16.8		Gravity / Leach Testwork Overall Results on VOC Composite Gravity Tails		63
Table 16.9		Gravity / Leach Testwork Overall Results on VFC Composite Gravity Tails		63
Table 16.10		Leach Residue and Reagent Consumption on VOC Composite Gravity Tails		63
Table 16.11		Leach Residue and Reagent Consumption on VFC Composite Gravity Tails		64
Table 16.12		Summary Process Design Criteria		65
Table 17.1		RC vs DC Summary Statistics		72
Table 17.2		Esaase Gold Deposit Domain Composite Statistics (Au g/t)		72
Table 17.3		Esaase Gold Deposit Density Statistics (t/m3)		73
Table 17.4		Esaase Gold Deposit Indicator Class Means		77
Table 17.5		Esaase Deposit Zone 100 Correlogram Models		79
Table 17.6		Esaase Deposit Zone 150 Correlogram Models		81
Table 17.7		Esaase Deposit Zone 200 Correlogram Models		82
Table 17.8		Esaase Gold Deposit Block Model Construction Parameters		84
Table 17.9		Esaase Gold Deposit Dry Bulk Density		84
Table 17.10		Esaase Gold Deposit Multiple Indicator Kriging Sample Search Parameters		88
Table 17.11		Esaase Gold Deposit Variance Adjustment Ratios (8mE x 10mN x 2.5mRL SMU)		89
Table 17.12		Esaase Deposit Confidence Levels of Key Criteria		90
Table 17.13		Esaase Deposit Grade Tonnage Report (Multiple Indicator Kriging: 8mE x 10mN x 2.5mRL Selective Mining Unit)		91
Table 19.1		Slope Geometry Recommendations		92
Table 19.2		Summary Mining Schedule 6.5 Mtpa Oxide, 5 Mtpa Thereafter, No Stockpiling		95
Table 19.3		Summary Capital Estimate (US$, 4Q09, +30% / -15%)		110
Table 19.4		Capital Estimate by Area (US$, 4Q09, +30% / -15%)		111
Table 19.5		Summary Operating Cost Estimate (US$, 4Q09)		19.4
Table 19.6		Operating Cost Summary by Expense Type (5 Mtpa Basis) (US$, 4Q09)		19.5
Table 19.7		Summary Manning Estimate		19.5
Table 19.8		Project Production Summary		19.6
Table 19.9		Project Cash Flow Summary		19.7
Table 19.10		Project Financial Measures Summary		19.7
Table 19.11		Sensitivity Analysis Inputs		19.10
				PAGE
FIGURES
Figure 1.1		Preliminary Project Development Schedule		10
Figure 4.1		Location of Esaase Concession SW Ghana		17
Figure 4.2		Keegan Concessions		19
Figure 7.1		Southwest Ghana Geology		24
Figure 7.2		Map of Esaase Concession IP Resistivity, Northeast Structures and Alluvial Mining		26
Figure 7.3		Core Photos of Lithology Types from the Esaase Deposit		26
Figure 7.4		Schematic Structural Model for the Esaase Deposit and Vicinity (Klipfel, 2009)		27
Figure 7.5		Typical Weathering Profile at the Esaase Project Looking North		28
Figure 9.1		Example of Folded and Broken Early Veins		29
Figure 9.2		Example of Sheeted Veining with Visible Gold		30
Figure 10.1		Gold in Soil Thematic Map Esaase (top) and Jeni River (bottom) Concessions		31
Figure 10.2		Gold in Soil Contour Map		32
Figure 10.3		VTEM 92 Meter Layered Earth Inversion (LEI)		33
Figure 11.1		Esaase Project Drillhole Collar Locations		37
Figure 11.2		Drillhole Locations by Drilling Type (Resource Estimate Database)		38
Figure 11.3		Esaase Project Drill Collar Locations Post Resource Estimate Drilling		39
Figure 17.1		Mineralisation Interpretation SE Oblique View		70
Figure 17.2		Mineralisation Interpretation Typical Sectional View		70
Figure 17.3		Weathering Interpretation, Local Grid 9,840m N		71
Figure 17.4		Log Histogram and Probability Plot Zone 100		74
Figure 17.5		Log Histogram and Probability Plot Zone 150		75
Figure 17.6		Log Histogram and Probability Plot Zone 200		76
Figure 17.7		Zone 100 Grade Variogram (Corellogram)		80
Figure 17.8		Zone 200 Grade Variogram (Correlogram)		83
Figure 19.1		Preliminary Project Development Schedule		109
Figure 19.2		Sensitivity Response of After-Tax IRR to Variation in Gold Price, Capital Cost, Processing Cost, Administration Cost, Mining Cost and Royalties		19.9
Figure 19.3		Sensitivity Response of After-Tax NPV Discounted at 5% to Variation in Gold Price, Capital Cost, Processing Cost, Administration Cost, Mining Cost and Royalties		19.10
Figure 19.4		Tornado-type Sensitivity Diagram for Variable Ranges of Sensitivity		19.11

ESAASE PROJECT

Page 1

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

1.0

SUMMARY

1.1

Introduction

Lycopodium Minerals Pty Ltd ("Lycopodium") was commissioned by Keegan Resources Incorporated ("Keegan") to coordinate an Independent Technical Report on the Esaase Gold Project in the country of Ghana, West Africa, in order to summarise the outcomes of the recent Preliminary Economic Assessment undertaken by Lycopodium and Coffey Mining Pty Ltd (Coffey Mining). Additionally, this technical report supports Keegan's news release dated April 6 2010. This report complies with disclosure and reporting requirements set forth in the National Instrument 43-101, Companion Policy 43-101CP, and Form 43-101 F1 .

1.2

General Summary

Keegan through its wholly owned subsidiary Keegan Resources (Ghana) Ltd owns the Esaase Gold Project (the Project), consisting of the adjoining Esaase and Jeni River Mining Leases, and the smaller Mpatoam and Mepom concessions in the Republic of Ghana (Ghana). The concessions are located approximately 35 km south west of the regional capital, Kumasi, and are 45 km north west of AngloGold Ashanti's Obuasi operations, which have been in continuous production for over 100 years.

The Republic of Ghana is a West African country covering 239,460 km² (about the size of Britain). It is one of the five African nations along the northern coastline of the Gulf of Guinea, and is bordered on the west by Cote d'Ivoire, to the north by Burkina Faso, and to the east by Togo. Ghana has substantial natural resources and a much higher per capita output than many other countries in West Africa.

Lycopodium was engaged by Keegan to coordinate this Independent Technical Report to provide a preliminary assessment of the scope of facilities and capital cost required to develop the open pit mine, concentrator, and other facilities, and to provide preliminary operating costs and preliminary economic assessment for a future gold mining operation.

The project scope from which the capital and operating costs have been developed is conceptual in nature and is based on limited information in some areas.

No site visit has been undertaken by Lycopodium as part of the preparation of this report. Lycopodium has, however, considerable experience in the study, development, engineering and construction of similar projects in Ghana through it's participation in multiple projects over a period of 16 years.

Contributions from various sources were compiled for this document as summarised in Table 1.1.

Table 1.1                             Study Contributors

Ownership, environment, community and permitting	Keegan
Geology and mining	Coffey Mining Pty Ltd (Coffey Mining)
Process plant	Lycopodium
Infrastructure and services	Keegan / Lycopodium / Coffey Mining
Tailings storage facility	Coffey Mining
Operating cost estimate	Lycopodium / Coffey Mining
Capital cost estimate (mining)	Coffey Mining
Capital cost estimate (process and ifrastructure)	Lycopodium
Preliminary economic assessment	Keegan / Lycopodium
Project development	Keegan / Lycopodium / Coffey Mining
Proposed work plan	Keegan / Lycopodium / Coffey Mining
Compilation of study document	Lycopodium

1.3

Mineral Resource Estimate

The Esaase Project area contains a system of gold-bearing quartz veins hosted by tightly folded Birimian-age sedimentary rocks.

To date, reverse circulation and diamond drilling has been completed at an adequate density to define an Inferred and Indicated Mineral Resources at Esaase. Mineralisation is currently open, particularly to the north and down dip. Additional satellite mineralisation has been identified at two locations to the west and northwest of the main deposit. This mineralisation does not yet have an associated grade estimate and does not form part of the current resource.

Coffey Mining has estimated the Mineral Resources for the Esaase Gold Project as at 28 February 2009. All grade estimation was completed using Multiple Indicator Kriging (`MIK') for gold. This estimation approach was considered appropriate based on review of a number of factors, including the quantity and spacing of available data, the interpreted controls on mineralisation, and the style of mineralisation. The estimation was constrained with geological and mineralisation interpretations.

The resource estimation was based on the available exploration drillhole database which was compiled by Coffey Mining. The database has been reviewed and validated by Coffey Mining prior to commencing the resource estimation study.

For the purpose of resource estimation, three mineralised domains were interpreted and were modelled on a lower cut-off grade of 0.3 g/t Au.

•

Footwall Domain: A moderately to steeply dipping zone hosting the bulk of the mineralisation and entirely contained within the sedimentary sequence.

•

Hangingwall Domain: A parallel to sub-parallel zone of mineralisation, structurally higher than the footwall domain.

•

South Domain: Previously undefined mineralisation to the south of the previous two domains.

Composites were also coded by the weathering profile.

A summary of the estimated resources for the Esaase deposit is provided in Table 1.2 below. It should be noted that mineral resources that are not mineral reserves do not have demonstrated economic viability.

Table 1.2

Resource Estimate (February 2009)

Esaase Deposit Grade Tonnage Report (Multiple Indicator Kriging; 8mE x 10mN x 2.5mRL Selective Mining Unit)
Lower
Cut-off Grade	Tonnes	Average Grade	Gold Metal
(g/t Au)	(Mt)	(g/t Au)	(Mozs)
Indicated
0.4	57.987	1.2	2.278
0.5	49.248	1.4	2.153
0.6	41.942	1.5	2.025
0.7	35.748	1.7	1.898
0.8	30.656	1.8	1.777
0.9	26.322	2.0	1.660
1.0	22.782	2.1	1.552
Inferred
0.4	41.664	1.2	1.653
0.5	34.054	1.4	1.546
0.6	28.573	1.6	1.451
0.7	24.430	1.7	1.365
0.8	20.649	1.9	1.275
0.9	17.914	2.1	1.201
1.0	15.852	2.2	1.139

Note: Appropriate rounding has been applied.

1.4

Mining

Coffey Mining undertook a mining study as part of the work summarised in this report. The study was based on the total resource, including Inferred.

Geotechnical stability modelling work was based on six oriented drill holes located within the Esaase deposit with the assessments of the core including rock mass quality, rock discontinuity orientations, and discontinuity characteristics.

Whilst limited hydrogeological work has been undertaken Coffey Mining believes that, based on existing information, groundwater may be present in significant quantity and it is probable that high pore pressures will exist within the open pit walls. Overall pit slope angles were based on the assumption that dewatering will result in dry pit walls.

The initial pit optimization for the Project was assessed for three processing rates, namely 3 Mtpa, 5 Mtpa and 7.5 Mtpa and based on the results, the 5 Mtpa scenario was adopted.

Owner mining using a fleet of 200 t class excavators and 90 t rated trucks formed the basis of the study.

The mine planning work was predicated on the pit optimisation results that were based on a gold price of $850/oz. The material breakdown for the pit shell that was selected as the basis for mine production scheduling is provided below.

Total [Mt]	Waste [Mt]	Strip Ratio [ w: o]	Mill Feed [Mt]	Au [g/t]
229	179	3.5	50	1.3

The Study indicates a 10 year project life at an elevated crusher feed target of 6.5 Mtpa of oxide mill feed for the first 2 years and 5 Mtpa of both oxide and fresh material combined thereafter. The elevated mill feed target for the first two years results in a gold production of approximately 260 koz per annum, with an average for the life of mine of approximately 200 koz per annum. The average annual total material movement ranges between 23 Mt and 27 Mt, with an average of approximately 25 Mtpa.

1.5

Metallurgy

Two programmes of metallurgical testwork have been completed on Esaase project material. The first phase in 2008 was based on testing of individual intercepts to determine metallurgical response. Following this, a second programme was completed in 2009 using composites from the same intercepts to determine a likely process flowsheet.

No significant metallurgical problems were identified in the most recent testwork program based on the VOC and VFC composites. The presence of coarse gold was identified and for the purposes of this study a flow sheet including gravity separation, followed by carbon in leach, CIL, was applied to the VOC and VFC composites with acceptable results.

Comparative gravity / CIL leach testwork at 80% passing 150 and 75 microns showed a recovery improvement of 0.5 to 1% for the finer grind. A grind optimisation study which considered the incremental cost benefit of the finer grind indicated a better return with a grind of 150 micron.

For the purpose of this study, a leach time of 24 hours has been used which is typical for the industry. This will be optimised during the next stage of testwork.

Gold recovery predictions are based on the testwork on the VOC and VFC samples only and have used the following methodology:

·

Head grade has been based on a weighted average of all the screen fire assay results for the composite in question.

·

Gravity extraction is assumed as 40% for the calculation with a concentrate leach extraction of 97% based on the master concentrate leach at 150 microns.

·

CIL extraction has been based on the average of the two duplicate 2 kg leach tests at 150 microns with leach residue grades calculated by duplicate 1,000 gram screen fire assay procedures.

·

Soluble loss has been assumed at 0.01 g/t.

·

Recovery is based on the calculated mass of gold extracted at each stage as follows:

·

gold in conc. leach solution + gold in CIL feed - gold in residue - solution loss.

Gold recovery predictions for the two main mineralisation types, is as follows:

Fresh material

94.4%

Oxide material

93.3%

1.6

Operating Cost Estimate

The mining-related operating cost estimates have been developed by Coffey Mining while the operating cost estimates for the process plant and administration have been developed by Lycopodium. The administration operating costs compiled by Lycopodium incorporate information from Coffey Mining and Keegan.

The base mine operating costs for owner mining, excluding mine equipment ownership costs, were estimated at approximately $1.97/t mined.

With the proviso that the anticipated power tariff used for the estimate has been advised by Keegan to be US$0.08/kWh, and that reagent consumptions are based on a limited suite of testwork, the process operating and administration cost estimates are considered to have an accuracy of —15% +20%. Coffey Mining quote an accuracy of ±40% for the mining operating costs. Costs are presented in United States dollars (US$) and are based on prices for the fourth quarter of 2009 (4Q09).

Table 1.3

Summary Operating Cost Estimate (US$, 4Q09)

Cost Centre	Design Basis 5.0 Mtpa		LOM Averaged
	Operating Cost (US$/t milled)	Operating Cost (US$/oz Au recovered)	Operating Cost (US$/t milled)	Operating Cost (US$/oz Au recovered)
Mine Operating Cost	8.91	224.82	8.83	224.80
Processing Cost	7.45	188.07	7.39	188.29
General and Administration Cost	1.17	29.62	1.15	29.25
Rehabilitation / Closure / Reclamation	0.40	10.09	0.40	10.09
Refining	0.16	4.00	0.16	4.00
Royalties	1.18	29.75	1.17	29.75
Total Operating Cost	19.28	486	19.09	486

Note:

1.

Process and administration costs developed by Lycopodium Minerals

2.

Mining and mining administration costs developed by Coffey Mining

3.

The LOM Averaged Operating Cost estimate have been based on a nominal throughput of 6.5 Mtpa during the first two years of operation, followed by a nominal throughput of 5.0 Mtpa, as used for the financial analysis

1.7

Capital Cost Estimate

The capital cost estimate for mining was based on estimates developed by Coffey Mining and was based on an owner-operated fleet.

The mine initial capital expenditure based on owner mining was estimated to be $48.2M. Mine sustaining capital, primarily comprising mine equipment replacement and additional tailings storage facility raises, was estimated at approximately $25.9M.

The capital cost for the process plant and infrastructure was estimated by Lycopodium based on the scope of facilities described in this document.

An allowance of $6 million was advised by Keegan to cover salary and other expenses related to maintaining an owner's project management team. Pre-production expenses were estimated and include an allowance for spare parts and provision of mobile equipment outside the mine fleet.

Table 1.4

Summary Capital Estimate (US$, 4Q09, +30% / -15%)

Mining · Equipment · Pre-production · Mining Infrastructure	$41.7 M $7.5 M $6.5 M	$ 55.7 M
Infrastructure (including roads, buildings, communications)		$ 14.6 M
Mineral Processing Plant		$ 91.8 M
Tailings		$ 9.4 M
Indirects		$ 46.0 M
Owners costs		$ 42.7 M
EPCM		$23.9 M
Subtotal		$ 284.2M
Contingency		$35.4 M
Subtotal (millions)		$319.5 M
Sustaining and Deferred capital		$24M
Total (Millions)		$343M

1.8

Economic Analysis

A preliminary economic assessment of the Esaase Gold Project has been conducted using a simple cash flow model.

The economic evaluation of the Esaase Gold Project has been based upon:

·

Capital cost estimate prepared by Lycopodium and Coffey Mining.

·

Mine schedule and mining operating cost estimates prepared by Coffey Mining.

·

Process and general and administration cost estimates prepared by Lycopodium.

·

Sustaining capital costs for mining and tailings operation provided by Coffey Mining.

·

Owners capital costs based on estimates prepared by Lycopodium Minerals or provided by Keegan.

·

Gold price of US$850 per oz provided by Keegan.

·

Ghanaian corporate tax rate of 25%.

·

Royalties payable 3.5%.

Table 1.5

Project Production Summary

	Basis of Estimate
Mining Schedule
Primary material mined	33.7	Mt
Oxide material mined	16.8	Mt
Waste mined	178.6	Mt
Total Material Mined	229.1	Mt
Total Mill Feed Processed	50.5	Mt
Mine Life	10	years
Contained Gold	2,105	koz Au
Recovered Gold	1,982	koz Au
Average Strip Ratio	3.5	(w : o)
Average Grade	1.30	gAu/t
Average Gold Recovery	94.0	%
Average Annual Plant Throughput	5.04	Mtpa
Average Annual Gold Production	198	koz Au / y

Table 1.6

Project Cash Flow Summary

Year 1-3 US$ Million		Project US$ Million	US$/oz Au Recovered	USM Milled	USM Mined
Mining Cost	149.7	474.2	239.29	9.40	2.34
Processing Cost	119.5	373.1	188.29	7.39	1.66
General and Administration Cost	8.8	29.2	14.75	0.58	0.13
Rehabilitation / Closure / Reclamation	6.0	20.0	10.09	0.40	0.09
Total Operating Cost	284.0	896.6	452	17.77	4.22
Smelting and Refining Cost	2.9	7.9	4.00	0.16	0.04
Royalties	21.7	59.0	29.75	1.17	0.26
Total Cash Cost	308.6	963.5	486	19.09	4.27
Revenue	620.3	1,704.3	850	33.38	7.46
Total Cash Cost	308.6	963.5	486	19.09	4.27
Operating Cash Flow (EBITDA)	311.7	740.8	364	14.28	3.19
Depreciation and Amortisation	322.1	359.9	182	7.13	1.59
Earnings Before Interest & Taxes (EBIT)	(10.4)	381.0	182	7.15	1.60
Interest	-	-	-	-	-
Gross Profit before Tax	(10.4)	381.0	182	7.15	1.60
Tax	-	85.2	14.75	0.58	0.13
Net Profit After Tax	(10.4)	295.7	167	6.57	1.47

Table 1.7                  Project Financial Measures Summary

Mine life	10	Years
Revenue from Gold	1,684	US$ M
Direct cash cost (operating cost only)	452	US$ / oz Au
Total cash cost (including royalties, refining)	488	US$ / oz Au
Capital expenditure (excl working capital)	344	US$ M
Initial capital investment	326	US$ M
Plant and equipment salvage	20	US$ M
Pre-Tax Economics
Free cash flow after cost allocation (undiscounted)	397	US$ M
Internal rate of return (IRR)	19.5	%
Project NPV (discounted at 5.0%)	221	US$ M
Payback period	3.33	Years
After-Tax Economics
Free cash flow after cost allocation (undiscounted)	312	US$ M
Internal rate of return (IRR)	17.2	%
Project NPV (discounted at 5.0%)	168	US$ M
Payback period	3.34	years

1.9

Project Development Schedule

While this report provides a preliminary assessment of the likely economics of the project, prior to committing to detail design, procurement and construction, the concepts outlined will need to be further developed and backed by auditable design data.

A more comprehensive program of metallurgical testwork is required to provide a firm basis of design for the process plant. This is planned for 2010.

Site visits and a more detailed analysis of needs are required to scope the infrastructure and services requirements of the project.

Baseline social and environmental surveys must be completed, environmental impact assessments submitted and approved and a range of project permits obtained.

Additional mining studies are required to finalise the mine design and develop a firm production schedule.

A high level preliminary schedule has been developed (refer Figure 1.1) providing an indication of the likely time scale required for project development. The schedule assumes that development will be based on the current mineral resources with no delays for additional exploration or definition drilling although it does not preclude these happening in parallel with the scheduled activities.

The schedule predicts first gold production at the beginning of Q214.

Figure 1.1

Preliminary Project Development Schedule

1.10

Conclusions and Recommendations

The geological understanding of the Esaase Gold Project has evolved greatly since the commencement of the Keegan exploration program. The knowledge acquired to date and exploration success over the last two years confirms the potential of Esaase and surrounding areas.

The results of the Preliminary Economic Assessment indicates that, subject to further confirmatory testwork and more detailed studies, there is potential for the Project to support the economic development of an open pit mine and treatment plant with a nominal production capacity of 200,000 ounces of gold per year.

Lycopodium considers that the proposed development strategy is entirely appropriate and reflects the potential of the Esaase Gold Project.

2.0

INTRODUCTION AND TERMS OF REFERENCE

Lycopodium Minerals Pty Ltd (Lycopodium) was engaged by Keegan to coordinate this Independent Technical Report summarising the results of a Preliminary Economic Assessment addressing the scope of facilities and capital cost required to develop the open pit mine, concentrator, and other facilities, and to provide preliminary operating costs for a future operation.

The project scope from which the capital and operating costs have been developed is conceptual in nature and is based on limited information in some areas.

No site visit has been undertaken by Lycopodium as part of the preparation of this report. Lycopodium has, however, considerable experience in the study, development, engineering and construction of similar projects in Ghana through it's participation in multiple projects over a period in excess of 10 years.

Coffey Mining have contributed to this report as defined in Table 1.1 and an author of this report, Brian Wolfe of Coffey Mining has made multiple visits to the Esaase Gold Project.

2.1

Sources of Information

Keegan technical staff supplied digital and hard copy data for the Project. In summary, the following key digital data were provided:

·

Drillhole database containing collar location, downhole survey, assay and geology data.

·

A 3-dimensional model of the topography.

·

A representative selection of the original assay sheets.

·

Quality control procedures and database.

·

Internal and external quality control data.

·

A bulk density dataset consisting of 6,121 determinations.

·

Representative cross-sections.

In addition information was supplied by Keegan in relation to project ownership, baseline social and environmental data as well as progress with stakeholder consultation and statutory reporting and permitting.

Technical and cost data was derived from a variety of testwork reports from reputable laboratories, information and budget pricing from vendors of mining equipment and suppliers of consumables and raw materials, as well as Lycopodium and Coffey Mining's in-house databases of information relevant to projects of this nature.

Lycopodium and Coffey Mining has made all reasonable enquiries to establish the completeness and authenticity of the information provided and identified, and a final draft of this report was provided to Keegan along with a written request to identify any material errors or omissions.

2.2

Participants

Resource Estimate

The resource estimation was completed by Coffey Mining Specialist Resource Consultant, Mr Brian Wolfe. Mr Wolfe is a professional geologist with 17 years experience in exploration geology, mining geology and geostatistical modelling and estimation of Mineral Resources. Mr Wolfe is a Member of the Australasian Institute of Mining and Metallurgy (AusIMM).

Mr Wolfe has the appropriate relevant qualifications, experience and independence to be considered a Qualified Person as defined in Canadian National Instrument 43-101 and Competent Person as defined in the Australasian JORC Code.

Mining

The mining review was completed by Coffey Mining Specialist Resource Consultant, Mr Harry Warries, who is a professional mining engineer with 20 years experience in mine optimisation, design, scheduling, cost estimation and cashflow analysis. He is a Principal Consultant with Coffey Mining and a Member of the AusIMM.

Mr Warries has the appropriate relevant qualifications, experience and independence to be considered a Qualified Person as defined in Canadian National Instrument 43-101.

Metallurgy, Process Plant and Infrastructure, Project Development Schedule, Project Economic Assessment

The technical review and conceptual design was completed by Lycopodium Principal Metallurgist, Mr Aidan Ryan. Mr Ryan is a professional metallurgist with 29 years experience undertaking feasibility studies and in the operation and design of mineral processing plants. Mr Ryan is a Member of the AusIMM.

Mr Ryan has the appropriate relevant qualifications, experience and independence to be considered a Qualified Person as defined in Canadian National Instrument 43-101.

Independence

Lycopodium Minerals Pty Ltd is part of the Lycopodium Ltd group, a highly respected Australian-based international consulting firm specialising in area of extractive metallurgy and the design and construction of mineral processing plants and associated infrastructure. Lycopodium Minerals Pty Ltd has completed multiple study briefs for similar projects for other clients in the minerals industry.

Coffey Mining is part of Coffey International Limited (CIL), a highly respected Australian-based international consulting firm specialising in the areas of exploration, geology, mining, metallurgy, geotechnical engineering, hydrogeology, hydrology, tailings disposal, environmental science and social and physical infrastructure.

The authors of this report do not, and are not aware of Lycopodium or Coffey Mining, having or have had previously any material interest in Keegan or related entities or interests. Lycopodium and Coffey Mining's relationship with Keegan is solely one of professional association between client and independent consultant. This report is prepared in return for fees based upon agreed commercial rates and the payment of these fees is in no way contingent on the results of the report.

2.3

List of Abbreviations

Table 2.1 below contains a list of abbreviations.

Table 2.1     List of Abbreviations

3.0

RELIANCE ON OTHER EXPERTS

Neither Lycopodium, Coffey Mining nor the authors of this report are qualified to provide extensive comment on legal and permitting issues, including status of tenure associated with the Esaase property referred to in this report.

Lycopodium and Coffey Mining have relied on advice provided by Keegan, in respect of the property ownership and the related legal dispute, regarding information provided by Ghanaian lawyers Enyonam Dedey and Innocent Akwayena, the latter of Rem Law Consultancy. Both lawyers have consented to the inclusion of their names in this report. This information has not been independently verified by Lycopodium or Coffey Mining. This report has been prepared on the understanding that the property is, or will be, lawfully accessible for evaluation, development, mining and processing.

4.0

PROPERTY DESCRIPTION AND LOCATION

4.1

Background Information on Ghana

Ghana is a West African country covering 239,460 km² (about the size of Britain). It is one of the five African nations along the northern coastline of the Gulf of Guinea, and is bordered on the west by Cote d’Ivoire, to the north by Burkina Faso, and to the east by Togo. The country consists mostly of low savannah regions with a hilly central belt of forest. Ghana’s distinguishing geographic feature is the Volta River, on which was built the Akosombo Dam in 1964. The damming of the Volta created the enormous Lake Volta, which occupies a sizeable portion of Ghana’s south-eastern territory. The country lies immediately north of the equator and has a largely tropical climate.

Ghana’s population is estimated at 23.8 million (July 2009), generally concentrated in the south of the country. The capital, Accra, is a modern coastal city with a population of approximately 2 million people. The second largest city, Kumasi, lies in the heart of the Ashanti region and has about 1.6 million people. Ghana has a large variety of African tribal or sub-ethnic units. The main groups include the Akan (45%), Moshi-Dagomba (15%), Ewe (12%) and Ga (7%) people. Birth rates are high compared with world averages and the annual rate of population growth is one of the highest in the world, although about average for sub-Saharan Africa. Ghana has a relatively young population, with almost one-half of the total population being under 20 years of age. More than two-thirds of the population live in rural areas. The majority of the population are Christian (69%). The northern ethnic groups are largely Muslim (16%). Indigenous beliefs (21%) are also practised throughout the country. English is the official language, while Twi is the most widely spoken African language. Ghana consists of 10 administrative regions. The country is bisected by the Greenwich meridian and operates on Greenwich Mean Time.

Throughout the first half of the twentieth century Ghana (then known as the Gold Coast) was a British colony. It was the first sub-Saharan country in colonial Africa to be granted independence on 6 March 1957. Following a national referendum, it became a republic in July 1960. Between 1966 and 1992, periods of democratic rule alternated with military rule. By 1992, the economy had stabilised, a new constitution was put in place and Ghana returned to democracy with the election of Jerry Rawlings as president. Rawling's National Democratic Congress party continued in power throughout the 1990s, being replaced by the New Patriotic Party in the 2000 democratic election. Ghana has now enjoyed 17 years of continuous democratic rule, with political freedoms and stability which are the envy of other African countries. Ghana is governed under a multiparty democratic system, with elected presidents allowed to hold power for a maximum of two terms of four years. The election held in December 2008 was won by the National Democratic Congress. The elected president is Atta Mills. Next elections are to be held in December 2012.

Ghana has a developing, mixed economy based largely on agriculture and mining although it is also an emerging oil producer. Despite economic difficulties, it is still one of the most developed countries in tropical Africa. The gross national product (GNP) is growing about as rapidly as the population. The GNP per capita is among the lowest in the world, though it is above average for western Africa. The domestic economy of Ghana is dominated by subsistence agriculture, which accounts for about 37% of the gross domestic product (GDP). Most of the working population (60%) grow food crops (plantain, cassava, maize, yams, rice, groundnuts, etc.) for local consumption. The most important cash crop is cocoa. Lesser cash crops include palm oil, rubber, coffee and coconuts. Cattle are farmed in northern Ghana. The most important source of foreign exchange is gold mining, followed by cocoa and timber products. Manganese, bauxite and diamonds are also mined. Tourism is growing rapidly. Gold represents Ghana's major export commodity. Ghana is the world's tenth and Africa's second largest producer of gold, with gold production of 2.0 Moz in 2005. The unit of Ghanaian currency is the Ghanaian Cedi. The exchange rate is presently around 1.4 Ghana Cedis to the US dollar.

Ghana has substantial natural resources and a much higher per capita output than many other countries in West Africa. Nevertheless, it remains dependent on international financial and technical assistance. Inflation, decreasing currency exchange rate and high interest rates have caused concern in recent years, but are improving with more stringent fiscal and monetary policies. Since the early 1980s, the government of Ghana has made a sustained effort to improve and liberalise the fiscal policies of the country in order to attract private investment and stimulate economic growth. Many state-owned companies have been privatised. The result has been a sustained period of real economic growth and an improvement in the country's balance of payments. However, persistent problems remain such as relatively high inflation and unemployment rates.

Under the constitution of Ghana the judiciary is independent of government and cannot be overruled by the president or the parliament. The head of the judiciary is the Chief Justice. The judiciary rules on civil, criminal and constitutional matters. The system includes the Supreme Court, the Court of Appeal, the High Court and Regional Tribunals. There is also a Judicial Council, with representatives from all parts of the justice system, which acts as a forum to observe and review the functioning of the judiciary and to recommend reforms to government. The constitution also dictates that there is an Attorney General who is a Minister of State and is the principal legal adviser to the government.

The capital Accra is serviced by multiple regional and international airlines. The principal ports of Tema and Takoradi are visited regularly by vessels servicing the Europe / West Africa trade routes and provide import / export services for the land locked countries to the north as well as Ghana itself.

Transport infrastructure within Ghana is comparatively good for the region. Since the early 1990’s multiple large, medium and small gold mining operations have been developed in Ghana with both their construction and ongoing operational logistic requirements being met by the two main ports and the internal road network.

Ghana has a relatively well developed power generation and distribution network managed by the Volta River Authority (VRA) and GRIDCo with power generated from the Volta River hydroelectric scheme as well as multiple thermal power stations. Total power capacity within the system is approximately 1,800 MW.

4.2

Project Location

The Esaase Gold Project is located in southwest Ghana, West Africa (Figure 4.1). It is located in the Amansi East district, in the Ashanti Region, approximately 35 km southwest of the regional capital Kumasi.

Figure 4.1                     Location of Esaase Concession SW Ghana

4.3

Property Description and Ownership

The Esaase Gold Project is located in the Amansi West District, in the Ashanti Region, approximately 35 km south west of the regional capital Kumasi. Keegan’s current land holdings consist of the adjoining Esaase and Jeni River mining leases and the smaller Mpatoam and Mepom exploration concessions. The Esaase Mining Lease is approximately 10 km in a north east direction by 3 km in a north west direction covering 28.77 km². The centre of the lease is located at 1º 53’ west, 6º 34’ north. The Jeni River Mining Lease is approximately 10 km in an east west direction and 5 km in a north south direction covering 45.54 km². The centre is located at 1º 98’ west 6 º 52’ north. The Mpatoam Concession is approximately 14 km in a north east direction by 0.8 km in a north west direction covering 8.68 km². The centre is located at 1º 57’ west, 6º 33’ north. The Mepom Concession is approximately 4 km in a north east direction by 0.8 km in a north west direction covering 2.69 km². The centre is located at 1º 56’ west, 6º 33’ north.

The lease / concession boundaries have not been legally surveyed, but are described by latitude and longitude via decree. Figure 4.2 below depicts a plan map of Keegan’s concessions, showing creeks, contours, roads, and concession boundaries.

Figure 4.2                    Keegan Concessions

4.4

Ownership and Encumbrances

Following the bankruptcy of Bonte Gold Mining (BGM) in 2002, the Bonte Liquidation Committee (BLC) was established on behalf of the Government of Ghana to manage the remaining assets.

The Esaase Mining Lease, including the camp facilities at Tetrem, was bought from the BLC by Sametro Company Ltd. (Sametro), a private Ghanaian company. On 3 May 2006, Keegan entered into an option agreement with Sametro to earn 100% of the Esaase Mining Lease that concluded with the successful transfer / assignment of the Esaase Mining Lease to Keegan on 8 June 2007.

On July 14, 2008, Sametro Company Limited ("Sametro") issued a Writ of Summons against Keegan Ghana and the Ghana Minerals Commission in the High Court of Ghana, seeking an order setting aside the Deed of Assignment of the Esaase Concession on the grounds of fraud, a declaration that the Esaase Concession is the property of Sametro, an order for perpetual injunction directed at Keegan Ghana and the Ghana Minerals Commission from interfering with Sametro's rights to the Esaase Concession, an order directed at the Ghana Minerals Commission to expunge from their records the Deed of Assignment which granted Keegan Ghana its ownership of the Esaase Concession plus (unspecified) damages and costs. Sametro's basic claim is that the individual who purported to handle the Deed of Assignment, Mr. Ekow Amua-Sekyi, was either not authorized to do so by Sametro and/or that he provided a deed to Keegan Ghana and the Ghana Minerals Commission which was not in fact signed, as it was purported to be, by the Managing Director of Sametro, Mr. Samuel Gordon Etroo. It is the legal position of Keegan Ghana that Mr. Amua-Sekyi was at all material times a director and agent of Sametro in the transactions between Sametro and Keegan Ghana, and that Keegan Ghana acted in good faith without any notice of irregularity. Keegan Ghana will in part rely upon sections 139-143 of the Ghana Companies Code, 1963 (Act 179) which confers statutory protection on innocent third parties from any fraud by directors, officers, servants and agents of corporate bodies in Ghana. Keegan Ghana believes that Sametro's claims are without merit and will vigorously defend its interests in the Esaase Concession.

Keegan purchased the Jeni River Mining Lease from the BLC, with the lease being transferred / assigned to Keegan on 11 March 2008.

All mining leases currently held by Keegan in Ghana were originally granted for a 30-year period, and can be renewed for additional 30- year terms if necessary. The Esaase and Jeni River mining leases are currently valid until 4 September and 22 March 2020 respectively. The mining leases will allow Keegan to carry out mining provided certain conditions and fee payments are maintained with the Ministry of Lands, Forestry and Mines. The agreements for both leases contain provisions for a 10% government free carried interest (standard in Ghana), a 3% government net smelter return (NSR), and a 0.5% NSR to the BLC.

The Mepom Concession was purchased from a private Ghanaian company, Mepom Mining Company, and transferred to Keegan on 29 June 2009. The Mpatoam Concession is a new concession created at the request of Keegan, and granted on 30 November 2009. Both concessions are covered by prospecting licenses.

The traditional surface rights on the property are owned by the Manso-Nkwanta Paramountcy Stool. At the exploitation stage, the Manso-Nkwanta Paramountcy Stool may apply to the Government of Ghana for the right to a share of the government royalties.

4.5

Environmental Liabilities

The BGM alluvial operation resulted in large land disturbance along the axis of the Bonte valley floor and rerouting of the courses of the relatively small Bonte and Jeni rivers. BGM successfully revegetated a substantial percentage of the disturbed land, however, some of the shallow mined areas, particularly on the Jeni River Mining Lease were not reclaimed by BGM at the time of bankruptcy / closure of the operation. A moderate-sized stockpile of washed gravels remains at the site of the Jeni River recovery plant. On the Esaase Mining Lease, a series of shallow impoundments were constructed by BGM as settling ponds for clay-rich sediments from the recovery plant. The impoundments have naturally revegetated since closure to become a series of wetlands.

Under the agreement with the BLC (acting on behalf of the Government of Ghana), Keegan assumes no liability for any existing environmental liabilities resulting from the operations of BGM on the Esaase Mining Lease. However, on the Jeni River Mining Lease, Keegan has agreed to remediate the existing environmental disturbance should Keegan undertake any mining operation.

5.0

ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1

Access

The Esaase Gold Project is accessed from Kumasi by taking an asphalt road west 10 km to the Bibiani Junction at Asenemuso and then south west 10 km to the village of Wioso. At Wioso, a secondary asphalt road is taken 8 km south to the village of Amankyea then by secondary gravel roads a further 11 km via the villages of Ahewerwa and Tetrem.

5.2

Physiography and Climate

The annual rainfall is in the range of 1,500 mm to 1,700 mm with temperatures ranging from 22°C to 36°C. A major rainy season occurs from April to July followed by a minor one from September through to October. Keegan has operated without cessation or delay throughout all of the rainy seasons to date.

The Esaase property is drained by the Bonte River, which is bounded on both sides by steep hills that reach heights of approximately 500 mASL. The area is predominantly subsistence farmland producing food crops such as plantain, corn, cassava, yam and tomatoes. Some cash crop such as cocoa and oil palm are also grown. About 30% of this land is covered with secondary forest and thick brush. The valley floor has been extensively alluvial mined.

5.3

Local Resources and Infrastructure

The Esaase exploration camp and surrounding communities are connected to the national electrical grid. However, the existing connection does not have the capacity to support a mining / processing operation. The most likely point for future connection to the grid is the Obuasi substation some 40 km to the southeast.

Mobile phone communication is accessible across the project site. A satellite dish is installed in the exploration camp for internet access. The nearest local clinic is situated at Esaase and there is a district hospital at Toase-Nkawie, 20 km from the exploration camp. Larger hospitals and most government offices are available in Kumasi. Food and general supplies are also purchased in Kumasi. The nearest police station is at Ahwerewa.

6.0

HISTORY

6.1

Ownership and Exploration History

The Bonte area has a long history of artisanal mining, associated with the Ashanti Kingdom. There is also evidence of adits driven by European settlers in the period between 1900 and 1939, although no documented records remain of this activity.

Drilling was conducted on the Bonte River valley alluvial sediments during 1966 and 1967 to determine alluvial gold potential. In March and September 1990 respectively, the Jeni River and Esaase mining leases were granted to BGM. BGM reportedly recovered an estimated 200,000 oz of alluvial gold on the Esaase Mining Lease and another 300,000 oz downstream on the Jeni River Mining Lease, prior to entering into receivership in 2002.

Since mid 2006, Keegan has undertaken an aggressive exploration program combining soil geochemistry and IP geophysical surveys, followed by diamond and reverse circulation exploration and resource drilling. It should be noted that previous alluvial gold production is of no relevance to Keegan’s exploration and development program, which is entirely focused on the discovery and exploitation of hard rock resources on the Esaase property.

6.2

Resource and Reserve History

A resource estimate was released by Coffey Mining on behalf of Keegan and detailed in a previous 43-101 report in April of 2009. This resource estimate has been used as the basis for this technical report on the Preliminary Economic Assessment of the Esaase Gold Project. At present there are no reserves at Esaase.

7.0

GEOLOGICAL SETTING

7.1

Regional Geology

The geology of Ghana is comprised predominantly of rocks of the Birimian (2.17-2.18Ga) and to a lesser extent of units of the Tarkwaian (2.12-2.14Ga, after Davis et al. 1994).

The Birimian consists of narrow greenstone (volcanic) belts, which can be traced for hundreds of kilometres along strike but are usually only 20 to 60 km wide, separated by wider basins of mainly marine clastic sediments. Along the margins of the basins and belts there appears to be considerable interbedding of basin sediments and volcanoclastic and pyroclastic units of the volcanic belts. Thin but laterally extensive chemical sediments (exhalites), consisting of cherts, fine-grained manganese-rich and graphitic sediments, often mark the transitional zones. The margins of the belts commonly exhibit faulting on local and regional scales. These structures are fundamentally important in the development of gold deposits for which the region is well known.

The Tarkwaian rocks, on the other hand, consist of a distinctive sequence of metasediments (quartzite, conglomerate and phyllite) occurring within a broad band along the interior of the Ashanti Belt. They host important paleoplacer gold deposits in the Tarkwa district. Equivalent rock types occur in other belts of the region but in relatively restricted areas. In the type locality at Tarkwa, the sequence is in the order of 2.5 km thick, whereas in the Bui belt, comparable units are about 9 km thick sediments that mark a rapid period of erosion and proximal deposition during the late-stage of an orogenic cycle.

All of the Birimian sediments and volcanics have been extensively metamorphosed; the most widespread metamorphic facies appears to be greenschist, although in many areas, higher temperatures and pressures are indicated by amphibolite facies.

Multiple tectonic events have affected virtually all Birimian rocks with the most substantive being a fold-thrust compressional event (Eburnean Orogeny) that affected both volcanic and sedimentary belts throughout the region and to a lesser extent, Tarkwaian rocks. For this reason, relative age relations suggest that final deposition of Tarkwaian rocks took place as the underlying and adjacent volcanic and sedimentary rocks were undergoing the initial stages of compressional deformation. Studies in the western part of the region (Milesi et al., 1992) have proposed several separate phases of folding and faulting suggesting a change in stress direction from northeast to southwest, to north to south. However, a regional synthesis by Eisenlohr (1989) has concluded that, although there is considerable heterogeneity in the extent and styles of deformation in many areas, most of the structural elements have common features, which are compatible with a single, extended and progressive phase of regional deformation involving substantial northwest-southeast compression.

Figure 7.1 below shows the geology of southwest Ghana highlighting the Keegan projects of Esaase and Asumura.

Figure 7.1                       Southwest Ghana Geology

7.2

Property Geology

The Esaase Project area contains a system of gold-bearing quartz veins hosted by tightly folded Birimian-age sedimentary rocks. Geological units on the Esaase property have been interpreted by a combination of airborne geophysical resistivity mapping (VTEM), resource definition drilling and associated outcrop mapping.

The rocks of the property can be divided into metasedimentary units with high electrical and EM resistivity and highly conductive rocks (Figure 7.2). Within the resource zone, the host rocks can be divided between phyllite and siltstone (ore zone predominant in hanging wall of resistivity break) and sandstone / greywacke (predominant in footwall of resistivity break). Host rocks to ore range from massive thinly layered phyllite through interlayered phyllite and siltstone, to massive silt and sandstone (Figure 7.3). Although recognizable stratigraphy appears to be present, the similarity of rock types, folding and faulting precludes correlation of individual stratigraphic units at this stage of core drill and outcrop density.

The structural architecture of the Esaase area is dominated by fold-thrust patterning followed by a late stage strike-slip deformation event. Open to tight, northwest-dipping (axial planes strike (020º to 035º), northeast plunging (30º to 70º) folds are asymmetric and climb to the southeast. Folds tighten and deformation increases systematically to the southeast as shear zones are approached. This patterning repeats itself on the 10 m to 100 m scale. Folding in the deformed siltstone / shale package is open to tight, locally approaching isoclinal. Fold orientation ranges from upright to moderately inclined with dips to the northwest. Folds are asymmetric and climb to the southeast, consistent with regional interpretations of tectonic transport to the southwest. The fold limbs steepen as high strain zones (shears/thrust faults) are approached from the northwest. Within zones of higher strain, weak to moderately developed transposition is common. The footwall side of a shear commonly shows low or lesser strain and repeats the pattern of low to high strain at the next shear. This pattern repeats itself at many scales (micro to macro), but for mapping purposes it is typically on the 10 m to 50 m scale. These northeast striking, northwest-dipping syn-kinematic shears, which roughly parallel fold axial planes appear to demarcate zones of mineralization. In many (but not all) instances, the basal shear / thrust, divides the more deformed, altered, mineralized and electrically conductive siltstone shale unit in the hanging wall from the more massively bedded and less deformed sandstone/greywacke in the footwall. It is common to see broken rock, often carbonaceous, at or near this basal contact indicating likely late brittle faulting. As fault planes cannot be measured on these surfaces, their orientation cannot be clearly determined; thus it cannot be conclusively determined whether this fault or series of fault provide a conclusive footwall boundary. The resistivity contrast provides the best evidence for this contact on a property wide scale and consistent gold assays provide the best evidence on a sectional scale (Klipfel, 2009).

The metasediments are intruded post-kinematically by dikes and small stocks of intermediate to felsic composition, i.e. tonalite to granodiorite. In the southern portion of the deposit, these intrusions are intensely brecciated and mineralized and occur at or near the footwall of mineralization and are themselves mineralized (Klipfel, 2009).

The existence of weathering profile on the Esaase Property is strongly influenced by topography (Figure 7.5). The typical weathering horizon in tropical settings in West Africa consists of laterite (+- duracrust), saprolite, oxidized bedrock, and bedrock (there is often a gradational zone, “saprock” between the saprolite and oxidized bedrock. At the higher elevations at Esaase, the laterite and saprolite, and much of the saprock has been weathered away, leaving behind oxidized bedrock. At intermediate elevations the weathering profile is mostly intact and may be covered by transported colluvium. At the lowest elevations, the entire profile is covered by either alluvium or residual tailings from previous alluvial operations (Klipfel, 2009).

Figure 7.2                  Map of Esaase Concession IP Resistivity, Northeast Structures and Alluvial Mining

Figure 7.3       Core Photos of Lithology Types from the Esaase Deposit

Figure 7.4           Schematic Structural Model for the Esaase Deposit and Vicinity (Klipfel, 2009)

Figure 7.5                Typical Weathering Profile at the Esaase Project Looking North

8.0

DEPOSIT TYPES

The target deposit being explored for is mesothermal quartz vein style mineralisation. This is by far the most important type of gold occurrence in West Africa and is commonly referred to as the Ashanti-type in recognition of the Obuasi area being the type locality and the largest gold deposit in the region. Milesi et al. (1992) recognize that these deposits are largely confined to tectonic corridors that are often >50 km long and up to several kilometres wide and they usually display complex, multi-phase structural features, which control the mineralisation.

The most common host rock is usually fine-grained metasediments, often in close proximity to graphitic, siliceous, or manganiferous chemical sediments. However, in some areas, mafic volcanics and belt intrusions are also known to host significant gold occurrences. Refractory type deposits feature early-stage disseminated sulphides in which pyrite and arsenopyrite host important amounts of gold overprinted by extensive late stage quartz veining in which visible gold is quite common and accessory polymetallic sulphides are frequently observed. This type includes important lode / vein deposits in Ghana such as at Obuasi, Prestea, Bogosu, Bibiani and Obotan. However, a second non-refractory style of gold mineralization occurs in which gold is not hosted within sulphide minerals either in early or late stage mineralization. These type deposits have lower sulphide content in general and in particular, lack the needle-like arsenopyrite that is common in the refractory type deposits. Such deposits include the Chirano and Ahafo type deposits (Stuart, 2007).

9.0

MINERALIZATION

Gold Mineralization on the Esaase Property occurs in quartz - carbonate veins hosted within parallel NE trending, moderately to steeply west dipping bodies of extremely deformed siltstone shale. One form of disseminated alteration most commonly noted in oxidized rocks is quartz-sericite-pyrite (QSP) alteration. This alteration type is not distinctly different in coloration in fresh core and is thus difficult to detect in that state. Surface weathering converts the sericite to white kaolinite creating a bright white color alteration distinguishable even at great distance when exposed in trenches, road cuts, and drill pads. At closer scale, pyrite pseudomorphs can be distinguished. The second stage consists of pervasive carbonate alteration in the form of carbonate porphyroblasts, particularly after andalusite in phyllitic rocks. Carbonate flooding is more prevalent in siltstone where precursor andalusite porphyroblasts did not form (Klipfel, 2009).

As mentioned in Section 7, quartz veining occurred within the mineralization envelopes over most of the duration of the extensive fold and thrust and strike slip deformation events noted in Section 7. Four stages of veins can be identified. These include an early unmineralized quartz only vein stage which has undergone deformation and brecciation. A second vein stage consists of myriad fine spider-web­like quartz-carbonate veins. These veins are also early and are consistently deformed and offset. The third stage consists of quartz-carbonate±sulfide veins with visible free gold. The associated sulfide is generally pyrite, but up to 15% of it can be chalcopyrite and minor arsenopyrite variably occurs as well. Finally, late stage post-mineral calcite veins crosscut all previous features (Klipfel, 2009).

Veins that contain visible gold overwhelmingly strike (350º to 020º), have sub vertical dips and are either planar or S-shaped. Thus they are oblique in orientation to the overall strike and dip of mineralization and appear to be bounded by aforementioned thrust faults and can thus be described as en echelon vein sets form en-echelon sets. As previously describe in Section 7, they likely were emplaced during a transition from fold thrust deformation to left lateral strike slip deformation (see Figure 7.4; Klipfel, 2009).

Figure 9.1            Example of Folded and Broken Early Veins

Figure 9.2                        Example of Sheeted Veining with Visible Gold

10.0

EXPLORATION

No modern-style lode exploration was completed on the Esaase Project prior to the commencement of exploration by Keegan in mid 2006.

10.1

Soil Sampling Program

Keegan commenced a soil sampling program upon acquisition of the Esaase Concession in June 2006 and has received assay results from over 4,000 soil samples. Sampling was undertaken on NE oriented lines spaced 100 - 400 meters apart with samples taken at 25 meter intervals along the lines. This program extended the initial soil sampling completed in March 2006 as part of initial due diligence on the concession. After the acquisition of the Jeni River Concession, Keegan expanded its soil program to the Jeni River Concession and has obtained over 2,100 samples from this concession using an identical sampling regime. Figure 10.1 shows the gold-in-soil contour map derived from these samples. Soil samples were obtained wherever there was not obvious alluvial disturbance or alluvial material and care was taken to sample below the organic horizon. As illustrated in Figure 7.5, the material below the organic horizon on ridge tops or steep slopes from higher elevations is weathered bedrock, whereas that taken nearer to the alluvial creek bottoms is underlain by colluvium, laterite, and/or saprolite. Drilling and trenching indicate that soil samples from weathered bedrock, on average, have gold levels within an order of magnitude of the underlying rock values. Soil samples from non bedrock sources (i.e. alluvial) tend to have much lower gold values than the underlying bedrock. As a result of this observation, Keegan has begun an auger sampling program in order to get samples at (or at leaser closer to) the bedrock / soil interface.

Figure 10.1              Gold in Soil Thematic Map Esaase (top) and Jeni River (bottom) Concessions

Figure 10.2                       Gold in Soil Contour Map

10.2

Geophysical Programs

An IP program was completed in 2006 which successfully identified significant faults that are interpreted as significant mineralization boundaries. In order to identify other such structures, Keegan contracted Geotech Ltd to perform an airborne VTEM geophysical program on the Esaase Property. The survey was carried out during the period 11 October 2007 to 25 October 2007. The principal geophysical sensors included Geotech’s versatile time domain electromagnetic system (VTEM). Ancillary equipment included a GPS navigation system and a radar altimeter. A total of 2,266 line-km were flown. In-field data processing involved quality control and compilation of data collected during the acquisition stage, using the in-field processing centre established in Ghana. The survey was flown at nominal traverse line spacing of 200 m. Flight line directions were N130°E/N50°W. The helicopter maintained a mean terrain clearance of 122 m.

The data was processed and interpreted by Condor Consulting, Inc., who performed AdTau time constant analysis on line data in order to determine the best time delay channels to use. Condor performed Layered Earth Inversions (LEI), generated depth slices for the survey and characterized the 2D and 3D nature of the survey.

The 10 channel map shown in Figure 10.3 is a relatively deep penetrating channel that avoids noise disturbance and provides an overall picture of the resistive characteristics of the rocks. The 92 meter Layered Earth Inversion is useful for a more detailed view of bedrock resistivity at the fresh bedrock surface.

The image indicates significant breaks changes in the resistivity values of the rocks along what are interpreted as NE oriented structures. These breaks correlate with the position and orientation of gold anomalies, which are expressed both in the surface soils that overlie these breaks and in the subsurface as indicated by drilling

Figure 10.3                    VTEM 92 Meter Layered Earth Inversion (LEI)

11.0

DRILLING

11.1

Introduction

Drilling at the Esaase Gold Project has been managed by Keegan and Coffey Mining geologists and to date has been constrained geographically to the south central portion of the Esaase Concession (see Figure 11.1). Surface Reverse Circulation (‘RC’) and Diamond Core (‘DC’) drilling has been completed at the project. The ongoing drill program is designed to test the mineralised corridor delineated from soil sampling, trenching, drilling and geophysical interpretations. The initial 14 diamond drillholes were completed by Eagle drilling contractors with the remainder completed by Geodrill contractors. Both of these drilling companies are reputable Ghana based companies providing RC and Diamond drilling services consistent with current industry standard. Table 11.1 summarises pertinent drilling statistics for all holes drilled on the Esaase concession at the time of commencement of the resource estimate study. A total of 569 drillholes had been completed at the date of this resource estimation study of which 467 of these drillholes (in the currently defined resource area) were used for the resource estimation study.

Keegan has subsequently drilled a total of 271 holes with 43,235 meters of combined core and RC (see Figure 11.3) since the previous 43-101 report was completed in April of 2009. Of that, 17,992 meters were drilled within the resource area for geotechnical, metallurgical and assay data. The remaining 25,243 meters were drilled in 197 stepout exploration holes. Approximately 10,000 meters of drill material awaits assay. The significant intercepts from the post resource estimate drilling are shown in Table 11.4.

Table 11.1   Esaase Gold Project Summary Drilling Statistics (Resource Estimate Database)

Type	Number	Type	Metres
RC holes	399	RC metres	69,953
RC pre-collars with Diamond tails	129	RC pre-collar with Diamond tail metres	38,318
Diamond holes	41	Diamond hole metres	9,688.7
Total Drillholes	569	Total Metres Drilled	117,959.7

Table 11.2         Esaase Gold Project Summary Drilling Statistics to Current Date

Type	Number	Type	Metres
RC holes	596	RC metres	95,196
RC pre-collars with Diamond tails	162	RC pre-collar with Diamond tail metres	48,779
Diamond holes	82	Diamond hole metres	17,220
Total Drillholes	840	Total Metres Drilled	161,195

11.2

Drilling Procedure

11.2.1

Accuracy of Drillhole Collar Locations

Drillhole collars were surveyed by a Coffey Mining surveyor utilising a Thales Promark 3 DGPS unit. This unit was validated as returning sub centimetre accuracy when compared to the topography pickup completed by Coffey Mining using a Geodimeter 610S total station. These instruments have an accuracy of better than 1 cm and are considered conventional.

11.2.2

Downhole Surveying Procedures

Drillholes were surveyed on approximately 50 m downhole intervals, using a Reflex EZ-Shot®, an electronic single shot instrument manufactured by Reflex of Sweden.

These measurements have been converted from magnetic to UTM Zone 30 North values. The factor used to convert between the two grids is -5 degrees.

11.2.3

Reverse Circulation Drilling Procedures

Keegan supervised RC and diamond drilling was completed by Geodrill using a UDR KL900-02 multipurpose track mounted rig. RC rods were 41/2 inch diameter and the drill bit used was a standard diameter.

11.3

Diamond Drilling Procedures

The initial 14 diamond drillholes (HQ and NQ diameters) were completed by Eagle Drilling using a Longyear 38 skid mounted diamond drill. The Geodrill rig utilised in the RC drilling is multipurpose and completed the remaining diamond component of drilling also. The core was oriented by a combination of the spear technique and the Ezimark orientation device.

11.4

RC and Core Sampling Procedures

The sampling procedures followed during RC and DC drilling are detailed in Section 12, as is the sample quality assessment.

11.5

Summary Results

It is not practical to include a listing of all sample results, as a total of 113,069 RC samples and diamond core samples have been collected to date. Table 11.2 summarises pertinent statistics relating to the RC and core sampling program which was utilized in this resource estimate.

Table 11.2             Esaase Project Drilling and Sampling Statistics

Method	Number	Average Length	Total Metres	Number of Assays
RC	399	175	69,953	69,515
RC precollars with diamond tails				19,216 (RC)
	129	297	38,318	18,406 (diamond)
Diamond	41	236	9,688.7	5,932
Total	569	85	117,959.7	113,069

The location of all drillhole collars colour coded by cumulative grade thickness is shown in Figure 11.1. It shows three other zones of drilled mineralization that are not currently included in the current resource. Figure 11.2 displays drilling within the resource area colour coded by type. Note the Figure shows drillholes not used in the grade estimate.

Table 11.4             Esaase Project Drilling Post Resource Estimate: Significant Intercepts

11.6

Drilling Orientation

The vast majority of drillholes in the west dipping Esaase mineralisation were collared at an orientation of approximately 100° (UTM). A small number of holes were drilled towards approximately 300°.

11.7

Topographical Control

Topography has been generated from a Total Station survey completed by Coffey Mining surveyors in 2007. This topography is to an accuracy of +/-30 cm and compares well with the drillhole collar survey data. Coffey Mining considers the topography to be of high confidence.

Figure 11.1                  Esaase Project Drillhole Collar Locations

Figure 11.2               Drillhole Locations by Drilling Type (Resource Estimate Database)

Figure 11.3             Esaase Project Drill Collar Locations Post Resource Estimate Drilling

12.0

SAMPLING METHOD AND APPROACH

12.1

RC Sampling and Logging

RC drill chips were collected as 1 m intervals downhole 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 2 kg was collected for submission to the laboratory for analysis. All 1 m samples were analysed.

RC chip trays were systematically compiled and logged with all bulk rejects being stored at the Keegan exploration camp in Esaase.

12.2

Diamond Core Sampling and Logging

The sampling of the core was subject to the discretion of the geologist completing the geological logging. Initially, nominally 2 m intervals were taken unless geological features were identified requiring smaller intervals. After December 2006, nominally 1 m intervals were taken. 9.8% of diamond core sampling was submitted as whole core with the remaining 90.2% was submitted as half core.

It should be noted that these sampling intervals are much smaller than the true width of mineralised zones, which is variable throughout the deposit, but is typically in excess of 30 m.

After the marking out of the required interval, the core was cut in half by electric diamond blade core saw. The cut is made 1 cm to the right (looking downhole) of the orientation line with the left side being retained and the other half broken up for assay.

In the upper oxide zone, where the core was too friable for diamond saw cutting, the procedure was to dry cut or cleave the core.

Core structure orientations were routinely recorded to assist in determining the controls on mineralisation, in establishing a reliable geological model for resource estimation, and to provide additional geotechnical information to determine likely blast fragmentation and pit stability characteristics.

The core is transferred from the trays and pieced together on a V-rail (angle iron) rack and the orientation line (bottom of hole), determined by the orientation tool recorded during drilling, is drawn along the entire length of the assembled core.

Geotechnical logging has recorded percentage core recovery, RQD percentage, rock type, weathering, rock strength and fractures per metre. This basic geotechnical logging is considered appropriate at this stage of project development.

12.3

Sample Recovery

Sample recovery for RC drilling was noted as very good and averages approximately 34 kg per metre drilled. Bulk sample weights have been systematically recorded for each metre drilled.

Sample recovery in Diamond holes was very good although recoveries for core from the moderate to highly weathered saprolite and highly fractured and brecciated zones returned poor recoveries. Keegan utilised HQ3 drilling to minimise the core loss in the weathered zones.

12.4

Sample Quality

The sampling procedures adopted for drilling are consistent with current industry best practise. Samples collected by diamond coring within the highly weathered zones are of moderate quality, with the remainder being high. Sample recoveries and quality for the RC drilling are high with drilling switching to diamond core once wet samples were noticed.

Dedicated drillhole twinning of the DC drilling and RC drilling has not been completed by Keegan at this stage and it is difficult to determine if any negative bias has resulted in the DC drilling due to the use of water. A number of the DC holes had poor recovery in the highly weathered zone and there exists potential to wash the fine gold associated with the fractures and veining and therefore underestimate the gold content.

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 15) generated and no distinct bias is evident.

13.0

SAMPLE PREPARATION, ANALYSIS AND SECURITY

13.1

Sample Security

The close scrutiny of sample submission procedures by Keegan technical staff, and the rapid submission of samples from drilling for analysis, provides little opportunity for sample tampering. Equally, given the umpire assaying via an external international laboratory and the regular ‘blind’ submission of international standards to both the primary and umpire assay facilities, any misleading analytical data would be readily recognised and investigated.

Current Keegan sampling procedures require samples to be collected in staple closed bags once taken from the rig. They are then transported to the Esaase camp to be picked up by the laboratory truck. The laboratory truck then takes them to the laboratory directly.

Reference material is retained and stored at the Keegan exploration camp at Esaase, as well as chips derived from RC drilling, half-core and photographs generated by Diamond drilling, and duplicate pulps and residues of all submitted samples. Assessment of the data indicates that the assay results are generally consistent with the logged alteration and mineralisation, and are entirely consistent with the anticipated tenor of mineralisation.

13.2

Analytical Laboratories

Preparation and assaying of samples from the Esaase deposit has been carried out at three independent laboratories:

SGS Tarkwa (SGS)

(from April 2007).

Transworld Tarkwa (TWL)

(from October 2006).

ALS Kumasi

(from November 2007).

13.3

Sample Preparation and Analytical Procedure

13.3.1

Transworld Tarkwa

The assay method applied by TWL Tarkwa for the Esaase drilling is summarised below. All aspects of sample preparation and analysis were undertaken at TWL Tarkwa.

•

Sample Preparation

-

3 kg or less of sample is dried, disaggregated, and jaw crushed to 3 mm

-

sample is pulverised to a nominal 95% passing -75 micron using an LM2 pulveriser

-

two pulp samples are taken for analysis and pulp storage.

•

Sample Analysis

-

50 g charge, Fire Assay fusion, lead collection, AAS determination to 0.1 ppm.

13.3.2

SGS Tarkwa

The methodology for the 50 g fire assay from the SGS Tarkwa laboratory is the same as that completed at TWL. All aspects of sample preparation and analysis were undertaken at SGS Tarkwa. SGS is part of the global group of SGS laboratories with ISO/IEC 17025 accreditation.

13.3.3

ALS Kumasi

The assay method applied by ALS Kumasi for the Esaase drilling is summarised below. All aspects of sample preparation and analysis were undertaken at ALS Kumasi. ALS is part of the global group ALS Laboratory Group with ISO 9001:2000 accreditation.

•

Sample Preparation

3 kg or less of sample is dried, disaggregated, and jaw crushed to 2 mm with a nominal 70% passing 2 mm

-

sample is pulverised to a nominal 85% passing -75 micron using an LM2 pulveriser

-

two pulp samples are taken for analysis and pulp storage.

•

Sample Analysis

-

50 g charge, Fire Assay fusion, lead collection, AAS determination to 0.1 ppm.

13.4

Bulk Density Determinations

A total of 6,121 bulk density determinations have been collected for the Esaase deposit by Coffey Mining. The readings were taken over a range of lithological and weathered profiles by Coffey Mining technicians. The procedure used is detailed below and works on the Archimedes Principle. A custom set of “Bulk Density” scales with a weighing hook located underneath (purchased from Corstor South Africa) was utilised for the measurements:

•

10 cm billet of clean dry (dried in an oven for 4 hours at 60^oC) core is weighed.

•

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 as Bulk Density core = [Mass core] / [(Mass air – Mass water) – (Mass wax / 0.9)].

A statistical analysis of the results is presented in Section 17.4.

13.5

Adequacy of Procedures

Analytical procedures associated with data generated to date are consistent with current industry practise and are considered acceptable for the style of mineralisation identified at Esaase. Quality control procedures are described in the next section (Section 14).

14.0

DATA VERIFICATION

14.1

Quality Control Procedures

The quality control procedures adopted by the Keegan and the relevant analytical laboratories are listed in point form below.

14.1.1

Keegan

Keegan has undertaken the procedures recommended by Coffey Mining from January 2007, and include:

·

Insertion of 16 (Geostats Standards and CDN Resource Standards) internationally certified standard reference material (5% of samples).

·

Insertion of Blank material (5% of samples).

·

RC Field duplicates taken (5% of samples).

·

Diamond Core Field duplicates completed by a second split at the 3 mm jaw crushing stage.

·

Submission of selected Umpire samples to SGS.

·

Review of the Keegan and the internal laboratory QC data on a batch by batch basis.

The assay quality control procedures applying to the various laboratories is summarised in the following sections.

14.1.2

SGS Tarkwa

The following quality control procedures are adopted by SGS which is part of the global group of SGS laboratories with ISO/IEC 17025 accreditation:

·

Cross referencing of sample identifiers (sample tags) during sample sorting and preparation with sample sheets and client submission sheet.

·

Compressed air gun used to clean crushing and milling equipment between samples.

·

Barren quartz ‘wash’ applied to the milling / pulverising equipment at the rate of 1:10.

·

Quartz washes assayed to determine the level of cross contamination.

·

Sieve tests are carried out on pulps at the rate of 1:50 to ensure adequate size reduction.

·

Assaying of certified standards at the rate of one per batch of 20.

·

A minimum of 5% (1:20) of the submitted samples in each batch are subject to repeat analysis.

·

Blank samples are inserted at the rate of approximately 1:30.

·

Industry recognised certified standards are disguised and inserted at a rate of 1:30.

·

Assaying of internal standards data.

·

Participation in two international round robin programs; LQSi of USA and Geostats of Australia.

14.1.3

Transworld Tarkwa

TWL applies most of the QC procedures used by SGS although it only participates in the Geostats round robin umpire assay program and it does not utilise the CCLAS computer system. TWL Tarkwa was acquired by Intertek Minerals Group in October 2008. Intertek Minerals Group includes Genalysis Laboratory Services Pty Ltd of Australia and operates in accordance with ISO/IEC 17025, which includes the management requirements of ISO 9001: 2000.

14.1.4

ALS Kumasi

The following quality control procedures are adopted by ALS which is part of the global group ALS Laboratory Group with ISO 9001:2000 accreditation:

·

Cross referencing of sample identifiers (sample tags) during sample sorting and preparation with sample sheets and client submission sheet.

·

Compressed air gun used to clean crushing and milling equipment between samples.

·

Barren ‘wash’ material applied to the milling / pulverising equipment at between sample preparation batches.

·

Quartz washes assayed prior to use to determine the level of cross contamination.

·

Sieve tests are carried out on pulps on a regular basis to ensure adequate size reduction.

·

Assaying of certified standards at the minimum rate of one per batch (dependant on batch size and assay technique).

·

A minimum of one of the submitted samples in each batch are subject to repeat analysis.

·

Blank samples are inserted at the beginning of each batch.

·

Participation in a number of international round robin programs which include CANMET of Canada and Geostats of Australia.

14.2

Quality Control Analysis

The quality control data analysed by Coffey Mining includes:

·

Standard and blanks (both Field and Laboratory).

·

RC Field duplicates.

·

Laboratory repeats.

·

Re-assayed pulps.

·

Umpire assaying.

The assay quality control data, as they pertain to resource estimates completed on the basis of data available, have been subset into the categories above, and reviewed separately.

The quality control data has been assessed statistically using a number of comparative analyses for available datasets. The objectives of these analyses were to determine relative precision and accuracy levels between various sets of assay pairs and the quantum of relative error. The results of the statistical analyses are presented as summary plots, which include the following:

•

Thompson and Howarth Plot showing the mean relative percentage error of grouped assay pairs across the entire grade range, used to visualise precision levels by comparing against given control lines.

•

Rank % HARD Plot, which ranks all assay pairs in terms of precision levels measured as half of the absolute relative difference from the mean of the assay pairs (% HARD), used to visualise relative precision levels and to determine the percentage of the assay pairs population occurring at a certain precision level.

•

Mean v's % HARD Plot, used as another way of illustrating relative precision levels by showing the range of % HARD over the grade range.

•

Mean vs % HRD Plot is similar to the above, but the sign is retained, thus allowing negative or positive differences to be computed. This plot gives an overall impression of precision and also shows whether or not there is significant bias between the assay pairs by illustrating the mean percent half relative difference between the assay pairs (mean % HRD).

•

Correlation Plot is a simple plot of the value of assay 1 against assay 2. This plot allows an overall visualisation of precision and bias over selected grade ranges. Correlation coefficients are also used.

•

Quantile-Quantile (Q-Q) Plot is a means where the marginal distributions of two datasets can be compared. Similar distributions should be noted if the data is unbiased.

Comments on the results of the statistical analyses for each laboratory are provided below.

14.2.1

Transworld Laboratory, Tarkwa

TWL Duplicate Repeats

At TWL, every 20th sample is duplicated. A duplicate is two separate samples taken from the total pulped sample. Duplicate repeats are analysed in the same batch and are therefore not subject to intra-batch variance. Only assays greater than 10 times the detection level (>=0.1 ppm Au) are included in the assessment and data are divided into drillcore (HQ and NQ, 177 assays) and riffle split 1 m RC drill chips (461 assays). Results show equivalent means between the duplicate repeats and precision within acceptable limits for both diamond core and RC samples.

TWL Pulp Respray

After initial calibration of the AAS with control standards, the batch is sprayed (the aspirator tube is placed in the DIBK layer and approximately 1 ml is sprayed into the AAS flame). On combustion, the absorbance is measured by the AAS and the strength of the absorbance is proportional to the gold concentration). At the end of spraying, the operator returns to every 10th samples and performs the same operation and this is the Pulp Respray. At the end, control samples are again presented to the AAS to verify that short term drift has not occurred. Only assays greater than 10 times the detection level (>=0.1 ppm Au) are included in the assessment for a total of 1,202 assays. Results show equivalent means between the duplicate repeats and precision well within acceptable limits.

TWL Check Repeats

Check repeats occur where high grade samples are encountered or where the result is out of sequence (e.g. 0.01-0.04-0.02-1.2-0.03: Result 1.2 is out of sequence and would be repeated). A repeat is a second 50 g sample taken from the same kraft envelope as the original analysis (Au1) and is thus different from the duplicate repeat. Check Repeats are analysed later than the original assay (in a different batch) and may therefore be subject to intra-batch variance compared with the original result. Only assays greater than 10 times the detection level (>=0.1 ppm Au) are included in the assessment and data are divided into drillcore (HQ and NQ, 265 assays) and riffle split 1 m RC drill chips (573 assays). Check Repeat analyses data to September 2007 was available for review. Results show equivalent means between the duplicate repeats and precision within acceptable limits for both diamond core and RC samples.

TWL Pulp Reassay

Only pulp reassays greater than or equal to 10 times the detection level (0.1 ppm Au) are considered for analysis and these comprise 1,615 riffle split 1 m RC drill chip assays. Results show equivalent means between the duplicate repeats and precision within acceptable limits.

TWL Lab Standards and Blanks Analysis

Six certified standards were inserted by TWL into the sample batches at a rate of one in twenty in addition to preparation blanks and reagent blanks at a similar rate. The supplied database only contains Lab standards analysis received to September 2007. A total of 3,512 standards and blanks assays are available for analysis. Results generally show a positive bias of between -0.44% to 3.05%. This positive bias is more evident for higher grade standards.

14.2.2

SGS Laboratory, Tarkwa

SGS Duplicate Second Split

This comprises RC (339) and diamond core (73) field duplicates and is achieved by taking a second split at the 3 mm jaw crushing stage of the sample preparation. Results show equivalent means and a high level of precision between the original and the reassay for both diamond core and RC samples.

SGS Replicate First Split

These assays represent a random repeat assay with four random repeats completed from each batch of 50 samples. A total of 582 Diamond core and 2,392 RC analyses are available for analysis. Results show equivalent means and an acceptable level of precision between the original and the reassay.

Lab Standards and Blanks Analysis

Four certified standards were inserted by SGS into the sample batches at a rate of one in twenty in addition to preparation blanks and reagent blanks at a similar rate. The supplied database only contains Lab standards analysis received to September 2007. A total of 938 standards and blanks assays are available for analysis. Results show a relative low bias of up to -2.09%.

14.2.3

ALS Laboratory, Kumasi

ALS Duplicate Second Split

This comprises RC (176) and diamond core (62) duplicates and is achieved by taking a second split at the 3 mm jaw crushing stage of the sample preparation. Results show equivalent means and a high level of precision between the original and the reassay for the diamond core samples. Results for the RC samples demonstrate a high level of precision between the original and the reassay however the second mean is 7.5% lower than the original assay.

ALS Replicate

These assays represent a random repeat assay of a second sample taken from the original pulp. A total of 223 diamond core and 892 RC analyses are available for analysis. Results show equivalent means for diamond core however the second mean for the RC samples is significantly lower than the original. Overall levels of precision between the original and the reassay are low for both diamond core and RC samples.

ALS intra Batch Analysis

These assays represent a random repeat assay analysed in a different assay batch to the first. Results show equivalent means and acceptable precision (although at the lower end) for both RC and diamond core samples.

Table 14.1

Transworld Laboratory Tarkwa Laboratory Submitted Blanks and Standards

Standard Name	Expected Value (EV)	+/-10% (EV) (9/t)	No of Analyses	Minimum (9/t)	Maximum (9/t)	Mean (9/t)	% Within +/- 10 of EV	% RSD (from EV)	% Bias (from EV)
TWL Submitted Blanks
Reagent Blank	0.005	0.0045 to 0.0055	612	0.005	0.02	0.005	98.86	18.96	1.80
Sample Blank	0.005	0.0045 to 0.0055	1370	0.005	0.02	0.005	98.98	12.47	1.17
TWL Submitted Standards
BM292	1.48	1.33 to 1.63	66	1.41	1.60	1.50	100	2.62	1.1
ST06_5322	1.04	0.94 to 1.14	108	0.97	1.11	1.04	100	2.52	-0.44
ST06_5356	1.04	0.94 to 1.14	466	0.97	1.12	1.05	100	2.26	0.48
ST17_2290	0.78	0.70 to 0.86	595	0.72	0.85	0.79	100	2.68	1.61
ST343	1.286	0.18 to 0.22	218	0.19	0.23	0.20	98.62	3.91	2.09
ST364	8.59	7.73 to 9.45	68	8.20	9.31	8.85	100	3.10	3.05

Table 14.2

Transworld Laboratory Tarkwa Field Submitted Blanks and Standards

Table 14.3

SGS Laboratory Tarkwa Laboratory Submitted Blanks and Standards

Standard Name	Expected Value (EV)	+/-10% (EV) (g/t)	No of Analyses	Minimum (g/t)	Maximum (g/t)	Mean (g/t)	% Within +/- 10 of EV	% RSD (from EV)	% Bias (from EV)
SGS Submitted Blanks
Reagent Blank	0.005	0.0045 to 0.0055	179	0.005	0.01	0.005	94.41	21.75	5.59
Sample Blank	0.005	0.0045 to 0.0055	157	0.005	0.02	0.005	92.99	36.84	9.55
SGS Submitted Standards
ST05_2286	2.36	2.12 to2.60	151	2.14	2.52	2.33	100	2.09	-1.27
ST14_6368	0.41	0.37 to 0.45	164	0.38	0.42	0.40	100	2.26	-2.90
ST21_5327	6.83	6.15 to 7.51	122	6.21	7.35	6.76	100	2.64	-1.03
ST37_8229	1.73	1.56 to 1.90	165	1.62	1.84	1.71	100	1.76	-1.30

Table 14.4

SGS Laboratory Tarkwa Field Submitted Blanks and Standards

Standard Name	Expected Value (EV)	+/-10% (EV) (g/t)	No of Analyses	Minimum (g/t)	Maximum (g/t)	Mean (g/t)	% Within +/- 10 of EV	% RSD (from EV)	% Bias (from EV)
Keegan Submitted Blanks
Sample Blank	0.01	0.009 to 0.011	2097	0.005	0.2	0.025	-	-	-
	Keegan Submitted Standards
CDN-BL-3	0.01	0.009 to 0.011	150	0.13	1.3	0.017	-	-	-
CDN- GS-15A	14.83	13.35 to 16.31	81	13.7	18.6	15.86	66.67	7.47	6.93
CDN-GS-1C	0.99	0.89 to 1.09	50	0.81	1.11	1.024	86	5.05	3.43
CDN_GS_30A	35.25	31.73 to 38.78	70	32.3	44.2	35.38	97.14	4.58	0.37
CDN-GS-P5B	0.44	0.40 to 0.48	161	0.27	0.54	0.47	55.90	8.44	7.30
G306-3	8.66	7.79 to 9.53	70	7.97	9.45	8.7	100	2.68	0.46
G396-5	7.36	6.62 to 8.10	81	6.51	7.54	6.91	97.53	1.86	-6.15
G901-1	2.58	2.32 to 2.84	40	2.56	2.77	2.63	100	1.76	2.07
G901-11C	1.34	1.21 to 1.47	222	1.18	1.43	1.30	99.55	1.97	-3.30
G901-7	1.52	1.37 to 1.67	78	1.29	1.72	1.39	35.90	7.83	-8.41
G901-9	0.69	0.62 to 0.76	373	0.53	0.81	0.65	69.44	6.03	-6.56
G905-10	6.75	6.08 to 7.43	271	6.06	7.22	6.80	99.63	2.30	0.69
G905-5	0.52	0.47 to 0.57	352	0.43	0.67	0.53	83.52	7.55	1.92
G995-1	2.74	2.47 to 3.01	410	2.38	3.40	2.66	96.10	4.14	-2.77
G997-9	5.16	4.64 to 5.68	344	4.08	5.65	4.95	97.97	3.23	-4.14
G995-6	7.18	6.46 to 7.90	48	6.65	7.38	6.87	100	1.41	-4.28

Table 14.5

ALS Laboratory Tarkwa Laboratory Submitted Blanks and Standards

Standard Name	Expected Value (EV)	+/-10% (EV) (g/t)	No of Analyses	Minimum (g/t)	Maximum (g/t)	Mean (g/t)	% Within +/- 10 of EV	% RSD (from EV)	% Bias (from EV)
ALS Submitted Blanks
Reagent Blank	0.005	0.0045 to 0.0055	179	0.005	0.01	0.005	94.41	21.75	5.59
Sample Blank	0.005	0.0045 to 0.0055	157	0.005	0.02	0.005	92.99	36.84	9.55
ALS Submitted Standards
ST05_2286	2.36	2.12 to2.60	151	2.14	2.52	2.33	100	2.09	-1.27
ST14_6368	0.41	0.37 to 0.45	164	0.38	0.42	0.40	100	2.26	-2.90
ST21_5327	6.83	6.15 to 7.51	122	6.21	7.35	6.76	100	2.64	-1.03
ST37_8229	1.73	1.56 to 1.90	165	1.62	1.84	1.71	100	1.76	-1.30

Table 14.6

ALS Laboratory Tarkwa Field Submitted Blanks and Standards

Standard Name	Expected Value (EV)	+/-10% (EV) (g/t)	No of Analyses	Minimum (g/t)	Maximum (g/t)	Mean (g/t)	% Within +/- 10 of EV	% RSD (from EV)	% Bias (from EV)
Keegan Submitted Blanks
Sample Blank 0.01 0.009 to 0.011 292 0.01 0.06 0.01 - - -
Keegan Submitted Standards
CDN-BL-3	0.01	0.009 to 0.011	61	0.01	0.04	0.01	-	-	-
CDN- GS-15A	14.83	13.35 to 16.31	50	11.00	16.85	14.78	96.00	5.48	-0.32
CDN_GS_30A	35.25	31.73 to 38.78	44	27.20	41.40	35.38	79.55	8.02	0.35
CDN-GS-P5B	0.44	0.40 to 0.48	143	0.28	0.50	0.43	89.51	7.03	-3.16
G306-3	8.66	7.79 to 9.53	111	7.89	9.83	8.78	97.30	4.72	1.41
G396-5	7.36	6.62 to 8.10	3	7.63	8.78	8.28	33.33	5.80	12.45
G901-1	2.58	2.32 to 2.84	49	2.19	4.00	2.70	75.51	10.10	4.69
G901-11C	1.34	1.21 to 1.47	141	1.06	1.59	1.34	89.36	6.36	-0.31
G901-7	1.52	1.37 to 1.67	76	1.33	1.97	1.52	96.05	5.43	0.00
G901-9	0.69	0.62 to 0.76	145	0.53	0.80	0.68	90.34	5.65	-1.41
G905-10	6.75	6.08 to 7.43	31	5.55	7.86	6.78	80.65	7.81	0.38
G905-5	0.52	0.47 to 0.57	149	0.36	0.60	0.50	87.25	6.87	-3.65
G995-1	2.74	2.47 to 3.01	146	1.81	4.33	2.89	70.55	10.20	5.64
G997-9	5.16	4.64 to 5.68	150	3.54	6.59	5.24	87.33	7.80	1.48
G995-6	7.18	6.46 to 7.90	90	6.75	7.96	7.34	96.67	3.50	2.17

14.2.4

Keegan QA/QC

Keegan Field Standards and Blanks

A total of 16 Certified Standards and one blank have been included in sample batches sent to TWL, ALS and SGS. A total of 11,507 assays were available for analysis. Where identifiable, outliers to the data which are obviously a misplaced standard have been removed from the data before analysis resulting in 9,818 valid standard assays.

Results show a moderate positive bias of up to 6.09% for Transworld Laboratories. There is no relationship between grade and bias. One standard shows negative bias of -5.33%.

Blind standards analysis at SGS shows a spread of bias with one standard displaying a significant negative bias of up to -8.41%. In addition, one standard shows a positive bias of 6.93%. Again, there is no relationship between grade and bias.

Blind standards analysis at ALS shows a spread of bias from -3.65% to 5.64%. Negative bias is apparent at lower grades and positive bias up to 5.64% is seen in two standards at 2.58 g/t Au and 2.74 g/t Au. For higher grade samples the bias approaches zero.

Keegan Field Duplicates

Field duplicates totalling 1,567, 1163 and 2,802 have been sent to TWL, ALS and SGS respectively. Diamond core field duplicates consist of a portion of the “coarse rejects” obtained after the crushing stage. RC field duplicates consist of a second sample split from the reject sample in the field. Only assays returning values greater than ten times the detection limits (>0.1 ppm Au) and less than 5 g/t Au have been considered in the analysis.

Results for TWL, SGS and ALS show equivalent means and acceptable precision for both RC and diamond core samples.

Keegan Assay Resplits (Umpire)

In January and February 2007 a total of 1,197 RC samples were re-split and sent for analysis at SGS Tarkwa (TWL was the primary laboratory for the initial analysis). Only assays >0.1 g/t Au are considered in the analysis and a total of 481 assay pairs are available for analysis. Results show a significantly lower mean (by 15.6%) for analysis completed at SGS (although this is significantly reduces if outliers to the data are removed).

SGS Tarkwa has been utilised as a primary laboratory for the project since February 2007 and umpire samples numbering 1,633 have subsequently sent to Genalysis of Perth for umpire analysis. Only assays >0.1 g/t Au are considered in the analysis and a total of 1,572 assay pairs are available for analysis. Results show equivalent assay means for the pairs between ALS and Genalysis and between SGS and Genalysis. The means of the assay pairs between TWL and Genalysis show high bias for TWL, a finding which is supported by Standards analysis (Section 14.2.4). Precision is less than acceptable for all comparisons and this requires investigation.

14.3

QA/QC Conclusions

Coffey Mining believes that the current QA/QC systems in place at Esaase to monitor the precision and accuracy of the sampling and assaying are adequate and should continue to be implemented. Pertinent conclusions from the analysis of the available QA/QC data include:-

•

Use of Certified Standard Reference material has shown a significant relative low bias for SGS Laboratories, Tarkwa.

•

Use of Certified Standard Reference material has shown a relative high bias for Transworld Laboratories, Tarkwa and this interpretation is supported by the umpire analysis program.

•

Repeat analyses have confirmed that the precision of sampling and assaying is generally within acceptable limits for sampling of gold deposits.

•

Umpire analysis at Genalysis in Perth has shown a lack of precision between the various laboratories. This is currently unexplained and requires investigation.

•

Other relevant conclusions are discussed throughout Section 14.

15.0

ADJACENT PROPERTIES

There are a number of operating mines in proximity (<100 km) to the Esaase Gold Project. They include world class gold deposits such as the Obuasi project operated by Anglo Ashanti, and the Akyem Gold project that is currently being developed by Newmont mining.

16.0

MINERAL PROCESSING AND METALLURGICAL TESTWORK

16.1

Metallurgical Testwork

Two programmes of metallurgical testwork have been completed on Esaase project material. The first phase in 2008 was based on testing of individual intercepts to determine metallurgical response. Following this, a second programme was completed in 2009 using composites from the same intercepts to determine a likely process flowsheet. This section summarises the results of both programmes and the derivation of key design criteria.

16.1.1

Samples

Samples of core were received from a total of 14 diamond drill holes and were designated as follows:

Table 16.1    Testwork Samples

Hole	Meterage (m)	Oxidation Zone
KEDD 291	9-16	Oxide
KEDD 291	42-50	Oxide
KEDD 297	33-41	Oxide
KEDD 297	84-91	Oxide
KEDD 6011C	3.53-9	Oxide
KEDD 6011C	31-35	Oxide
KEDD 6016	31-39	Oxide
KEDD 297	118-126	Transition
KEDD 297	143-150	Transition
KEDD 152	283-290	Fresh
KEDD 152	348-355	Fresh
KEDD 161	281-288	Fresh
KEDD 161	303-309	Fresh
KEDD 199	181-188	Fresh
KEDD 231	166-172	Fresh
KEDD 231	185-192	Fresh
KEDD 320	224-240	Fresh
KEDD 336	329-336	Fresh

These samples formed the feed to both the early testwork in 2008 and the more recent testwork in 2009. The early testwork focussed on testing of individual intercepts to determine metallurgical response. The recent testwork was based on treatment of composites to determine a likely flowsheet. These testwork programmes are summarised separately in the following sections.

16.1.2

Initial Program

In July 2008, Keegan commissioned Lycopodium to provide assistance in assessing alternative process routes and the high level economics of the project. Lycopodium managed a preliminary metallurgical test program, which was completed by Amdel Mineral Laboratories Ltd in Perth, Western Australia.

The variability testwork was designed to be diagnostic in nature, rather than predictive of final plant design conditions. Each sample had a grind time determination, followed by grinding to 80% -45 N and cyanidation. The fine grind size was intended to produce results indicative of the potential maximum recovery, including possible regrinding of a flotation concentrate.

The first stage of the test program involved selection and testing of a range of HQ core samples, intended to:

·

Gain an understanding of the range of metallurgical response from the mineralisation types at Esaase.

·

Determine how many composite samples would need to be tested to establish preliminary metallurgical design parameters.

·

Identify any issues requiring further investigation, including refractory gold in sulphides,
presence of preg-robbing carbonaceous material and the presence of coarse gold.

The half HQ diamond core intervals for the variability testwork were despatched from Accra on 21 November 2008.

16.1.3

Main Findings

•

The variability samples selected provide good coverage of the current mineralisation spatially and in respect of parameters such as grade, oxidation state and sulphide content.

•

Strongly oxidised ores averaged 91% gold extraction in agitation leaching and 87% in coarse bottle rolls (indicative of heap leaching), although the latter tests were on a smaller sample set. Cyanide consumed in agitation leaching was 0.7 kg/t, but only 0.2 kg/t in the bottle rolls.

•

Moderately to weakly oxidised material averaged 71% extraction in agitation leaching, but were highly variable. Bottle roll extractions on the same samples averaged 65%. Cyanide consumption in agitation leaching was high, at 1.6 kg/t, but only 0.3 kg/t in the bottle rolls.

•

Fresh rock averaged 54% extraction in agitation leaching, but varied from 15 - 91 %. No bottle rolls were done on fresh material. Cyanide consumption was high at 1 kg/t, but lime addition rates were low at 0.4 kg/t.

•

Approximately 30% of the gold detected in the samples tested was coarser than 106 N. Possibly a total of up to 70% of the gold may be close to or coarser than 106 N in size and a further 20% accessible at sizes coarser than 45 N.

•

The presence of coarse gold resulted in reduced extraction which is assumed to be simply from the coarser grain size.

•

For all samples, increased manganese concentrations correlated with reduced gold extraction rates and/or reduced final extraction. The reason for the correlation is unclear, but the effect may warrant further work.

•

There is no evidence that elements such as tellurium, copper or antimony are affecting gold extraction. Other complications such as high mercury or silver concentrations are also absent.

•

All samples contained organic carbon (TOC). TOC concentrations varied with oxidation state, with strongly oxidised material containing on average 0.06% TOC, moderate to weakly oxidised 0.2% and fresh 0.35%. Gold extraction by coarse bottle rolls (10 day duration) showed a very strong relationship with TOC concentrations, declining from 90% extraction at 0.05% TOC to 60% at 0.3% TOC.

•

In the much more rapid agitation leach tests, there was no detectable preg-robbing effect on oxide and transition material. Only one fresh sample showed direct evidence of preg­robbing, in the form of a declining leach curve and a very low final extraction (19%). However, one other sample with even lower final gold extraction (15%) had the highest TOC concentration, at 0.56%. It is suspected that the presence of slowly leaching coarse gold both contributed to the low extraction and masked the effects of preg-robbing.

•

Evidence to date suggests that refractory gold in sulphides may not be a major problem at Esaase. Pyrite concentrations are relatively low, with none higher than 1%. Calculated arsenopyrite concentrations were quite high, ranging from <0.1 to 2.8%, but showed no relationship to gold extractions. Comparing size by size recoveries on oxide and fresh material suggests less than 5% of gold is finely occluded in sulphides. However, further mineralogical investigations are required to confirm this.

•

The lower gold extraction values achieved in some transition and fresh material show that Esaase has some metallurgical variability. The evidence to date points to the presence of coarse gold being the main cause of low extraction, rather than gold in sulphides. If this proves to be the case, there are good prospects for significant improvements in gold recovery if utilising continuous processing (gravity concentration, flotation and/or CIL). The prospects for heap leaching on fresh and transition material would be poor.

•

The best available predictor of potential heap leach recovery within the geological database was the logged (not modelled) oxidation state of the material, with strongly oxidised material having the best bottle roll extractions. This appears to be due to the low concentrations of carbonaceous material (indicated by low TOC).

•

Bulk density and rock porosity measurements can be used to distinguish fresh material from oxidised, but do not otherwise correlate with bottle roll extraction.

•

Samples modelled into the transition zone could not be uniquely distinguished from oxide material by any measured parameter, although on average their metallurgical performance lay between that of strongly oxidised and fresh material.

•

Gold extractions do not correlate with head grade. Residue grades varied rather than being constant. Thus, the results of metallurgical testwork are unlikely to be much affected by the grade of the samples, within reasonable limits.

•

Gravity / flotation / concentrate and tails leaching of the fresh material gave extractions of 92% at a grind of 150 microns. This improved to 97% at a grind of 75 microns.

•

Intermittent coarse bottle rolls on oxide material crushed to 22, 12 and 6 mm showed extractions of 78, 77 and 86% after 10 days. A follow up column leach crushed to 12 mm and agglomerated with 4 kg/t cement gave an extraction of 77.7% after 64 days. At this point the leach curve was still rising. Overall cyanide consumption was low at 0.19 kg/t.

•

Mineralogical investigation of an oxide and fresh sample showed limited results as the gold occurrences were rare. Most of the gold by volume which was observed was coarse.

•

All fresh samples had similar grindabilities, but oxide and transition samples were much more variable. The hardest oxide samples required similar energy to the hardest fresh samples. Although not quantifiable, it is likely that the presence of vein quartz influenced grinding energy in oxide and transition ores. Comminution testwork gave the following results on specific intercepts:

Table 16.2                      Bond Index Determination

	Bond Ball Work Index kWh/t	Bond Rod Work Index kWh/t	Bond Abrasion Index g
KEDD 510 (Fresh) KEDD 511+521 (oxide)	14.8 9.2	17.5 8.8	0.2557 0.0839

Table 16.3             SMC Test Results

	A x b		t 10 @ 1 kWh/t
	Value	Category	Value	Category
KEDD 510 (Fresh) KEDD 511+521 (oxide) Oxide Master Composite	38.7 93.9 141.5	Mod hard Soft Very soft	28.2 48.6 55.0	Mod hard Soft Very soft

The two test programmes are consistent and show that the fresh material is moderately hard with energy for grinding in the mid range. The oxide material is soft with a low energy demand.

In summary, the presence of coarse gold indicates that a gravity concentration stage is warranted and this may even out some of the variability in agitated leaching which showed in the testwork.

16.1.4

Recent Testwork

The recent testwork program completed at Amdel in 2009 used the same sample sources as those reported above. However, the 2009 programme was designed to develop a process flowsheet at the scoping level which could be used as a basis for completing a preliminary economic assessment. Further work would be required to characterise the metallurgical performance of the material types intended to be processed. As a result two major composites, variability fresh composite (VFC) and variability oxide composite (VOC) were prepared to represent the overall mineralised zones intended for treatment.

Table 16.4         VFC Composite

Composite	Mass to VFC Composite (kg)
KEDD 231 185-192 m	8.5
KEDD 199 181-188 m	5.6
KEDD 152 283-290 m	9.7
KEDD 320 224-240 m	9.5
KEDD 152 348-355 m	6.8
KEDD 161 281-288 m	8.1
KEDD 161 303-309 m	3.9
KEDD 336 329-336 m	4.4
KEDD 510 124.5-142.5 m	9.0
KEDD 522 77.6-108.5 m	15.5
Total	81.0

Table 16.5     VOC Composite

Composite	Mass to VOC Composite (kg)
KEDD 291 9-16 m	5.0
KEDD 6016 31-39 m	10.1
KEDD 511 32.5-49.5 m	5.0
KEDD 521 16.0-58.7 m	13.0
Total	33.0

From there, testwork was designed to address comminution, gravity separation and CIL leaching. From this, recovery predictions and reagent consumption was developed. With the exception of significant levels of total organic carbon and arsenic in the VFC composite both of the testwork composites were low in deleterious elements.

16.1.5

Comminution Testwork

Following the preliminary comminution testwork completed in the previous testwork, a series of work index determinations was completed on the two composites above. Results are presented in Table 16.6 below.

Table 16.6       Bond Index Determination

	Bond Ball Work Index kWh/t	Bond Rod Work Index kWh/t
VFC VOC	15.3 9.1	19.3 NA

Results are consistent with the previous work.

16.1.6

Coarse Ore Bottle Roll Testwork

Coarse ore bottle roll leach tests on composites VFC and VOC produced mixed results with a low recovery for the VFC composite, of 24.3%, and a moderate recovery for the VOC composite, of 73.9%. Reagent consumptions were moderate for both samples.

Table 16.7       Coarse Bottle Roll Testwork

Composite	Reagent Consumption		Au Recovery (%)	Au Calc Head Grade (g/t)	Au Residue Grade (g/t)
	NaCN (kg/t)	Lime (kg/t)
VFC VOC	0.53 0.56	0.38 0.75	24.3 73.9	1.41 0.991	1.07 0.259

Results indicate that while heap leaching may be possible on oxide material it is not possible on fresh rock.

16.1.7

Gravity / Leach Testwork

Gravity / CIL leach testwork was conducted on both composites at 150 and 75 micron grind sizes. Results at 150 micron showed 94% extraction for the VOC composite and 95% extraction for the VFC composite. These extractions improved somewhat with a finer grind.

The calculated and screen fire assay heads show some variation in head grade which appears to be due to the presence of coarse gold. All CIL residue grades were determined by duplicate 1,000 gram screen fire assay procedures.

Table 16.8        Gravity / Leach Testwork Overall Results on VOC Composite Gravity Tails

Fraction	AL 38 (150 p)		AL 39 (150 p)		AL 40 (75 p)		AL 41 (75 p)
	Au Grade (g/t)	Au Dist (%)	Au Grade (g/t)	Au Dist (%)	Au Grade (g/t)	Au Dist (%)	Au Grade (g/t)	Au Dist (%)
Gravity Concentrate + Leached Gravity Tail Un Leached Gravity Tail	1.16 0.073	94.1 5.9	1.33 0.065	95.4 4.6	1.11 0.034	97.1 2.9	1.17 0.040	96.7 3.3
Total	1.23	100.0	1.39	100.0	1.14	100.0	1.21	100.0
Head Assay - Screen Fire Assay	0.95		0.95		0.95		0.95

Table 16.9          Gravity / Leach Testwork Overall Results on VFC Composite Gravity tails

Fraction	AL 38 (150 p)		AL 39 (150 p)		AL 40 (75 p)		AL 41 (75 p)
	Au Grade (g/t)	Au Dist (%)	Au Grade (g/t)	Au Dist (%)	Au Grade (g/t)	Au Dist (%)	Au Grade (g/t)	Au Dist (%)
Gravity Concentrate + Leached Gravity Tail Un Leached Gravity Tail	1.78 0.087	95.4 4.6	1.80 0.085	95.6 4.4	2.23 0.074	96.8 3.2	2.26 0.066	97.2 2.8
Total	1.86	100.0	1.88	100.0	2.30	100.0	2.33	100.0
Head Assay - Screen Fire Assay	1.55		1.55		1.55		1.55

16.1.8

Leach Testwork Reagent Consumption

Leach test analysis of the VOC composite gravity tails at P80 75 and 150 pm produced residue grades of below 0.10 g/t, with moderate consumptions of both NaCN and lime. Leach residue grades were approximately 0.03 g/t lower in the 75 pm leach tail.

Leach test analysis of the VFC composite gravity tails at P80 75 and 150 pm also produced residue grades of below 0.10 g/t, with moderate consumptions of NaCN and lime. Leach residue grades were approximately 0.015 g/t lower in the 75 pm leach tail.

Table 16.10

     Leach Residue and Reagent Consumption on VOC Composite Gravity Tails

Grind Size (Pea, N)	Reagent Consumption		Au Recovery (%)	Au Calc Head Grade (g/t)	Au Residue Grade (g/t)
	NaCN (kg/t)	Lime (kg/t)
150	0.45	0.73	89.5	0.692	0.073
150	0.46	0.65	92.5	0.860	0.065
75	0.54	0.71	92.5	0.456	0.034
75	0.40	0.71	92.4	0.526	0.040

Table 16.11    Leach Residue and Reagent Consumption on VFC Composite Gravity Tails

Grind Size (Pis, II)	Reagent Consumption		Au Recovery (%)	Au Calc Head Grade (g/t)	Au Residue Grade (g/t)
	NaCN (kg/t)	Lime (kg/t)
150	0.35	0.36	84.6	0.564	0.087
150	0.38	0.31	85.4	0.580	0.085
75	0.38	0.32	87.3	0.579	0.074
75	0.36	0.31	89.1	0.604	0.066

The above data has formed the basis for the design criteria for the preliminary economic assessment. Further work will be required to fully characterise and optimise the metallurgical performance of the mineralisation types intended to be processed.

16.1.9

Interpretation of Results

On the basis of the 2008 testwork, it was determined that a flowsheet comprising gravity separation and CIL processing was most likely to offer a cost effective processing solution. As a result, the 2009 programme targeted this process flowsheet. The key design parameters to be determined were grind size, likely gravity recovery and CIL residence time.

Comparative gravity / carbon in leach testwork at 80% passing 150 and 75 microns showed a recovery improvement of 0.5 to 1% for the finer grind. A grind optimisation study which considered the incremental cost benefit of the finer grind indicated a better return with a grind of 150 micron. This was adopted for the study.

Testwork on the two composites (VOC and VFC) has shown gravity gold recovery of 49% for the VOC at 150 micron and 73% for the VFC at 150 micron based on gravity testing of a 9 kg sample. Typical industry standards allow for a 30% bleed of circulating load to be sufficient to capture the vast majority of gravity recoverable gold. As Esaase has a higher than average gravity component, the design parameter has been increased to 35%.

A leach time of 48 hours without sub-sampling was used for the purpose of determining the CIL extraction from gravity tail for the recent testwork. This was also used for the previous testwork. This was intended to obtain an ultimate recovery for the purpose of a gold balance across the circuit. For the purpose of this study, a leach time of 24 hours has been used which is typical for the industry. This will be optimised during the next stage of testwork.

16.1.10

Performance Predictions

Gold recovery predictions are based on the testwork on the VOC and VFC samples only and have used the following methodology:

·

Head grade has been based on a weighted average of all the screen fire assay results for the composite in question.

·

Gravity extraction is assumed as 40% for the calculation with a concentrate leach extraction of 97% based on the master concentrate leach at 150 microns.

·

CIL extraction has been based on the average of the two duplicate 2 kg leach tests at 150 microns with leach residue grades calculated by duplicate 1,000 gram screen fire assay procedures.

·

Soluble loss has been assumed at 0.01 g/t.

·

Recovery is based on the calculated mass of gold extracted at each stage as follows:

·

gold in conc. leach solution + gold in CIL feed - gold in residue - solution loss.

The result for the two main mineralisation types is as follows:

Fresh material

94.4%

Oxide material

93.3%

16.1.11

Suitability of Available Testwork

It is the opinion of Lycopodium that the available testwork is adequate to support the findings of the Preliminary Economic Assessment as summarised in this report.

16.2

Process Description

16.2.1

Summary Design Criteria

Based on the testwork a preliminary process design criteria was developed. Table 16.12 below summarises the key design criteria for the study.

Table 16.12          Summary Process Design Criteria

Parameter	Units	Value	Source
Annual throughput	tpa	5,000,000 fresh ore	Client/Consultant
Operating hours	h	8,000	Lyco/typical
Head grade	g Au/t	1.60	Client
Primary grind P80	microns	150	Testwork
Gravity recovery	%	35	Lyco/testwork
Leach residence time	h	24	Lyco/typical
CIL recovery	%	95	Testwork
Overall recovery	%	96.8	Calc

16.2.2

Process Overview

The Esaase process plant will consist of crushing, ore stockpiling and reclaim, grinding incorporating gravity recovery of coarse gold, thickening, carbon in leach (CIL), elution and gold recovery, and cyanide destruction circuits, with associated services and ancillaries, rated to treat 6.5 million tonnes per annum of oxide and 5 million tonnes per annum of fresh open pit material and recover 200,000 ounces of gold as dore bars ready for shipment to a refinery.

The processing facility is based on a simple, robust flowsheet providing a modest capital cost for rated capacity without excessive operating cost penalty, and with a design life of 10+ years.

The conceptual plant design will, when detailed, meet the requirements of the International Cyanide Management Code (ICMC).

The process flowsheet is based on a single stage crush, SAG mill and ball mill circuit. This is a conservative approach taken in the absence of comprehensive comminution testwork data. For the relatively coarse grind required by the Esaase ores a single stage mill may be appropriate, and will be investigated at a later study stage, as it offers some saving in capital. The SAG mill circuit incorporates recycle crushing. This has been incorporated as deferred capital as it has been assumed it will not be required in the first two years of operations when mill feed will be predominantly softer oxide ores.

In the absence of a full suite of testwork data the plant design adopted for the Preliminary Economic Assessment is based on a number of assumptions regarding the metallurgical performance of the Esaase ores, assumptions of this nature are consistent with the level of definition available at this stage of project development.

16.2.3

Primary Crushing, Ore Stockpile and Reclaim

ROM ore will be fed directly into the feed chamber of a 42 - 65 gyratory crusher. Crusher product will be drawn from the surge bin beneath the crusher by an apron feeder and conveyed to a coarse ore stockpile (COS) with a nominal 10,000 tonne live capacity.

Ore will be reclaimed from the stockpile by three apron feeders in a concrete reclaim chamber below the stockpile. Ore from the reclaim feeders will be conveyed to the SAG mill. At any one time two or three of the reclaim feeders will be in operation to mitigate the impact of size segregation on the stockpile on the downstream milling operation.

The crushing plant will operate on both day and night shifts. During periods when the crusher is not available ore for up to nominally 16 hours of mill operation can be recovered from the live portion of stockpile. Additional ore can be reclaimed from the dead portion of the stockpile using bulldozers.

16.2.4

Grinding and Classification

The grinding circuit will consist of a SAG and ball mill in closed circuit with a cyclone cluster providing a grind of 80% passing 150 microns.

Preliminary calculations, in the absence of comprehensive comminution testwork data, indicate mill installed power requirement of 7 MW for the SAG mill and 5.5 MW for the ball mill. In practice the power split may be changed to an installed 6 or 6.5 MW per mill to provide commonality of drive trains.

The coarse grind may be amenable to the use of a single stage milling circuit to reduce capital cost. This will be investigated when more detailed information is available.

16.2.5

Gravity Concentration

A portion of the classification cyclone feed will be bled from the cyclone cluster and fed to a gravity concentration circuit to recover coarse free gold. The circuit will consist of a pair of scalping screens and two 48” centrifugal concentrators. Gravity gold concentrate will be leached in an intensive leach reactor with the gold bearing pregnant solution recovered in a dedicated electrowinning cell.

16.2.6

Pre-leach Thickener

Classification cyclone overflow will gravitate to the pre-leach thickener via trash removal screens. The ore slurry will be thickened to nominally 50% solids before feeding the leach train. In practice it may not be possible to thicken the oxide ore slurry to 50% solids. This will decrease the residence time in the leach circuit. Further optimisation of thickener performance and leach residence times will be undertaken when test work data is available.

16.2.7

Carbon in Leach (CIL)

Pre-leach thickener underflow will be leached in a train of seven 3,250 m³ CIL tanks providing a nominal residence time of 26 hours at a mill feed rate of 5 million tonnes per annum. The lower residence time when treating a higher tonnage of oxide ore will be compensated for by faster leach kinetics.

To meet the requirements of the low head grade to the leach circuit the concentration of carbon in the CIL tanks will be nominally 10 grams per litre with an advance rate of 8 tonnes of carbon per day for a maximum carbon loading of approximately 2,500 grams of gold per tonne.

16.2.8

Elution and Gold Recovery

Gold will be stripped from the loaded carbon in an 8 tonne capacity split Anglo elution circuit and electrowon onto stainless steel cathodes. The capacity of the elution circuit is such that a single strip per day is required. This provides some “catch up” capacity by adopting around the clock stripping if required.

Barren carbon is regenerated in a horizontal kiln before returning to the CIL circuit.

Gold recovered from the stainless steel cathodes is smelted to doré bars for shipment to a refinery. The electrowinning cells, smelting furnace and associated equipment are located in a secure gold room with appropriate access control and security features such as closed circuit TV monitoring and intruder detection and alarms.

16.2.9

Cyanide Detoxification

A sulphur dioxide / air detoxification circuit will be provided to reduce WAD cyanide levels in plant tailings to below those required under the ICMC. Sodium metabisulphite will be the source of sulphur dioxide for the reactors.

16.2.10

Services and Water

Reagents

All reagents required for the operation are currently being used in commercial quantities by other mining operations in the region.

Raw Water

Raw water will be delivered to the plant site by pipeline from the pit dewatering pumps and borefield and stored on site in a HDPE lined dam.

Process Water

Process water will be supplied from the tailings storage facility decant supplemented by raw water if necessary. Process water will be stored on site in a HDPE lined dam.

17.0

MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

Coffey Mining has estimated the Mineral Resource for the Esaase Gold Project as at 28 February 2009. All grade estimation was completed using Multiple Indicator Kriging (‘MIK’) for gold. This estimation approach was considered appropriate based on review of a number of factors, including the quantity and spacing of available data, the interpreted controls on mineralisation, and the style of mineralisation. The estimation was constrained with geological and mineralisation interpretations.

17.1

Database Validation

The resource estimation was based on the available exploration drillhole database which was compiled by Coffey Mining. The database has been reviewed and validated by Coffey Mining prior to commencing the resource estimation study.

Data included samples from extensive trenching, but only the RC and diamond drilling sample data were included for use in the modelling process. A total of 467 RC and diamond drillholes were used in the resource modelling study.

The database was validated in Micromine software and the checks made to the database prior to loading into Vulcan included:

·

No overlapping intervals.

·

Downhole surveys at 0 m depth.

·

Consistency of depths between different data tables.

·

Check gaps in the data.

A total number of 145 samples from 6 sample batches were destroyed during a fire at the SGS laboratory in Tarkwa. These samples have been replaced by -999 in the database. Other changes that were made to the database prior to loading into Vulcan included:

·

Replacing less than detection samples with half detection.

·

Replacing intervals with no sample with -999.

·

Replacing intervals with assays not yet received with -999.

The resource dataset has been described in Section 11. In summary, samples were composited to 3 m down-hole lengths with residual intervals less than 1.5 m length being deleted from the composite file. Prior to deletion of composites less than 1.5 m, statistical analysis was undertaken to determine the impact on mean gold grades. Deletion of these composites was deemed to have negligible impact on mean grades and was therefore appropriate. The resulting file contained 13,496 composites with gold grades within mineralised domains.

17.2

Geological Interpretation and Modelling

Based on grade information and geological observations, oxidation and mineralised domain boundaries have been interpreted and wireframes modelled to constrain resource estimation for the Esaase deposit. Interpretation and digitising of all constraining boundaries has been undertaken on cross sections orientated at 100º (drill line orientation). The resultant digitised boundaries have been used to construct wireframe surfaces or solids defining the three-dimensional geometry of each interpreted feature. The interpretation and wireframe models have been developed using the Vulcan mine planning software package.

17.2.1

Mineralisation Interpretation

For the purpose of resource estimation, three mineralised domains were interpreted and were modelled on a lower cut-off grade of 0.3 g/t Au. The domains are listed below and depicted in Figure 17.1 and Figure 17.2.

·

Footwall Domain: Designated Zone 100 (designated 1 in the previous resource estimate). A moderately to steeply dipping zone hosting the bulk of the mineralisation and entirely contained within the sedimentary sequence. This domain dips more steeply towards the north and is depicted in Figure 17.1 on the right.

·

Hangingwall Domain: Designated Zone 150 (designated 2 in the previous resource estimate). A parallel to sub-parallel zone of mineralisation, structurally higher than the footwall domain and depicted on the left in Figure 17.1.

·

South Domain: Designated Zone 200. Previously undefined mineralisation to the south of the previous two domains and depicted on the left in Figure 17.1.

Figure 17.1                      Mineralisation Interpretation SE Oblique View

Figure 17.2               Mineralisation Interpretation Typical Sectional View

17.2.2

Weathering Interpretation

Composites were also coded by the weathering profile. The profile has been modelled from drill data and comprises of strongly weathered saprolite, moderately weathered saprock, transition material and fresh units. In general, weathering surfaces broadly parallel the topographical profile, although weathering tends to be deeper within zones of mineralisation and tends to parallel the footwall to the mineralisation where the footwall approaches the surface. On some sections, the intermixing of the weathering types can be quite complicated. All of the weathering surfaces have been utilised in terms of dividing the gold mineralisation into secondary domaining for statistical analysis and incorporated into the resource model for the purposes of assigning tonnage factors. Figure 17.3 is an example section (at 9,840mN, local grid) showing the distribution of weathering types and the interpreted position of the top of fresh rock.

Figure 17.3                  Weathering Interpretation, Local Grid 9,840mN

17.3

Statistical Analysis

The lengths of the samples were statistically assessed prior to selecting an appropriate composite length for undertaking statistical analyses, variography and grade estimation. Summary statistics of the sample length indicates that 94.3% of the samples were collected at 1 m intervals, 2.8% was collected at 2 m intervals and the remainder (2.8%) was sampled at irregular intervals less than 3 m.

Statistical analysis was undertaken based on 3 m composites of the gold assay data for the resource dataset drilling completed at Esaase. All composites inside the wireframes were flagged as separate domains. A total of 13,496 composites were used in the modelling process from a total of 467 RC and diamond drillholes.

Summary statistics were generated to compare assayed RC samples and DC samples. Only assays with values greater than 0.3 g/t Au were considered. These are presented in Table 17.1. The means of two types of sampling are similar with the medians being equivalent. Differences may be explained by the effect of high grade outliers.

Separate statistics were generated for each domain. The data was further subdivided, and flagged, into sub-domains based on weathering profile. Summary statistics for each modelled domain are presented in Table 17.2.

Table 17.1       RC vs DC Summary Statistics

Item	RC	DC
Count	3,989	1,858
Minimum	0.303	0.301
Maximum	75.815	42.290
Mean	1.5	1.435
Median	0.773	0.745
Standard Deviation	3.085	2.742
Variance	9.516	7.519
CV	2.057	1.91

Table 17.2       Esaase Gold Deposit Domain Composite Statistics (Au g/t)

Domain	Sub-Domain	N	Min	Max	Mean	Median	Std Dev	variance	CV
	Strongly Oxidised	610	0.005	14.083	0.87	0.44	1.475	2.176	1.695
	Moderately Oxidised	949	0.005	75.813	0.923	0.283	3.263	10.646	3.536
Zone 100	Transition	771	0.005	47.387	1.031	0.307	3.192	10.188	3.095
	Fresh	3689	0.005	34.61	0.671	0.25	1.499	2.247	2.235
	All	6,019	0.005	75.813	0.777	0.277	2.144	4.596	2.76
	Strongly Oxidised	752	0.005	71.643	0.689	0.193	2.975	8.849	4.319
	Moderately Oxidised	1460	0.005	35.568	0.609	0.2	1.557	2.424	2.558
Zone 150	Transition	957	0.005	38.6	0.582	0.17	1.919	3.683	3.299
	Fresh	2281	0.005	34.377	0.471	0.13	1.499	2.247	3.182
	All	5,450	0.005	71.643	0.557	0.163	1.859	3.457	3.337
	Strongly Oxidised	123	0.018	36.487	1.325	0.435	3.664	13.428	2.766
	Moderately Oxidised	539	0.005	35.717	0.768	0.25	2.298	5.282	2.992
Zone 200	Transition	154	0.007	14.51	0.824	0.313	1.619	2.622	1.965
	Fresh	1211	0.005	42.29	0.826	0.25	2.349	5.518	2.844
	All	2,027	0.005	42.29	0.841	0.27	2.391	5.718	2.844
All Domains		13,496	0.005	75.813	0.698	0.223	2.078	4.319	2.978

Figure 17.4 to 17.4.3 shows log histograms and probability plots of gold grades. Populations of gold grades are close to lognormal and show strong positive skewness for both domains and this is typical of many gold deposits. The coefficients of variation (`CV') are moderately high indicating that it may be difficult to maintain a high degree of selectivity in mining.

Bulk density determinations were coded by weathering interpretation in the database and density values for the weathering subdivisions were subsequently extracted from the database. Histograms, log histograms and probability plots were generated and examined. Samples exist in the database that pertains to areas outside of the resource area and these have been excluded prior to examination of the data. Summary statistics are presented in Table 17.3 below.

Table 17.3            Esaase Gold Deposit Density Statistics (t/m³)

	Strongly Oxidised	Weakly Oxidised	Transition	Fresh
Number	166	235	178	5,286
Minimum	1.66	1.39	1.38	1.07
Maximum	3.01	2.95	3.17	4.12
Mean	2.312	2.416	2.51	2.75
Std Dev	0.174	0.161	0.29	0.15
Variance	0.03	0.03	0.08	0.02
Coeff Var	0.075	0.067	0.114	0.055

Conditional statistics for data within each domain to be estimated by Multiple Indicator Kriging are listed in Table 17.4.

Figure 17.4         Log Histogram and Probability Plot Zone 100

Figure 17.5        Log Histogram and Probability Plot Zone 150

Figure 17.6          Log Histogram and Probability Plot Zone 200

Table 17.4         Esaase Gold Deposit Indicator Class Means

Domain
Zone 100 Zone 150
Probability Threshold	Grade Threshold	Class Mean	Probability Threshold	Grade Threshold	Class Mean
0.362	0.15	0.058	0.479	0.15	0.054
0.472	0.25	0.194	0.593	0.25	0.195
0.587	0.40	0.318	0.703	0.40	0.317
0.670	0.55	0.468	0.766	0.55	0.468
0.727	0.70	0.616	0.817	0.70	0.62
0.770	0.85	0.771	0.850	0.85	0.77
0.804	1.00	0.92	0.877	1.00	0.912
0.844	1.20	1.092	0.904	1.20	1.094
0.885	1.50	1.333	0.927	1.50	1.341
0.908	1.75	1.61	0.941	1.75	1.615
0.931	2.10	1.899	0.953	2.10	1.91
0.945	2.48	2.266	0.962	2.48	2.3
0.958	2.99	2.705	0.971	2.99	2.682
0.971	4.12	3.516	0.982	4.12	3.413
0.980	5.35	4.78	0.989	5.35	4.646
0.991	8.85	6.607	0.994	8.85	6.777
Max	Max	12.562	Max	Max	13.611
Zone 200
Probability	Grade
Threshold	Threshold	Class Mean
0.369	0.15	0.056
0.482	0.25	0.195
0.596	0.40	0.32
0.679	0.55	0.469
0.742	0.70	0.619
0.781	0.85	0.766
0.815	1.00	0.917
0.842	1.20	1.094
0.884	1.50	1.358
0.902	1.75	1.641
0.921	2.10	1.9
0.937	2.48	2.289
0.951	2.99	2.711
0.968	4.12	3.575
0.978	5.35	4.696
0.988	8.85	6.928
Max	Max	13.749

17.4

Variography

17.4.1

Introduction

Variography is used to describe the spatial variability or correlation of an attribute (gold, silver etc.). The spatial variability is traditionally measured by means of a variogram, which is generated by determining the averaged squared difference of data points at a nominated distance (h), or lag (Srivastava and Isaacs, 1989). The averaged squared difference (variogram or y(h)) for each lag distance is plotted on a bivariate plot, where the X-axis is the lag distance and the Y-axis represents the average squared differences (y(h)) for the nominated lag distance.

Several types of variogram calculations are employed to determine the directions of the continuity of the mineralisation:

·

Traditional variograms are calculated from the raw assay values.

·

Log-transformed variography involves a logarithmic transformation of the assay data.

·

Gaussian variograms are based on the results after declustering and a transformation to a Normal distribution.

·

Pairwise-relative variograms attempt to 'normalise' the variogram by dividing the variogram value for each pair by their squared mean value.

·

Correlograms are 'standardized' by the variance calculated from the sample values that contribute to each lag.

Fan variography involves the graphical representation of spatial trends by calculating a range of variograms in a selected plane and contouring the variogram values. The result is a contour map of the grade continuity within the domain.

The variography was calculated and modelled in the geostatistical software, Isatis. The rotations are tabulated as input into Isatis (geological convention), with X representing rotation around Z axis, Y representing rotation around Y' axis and Z representing rotation around X". Dip and dip direction of major, semi-major and minor axes of continuity are also referred to in the text. Modelled correlograms were generally shown to have good structure and were used throughout.

17.5

Esaase Deposit Variography

Grade and indicator variography was generated to enable grade estimation via MIK and change of support analysis to be completed. In addition, Gaussian variograms were also used as part of the change of support process. Seven indicator thresholds were investigated for each domain. Interpreted anisotropy directions correspond well with the modelled geology and overall geometry of the interpreted domains.

17.5.1

Zone 100

Grade variography shows good structure and displays moderate anisotropy between the major and semi-major axes. Two spherical models have been fitted to the experimental correlogram, with the correlogram exhibiting a high relative nugget effect (calculated by dividing the nugget variance by the sill variance) of 40%. The short-range structure, which has been modelled with ranges of 30 m, 15 m and 4 m for the major, semi-major and minor axis respectively, accounts for 75% of the non-nugget variance. The overall ranges fitted to the Zone 1 correlogram are 90 m, 55 m and 13 m for the major, semi-major, and minor axis respectively.

The interpreted major direction of continuity dips at 9° towards 036°30'. The modelled grade variogram plot is provided in Figure 17.7.

Modelled indicator correlograms display a range of relative nugget values from 36% to 65% and this is broadly comparable with the grade variogram nugget of 40%. Table 17.5 presents the fitted grade and indicator variogram models for Zone 100.

Table 17.5       Esaase Deposit Zone 100 Correlogram Models

Grade Variable or Indicator Threshold		Nugget (CO)	Rotation (Isatis)			Structure 1				Structure 2
						Sill 1 (C1)	Range (m)			Sill 2 (C2)	Range (m)
			Z	Y	X		Major	Semi Major	Minor		Major	Semi Major	Minor
Grade Variography
Gold (Au g/t)		0.4	40	-70	-10	10.45	30	15	4	10.15	90	55	13
	Indicator Variography
0.15(1)		0.36	40	-70	-10	0.42	26	19	10	0.22	130	85	34
0.25(1)		0.38	40	-70	-10	0.42	25	18	9	0.2	125	80	32
0.40		0.4	40	-70	-10	0.4	24	17	8	0.2	120	75	30
0.55(2)		0.45	40	-70	-10	0.385	23	17	7.5	0.165	115	75	27.5
0.70		0.5	40	-70	-10	0.37	22	17	7	0.13	110	75	25
0.85(3)		0.525	40	-70	-10	0.345	21	16	6.5	0.13	107.5	72.5	23.5
1.00		0.55	40	-70	-10	0.32	20	15	6	0.13	105	70	22
1.20(4)		0.565	40	-70	-10	0.31	19	14	5.5	0.125	102.5	67.5	16.5
1.50		0.58	40	-70	-10	0.3	18	13	5	0.12	100	65	11
1.75(5)		0.59	40	-70	-10	0.29	18	12.5	4.5	0.12	97.5	62.5	11
2.10		0.6	40	-70	-10	0.28	18	12	4	0.12	95	60	11
2.48(6)		0.61	40	-70	-10	0.28	16.5	11.5	4	0.11	87.5	57.5	10.5
2.99		0.62	40	-70	-10	0.28	15	11	4	0.1	80	55	10
4.12(7)		0.625	40	-70	-10	0.275	14	10.5	4	0.1	67.5	45	9
5.35		0.63	40	-70	-10	0.27	13	10	4	0.1	55	35	8
8.65)8)		0.65	40	-70	-10	0.25	12	9	3	0.1	40	25	5

Note:

1)  Assumed model based on 0.40 Au g/t variogram model

2)

Assumed model based on 0.40 Au g/t and 0.70 Au g/t variogram models

3)

Assumed model based on 0.70 Au g/t and 1.00 Au g/t variogram models

4)

Assumed model based on 1.00 Au g/t and 1.50 Au g/t variogram model

5)

Assumed model based on 1.50 Au g/t and 2.10 Au g/t variogram model

6)

Assumed model based on 2.10 Au g/t and 2.99 Au g/t variogram models

7)

Assumed model based on 2.99 Au g/t and 5.35 Au g/t variogram models

8)

Assumed model based on 5.35 Au g/t variogram model

Figure 17.7      Zone 100 Grade Variogram (Corellogram)

17.5.2

Zone 150

Grade variography shows good structure and displays moderate anisotropy between the major and semi-major axes. Two spherical models have been fitted to the experimental correlogram, with the correlogram exhibiting a moderate relative nugget effect of 40%. The short range spherical model has been fitted with minor ranges of 40 m, 25 m and 4 m. A second spherical model has been fitted with overall ranges of 90 m, 60 m and 23 m for the major, semi-major and minor axis respectively. The short range structure accounts for three quarters of the non nugget variance.

The interpreted major direction of continuity dips at 7º30’ towards 38º30’. Table 17.6 presents the fitted grade variogram and indicator variogram models for Zone 150.

Table 17.6         Esaase Deposit Zone 150 Correlogram Models

Grade Variable or Indicator Threshold		Nugget (C0)	Rotation (Isatis)			Structure 1				Structure 2
						Sill 1 (C1)	Range (m)			Sill 2 (C2)	Range (m)
			Z	Y	X		Major	Semi Major	Minor		Major	Semi Major	Minor
Grade Variography
Gold (Au g/t)		0.4	45	-50	-10	0.45	40	25	4	0.15	90	60	23
	Indicator Variography
0.10(1)		0.38	45	-50	-10	0.37	44	27	8	0.25	100	60	29
0.20(1)		0.39	45	-50	-10	0.36	42	26	7	0.25	95	55	27
0.30		0.4	45	-50	-10	0.35	40	25	5	0.25	90	50	25
0.40(2)		0.44	45	-50	-10	0.35	40	22.5	5	0.21	90	42.5	23.5
0.50		0.48	45	-50	-10	0.35	40	20	5	0.17	90	35	22
0.60(3)		0.495	45	-50	-10	0.335	39	19	5	0.17	90	35	21
0.75		0.51	45	-50	-10	0.32	38	18	5	0.17	90	35	20
0.85(4)		0.52	45	-50	-10	0.315	36.5	17.5	5	0.165	80	35	18
1.00		0.53	45	-50	-10	0.31	35	17	5	0.16	70	35	16
1.30(5)		0.545	45	-50	-10	0.305	32.5	17	4.5	0.15	65	35	14.5
1.65		0.56	45	-50	-10	0.3	30	17	4	0.14	60	35	13
2.10 (6)		0.57	45	-50	-10	0.295	27.5	17	3.5	0.135	57.5	35	12
2.45		0.58	45	-50	-10	0.29	25	17	3	0.13	55	35	11
2.9(7)		0.6	45	-50	-10	0.28	22.5	16	3	0.12	50	32.5	10.5
3.50		0.62	45	-50	-10	0.27	20	15	3	0.11	45	30	10
6.00(8)		0.64	45	-50	-10	0.25	19	14	3	0.11	40	25	8

Note:

1)

Assumed model based on 0.30 Au g/t variogram model

2)

Assumed model based on 0.30 Au g/t and 0.50 Au g/t variogram models

3)

Assumed model based on 0.50 Au g/t and 0.75 Au g/t variogram models

4)

Assumed model based on 0.75 Au g/t and 1.00 Au g/t variogram model

5)

Assumed model based on 1.00 Au g/t and 1.65 Au g/t variogram model

6)

Assumed model based on 1.65 Au g/t and 2.45 Au g/t variogram models

7)

Assumed model based on 2.45 Au g/t and 3.50 Au g/t variogram models

8)

Assumed model based on 3.50 Au g/t variogram model

17.5.3

Zone 200

Grade variography shows good structure and displays moderate anisotropy between the major and semi-major axes. Two spherical models have been fitted to the experimental correlogram, with the correlogram exhibiting a moderate relative nugget effect of 38%. The spherical model has been fitted with minor ranges of 40 m, 20 m and 8 m. A second spherical model has been fitted with overall ranges of 130 m, 90 m and 28 m for the major, semi-major and minor axis respectively. The short range structure accounts for approximately two thirds of the non nugget variance.

The interpreted major direction of continuity dips at 51º towards 261º12’. Table 17.7 presents the fitted grade variogram and indicator variogram models for Zone 200 while the grade variogram plot is provided in Figure 17.8.

Table 17.7              Esaase Deposit Zone 200 Correlogram Models

Grade Variable or Indicator Threshold		Nugget (C0)	Rotation (Isatis)			Structure 1				Structure 2
						Sill 1 (C1)	Range (m)			Sill 2 (C2)	Range (m)
			Z	Y	X		Major	Semi Major	Minor		Major	Semi Major	Minor
Grade Variography
Gold (Au g/t)		0.38	225	65	60	0.4	40	20	8	0.22	130	90	28
	Indicator Variography
0.10(1)		0.38	225	65	60	0.4	40	20	8	0.22	130	90	28
0.20(1)		0.38	225	65	60	0.4	40	20	8	0.22	130	90	28
0.30		0.38	225	65	60	0.4	40	20	8	0.22	130	90	28
0.40(2)		0.395	225	65	60	0.39	37.5	20	7.5	0.215	125	90	25.5
0.50		0.41	225	65	60	0.38	35	20	7	0.21	120	90	23
0.60(3)		0.43	225	65	60	0.375	32.5	19	6.5	0.195	107.5	85	21.5
0.75		0.45	225	65	60	0.37	30	18	6	0.18	95	80	20
0.85(4)		0.465	225	65	60	0.37	29	16.5	5.5	0.165	92.5	72.5	20
1.00		0.48	225	65	60	0.37	28	15	5	0.15	90	65	20
1.30(5)		0.495	225	65	60	0.37	26.5	14	5	0.135	85	57.5	20
1.65		0.51	225	65	60	0.37	25	13	5	0.12	80	50	20
2.10 (6)		0.525	225	65	60	0.36	22.5	12	5	0.115	77.5	42.5	19.5
2.45		0.54	225	65	60	0.35	20	11	5	0.11	75	35	19
2.9(7)		0.56	225	65	60	0.335	20	9.5	4.5	0.105	75	30	16
3.50		0.58	225	65	60	0.32	20	8	4	0.1	75	25	13
6.00(8)		0.6	225	65	60	0.3	18	7	3	0.1	70	23	12

Note:

1)

Assumed model based on 0.30 Au g/t variogram model

2)

Assumed model based on 0.30 Au g/t and 0.50 Au g/t variogram models

3)

Assumed model based on 0.50 Au g/t and 0.75 Au g/t variogram models

4)

Assumed model based on 0.75 Au g/t and 1.00 Au g/t variogram model

5)

Assumed model based on 1.00 Au g/t and 1.65 Au g/t variogram model

6)

Assumed model based on 1.65 Au g/t and 2.45 Au g/t variogram models

7)

Assumed model based on 2.45 Au g/t and 3.50 Au g/t variogram models

8)

Assumed model based on 3.50 Au g/t variogram model

Figure 17.8        Zone 200 Grade Variogram (Correlogram)

17.6

Block Modelling

17.6.1

Introduction

A three-dimensional block model was constructed for the Esaase deposit, covering all the interpreted mineralisation zones and including suitable additional waste material to allow later pit optimisation studies.

17.6.2

Block Construction Parameters

A sub-block model was used to construct the Esaase mineralisation and background models (Table 17.8). Block coding was completed on the basis of the block centroid, wherein a centroid falling within any wireframe was coded with the wireframe solid attribute. The block model was rotated to 045º to adequately represent the overall strike direction of mineralisation.

Table 17.8    Esaase Gold Deposit Block Model Construction Parameters

	Origin (m)	Extent (m)	Parent/Sub Block Size
Easting	619,517.145	1,100	10/2.5
Northing	723,617.143	3,400	40/2.5
RL	-250	750	5/1

The parent block size was selected on the basis of the average drill spacing (40 m section spacing) and the variogram models, which indicate estimation of blocks smaller than the data spacing is not practical. A parent block size of 10mE x 40mN x 5mRL was selected as appropriate. Sub-blocking to a 2.5mE x 2.5mN x 1 mRL size was completed to ensure adequate volume representation.

The attributes coded into the block models included the weathering and mineralisation models. A visual review of the wireframe solids and the block model indicates robust flagging of the block model.

Bulk density has been coded to the block model based on the weathering profile. The average bulk density for each subdivision, as presented in Table 17.9, was coded via a block model script. A description of the density measurement methodology can be found in Section 14.4.

Table 17.9    Esaase Gold Deposit Dry Bulk Density

Oxidation State	DBD t/m3
Strongly Oxidised	2.31
Weakly Oxidised	2.42
Transition	2.51
Fresh	2.75

17.7

Grade Estimation

17.7.1

Introduction

Resource estimation for the Esaase mineralisation was completed using MIK within all Domains. Ordinary Kriging, Inverse Distance Squared and Nearest Neighbour estimates were also completed within these domains to allow comparison with the post processed Etype mean.

Grade estimation was carried out using the Vulcan implementation of the GSLIB kriging algorithms. Calculation of selective mining unit estimates was undertaken using the Coffey Mining developed scripts. A description of the MIK estimation methodology is provided in Section 17.8.2.

17.7.2

The Multiple Indicator Kriging Method

The MIK technique is implemented by completing a series of Ordinary Kriging (“OK”) estimates of binary transformed data. A composite sample, which is equal to or above a nominated cutoff or threshold, is assigned a value of 1, with those below the nominated indicator threshold being assigned a value of 0. The indicator estimates, with a range between 0 and 1, represent the probability the point will exceed the indicator cutoff grade. The probability of the points exceeding a cutoff can also be considered broadly equivalent to the proportion of a nominated block that will exceed the nominated cutoff grade.

The estimation of a complete series of indicator cut-offs allows the reconstitution of the local histogram or conditional cumulative distribution function (ccdf) for the estimated point. Based on the ccdf, local or block properties, such as the block mean and proportion (tonnes) above or below a nominated cutoff grade can be investigated.

Post MIK Processing - E-Type Estimates

The E-type estimate provides an estimate for the grade of the total block or bulk-mining scenario. This is achieved by discretising the calculated ccdf for each block into a nominated number of intervals and interpolating between the given points with a selected function (e.g. the linear, power or hyperbolic model) or by applying intra-class mean grades. The sum of all these weighted interpolated points or mean grades enables an average whole block grade to be determined.

The following example shows the determination of an Etype estimate for a block containing three indicator cutoffs.

The indicator cutoffs and associated probabilities calculated are:

Indicator	Cutoff Grade Au g/t	Indicator Probability (cumulative)
minimum grade *	0	0.00 **
indicator 1	1	0.40
indicator 2	2	0.65
indicator 3	3	0.85
maximum grade *	4	1.00 **

Note : * Cutoff grades determined by the user.

** Indicator probability is assumed at the minimum and maximum cutoff.

The whole block grade can now be determined in this block with the following parameters used for the purposes of the interpolation:

•

Number of discretisation intervals:

            4.

•

Linear extrapolation between all points (median grade between nominated cutoffs).

The worked example is then calculated with the following steps:

•

Interval 1 (0 - 1 g/t Au)

median grade x probability / proportion attributed to the interval (0.5 g/t Au x 0.40 = 0.200).

•

Interval 2 (1 - 2 g/t Au)

median grade x proportion (1.5 g/t Au x 0.25 = 0.375).

•

Interval 3 (2 - 3 g/t Au)

median grade x proportion (2.5 g/t Au x 0.20 = 0.500).

•

Interval 4 (3 - 4 g/t Au)

median grade x proportion (3.5 g/t Au x 0.15 = 0.525).

•

Calculate total grade

average all calculated intervals ((0.2+0.375+0.500+0.525)/1) = 1.60 g/t Au.

It is also possible from this example to calculate the proportion and grade above a nominated cutoff (e.g. 2g/t - at sample support or complete selectivity). The following steps would be undertaken to calculate the tonnes and grade at sample selectivity using a 2 g/t cutoff:

•

Interval 3 (2 - 3 g/t Au)  

median grade x proportion (2.5 g/t Au x 0.20 = 0.500).

•

Interval 4 (3 - 4 g/t Au)  

median grade x proportion (3.5 g/t Au x 0.15 = 0.525).

•

Calculate total grade  

average

all calculated intervals ((0.500+0.525)/0.35) = 2.93 g/t Au with 0.35% of the block above the cutoff.

The effect of using a non-linear model to interpolate between cutoffs is to shift the grade weighting associated with that cutoff away from the median. For Esaase, the intra-class means based on the cut composite data have been used to reconstitute the ccdf and produce block statistics.

It is noted, however, that the calculation of the E-type estimate and complete selectivity often does not allow mine planning to the level of selectivity which is proposed for production. To achieve an estimate which reflects the levels of mining selectivity envisaged, a selective mining unit (“SMU”) correction is often applied to the calculated ccdf.

Support Correction (Selective Mining Unit Estimation)

A range of techniques are known to produce a support correction and therefore allow for selective mining unit emulation. The common features of the support correction are:

·

Maintenance of the mean grade of the histogram (Etype mean).

·

Adjustment of the histogram variance by a variance adjustment factor (f).

The variance adjustment factor, used to reduce the histogram or ccdf variance, can be calculated using the variogram model. The variance adjustment factor is often modified to account for the likely grade control approach or information effect’.

In simplest terms, the variance adjustment factor takes into account the known relationship derived from the dispersion variance.

Total variance = variance of samples within blocks + variance between blocks.

The variance adjustment factor is calculated as the ratio of the variance between the blocks and the variance of the samples within the blocks, with a small ratio (e.g. 0.10) indicating a large adjustment of the ccdf variance and large ratio (e.g. 0.80) representing a small shift in the ccdf.

Two simple support corrections that are available include the Affine and Indirect Lognormal correction, which are both based on the permanence of distribution. The discrete Gaussian model is often applied to global change of support studies and has been generated on the composite data set as a comparison. The indirect lognormal correction was applied to the Esaase MIK grade estimates.

Indirect Lognormal Correction

The indirect lognormal correction can be implemented by adjusting the quantiles (indicator cutoffs) of the ccdf with the variance adjustment factor so that the adjusted ccdf represents the statistical characteristics of the block volume of interest.

This is implemented with the following formula:

At the completion of the quantile adjustments, grades and tonnages (probabilities are then considered a pseudo tonnage proportion of the blocks) at a nominated cutoff grade can be calculated using the methodology described above (Etype). The indirect lognormal correction, as applied to the Esaase deposit, is the best suited of the common adjustments applied to MIK to produce selective mining estimates for positively skewed distributions.

17.8

Multiple Indicator Kriging Parameters

MIK estimates were completed for relevant domains using the indicator correlogram models (Section 17.6), and a set of ancillary parameters controlling the source and selection of composite data. The sample search parameters were defined based on the variography and the data spacing, and a series of sample search tests performed in Isatis geostatistical software. A total of 16 indicator thresholds were estimated for all Domains (see Tables 17.5, 17.6 and 17.7).

The sample search parameters are provided in Table 17.10. Soft boundaries were used in the estimation pass 1 for Zones 100 and 150 which allows samples lying within either domain to be used for the estimation of the other. For successive estimation passes, hard boundaries were used which does not allow samples lying within Zone 100 to be used for the estimation of Zone 150. This strategy allows adequate estimation in areas where the two estimation domains are adjacent to each other. Hard domain boundaries were used for the estimation of Zone 200 throughout. A three-pass estimation strategy was applied to each domain, applying progressively expanded and less restrictive sample searches to successive estimation passes, and only considering blocks not previously assigned an estimate. In addition, Zone 150 was divided into north and south portions to allow for a change in dip in the domain.

Table 17.10   Esaase Gold Deposit Multiple Indicator Kriging Sample Search Parameters

All relevant statistical information was recorded to enable validation and review of the MIK estimates. The recorded information included:

•

Number of samples used per block estimate.

•

Average distance to samples per block estimate.

•

Estimation flag to determine in which estimation pass a block was estimated.

•

Number of drillholes from which composite data were used to complete the block estimate.

The MIK estimates were reviewed visually and statistically prior to being accepted. The review included the following activities:

•

Comparison of the Etype estimate versus the mean of the composite dataset, including weighting where appropriate to account for data clustering.

•

Visual checks of cross sections, long sections, and plans.

Alternative estimates were also completed to test the sensitivity of the reported model to the selected MIK interpolation parameters. An insignificant amount of variation in overall grade was noted in the alternate estimations.

Applying the modelled variography, variance adjustment factors were calculated to emulate a 8mE x 10mN x 2.5mRL selective mining unit (“SMU”) via the indirect lognormal change of support. The intra-class composite mean grades (Table 17.4.4) were used in calculating the whole block and SMU grades. The change of support study also included the calculation of the theoretical global change of support via the discrete Gaussian change of support model.

An ‘information effect’ factor is commonly applied to the originally derived panel-to-block variance ratios to determine the final variance adjustment ratio. The goal of incorporating information effect is to calculate results taking into account that mining takes place based on grade control information. There will still be a quantifiable error associated with this data and it is this error we want to incorporate. This is achieved in practice by running a test kriging estimation of an SMU using grade control data (the results required to incorporate this option in the change of support do not depend on the assay data so the grade control data can be hypothetical). The incorporation of the information effect is commonly found to be negligible, however can have a significant effect in some cases. In this case, the information effect factor was found to have a minor effect and has been incorporated in the calculation.

The variance adjustment ratios are provided in Table 17.11.

Table 17.11 Esaase Gold Deposit Variance Adjustment Ratios  (8mE x 10mN x 2.5mRL SMU)

Zone	100	150	200
Variance adjustment factor (f)	0.28	0.13	0.33

17.9

Resource Classification

The grade estimates have been classified as Indicated and Inferred in accordance with N43-101 guidelines based on the confidence levels of the key criteria that were considered during the resource estimation. Key criteria are tabulated below.

Table 17.12         Esaase Deposit Confidence Levels of Key Criteria

Items	Discussion	Confidence
Drilling Techniques	RC/Diamond - Industry standard approach	High
Logging	Standard nomenclature and apparent high quality	High
Drill Sample Recovery	Drill core and RC recovery adequate	High
Sub-sampling Techniques and Sample Preparation	Industry standard for both RC and Diamond	High
Quality of Assay Data	Available data shows negative bias for SGS and positive bias for TWL. Poor reproducibility of individual assay results by umpire laboratory is unexplained.	Moderate
Verification of Sampling and Assaying	No drillhole twinning to reproduce original drill intercepts. Dedicated twin drilling is recommended.	High
Location of Sampling Points	Survey of all collars with adequate downhole survey. Investigation of available downhole survey indicates expected deviation.	High
Data Density and Distribution	Core mineralisation defined on a notional 40mE x 40mN drill spacing or better. Other areas more broadly spaced to approximately 80mN spaced lines (40mE spacing) reflecting a lower confidence.	Moderate
Audits or Reviews	Coffey Mining is unaware of external reviews	N/A
Database Integrity	Minor errors identified and rectified	High
Geological Interpretation	The broad mineralisation constraints are subject to a large amount of uncertainty concerning localised mineralisation trends as a reflection of geological complexity. Closer spaced drilling is recommended to resolve this issue.	Moderate
Rock Dry Bulk Density	DBD measurements taken from drill core, DBD applied is considered robust when compared with 3D data.	Moderate to high below top of transition, low in oxide material
Estimation  and Modeling Techniques	Multiple Indicator Kriging	High
Mining Factors or Assumptions	8mE by 10mN by 2.5mRL SMU	Moderate

17.10

Resource Reporting

A summary of the estimated resources for the Esaase deposit is provided in Table 17.13 below. It should be noted that mineral resources that are not mineral reserves do not have demonstrated economic viability.

Table 17.13      Esaase Deposit Grade Tonnage Report

(Multiple Indicator Kriging; 8mE x 10mN x 2.5mRL Selective Mining Unit)

Lower Cutoff Grade (g/t Au)	Tonnes (Mt)		Average Grade (g/t Au)	Gold Metal (Mozs)
		Indicated
0.4	57.987		1.2	2.278
0.5	49.248		1.4	2.153
0.6	41.942		1.5	2.025
0.7	35.748		1.7	1.898
0.8	30.656		1.8	1.777
0.9	26.322		2.0	1.660
1.0	22.782		2.1	1.552
		Inferred
0.4	41.664		1.2	1.653
0.5	34.054		1.4	1.546
0.6	28.573		1.6	1.451
0.7	24.430		1.7	1.365
0.8	20.649		1.9	1.275
0.9	17.914		2.1	1.201
1.0	15.852		2.2	1.139

Note: Appropriate rounding has been applied.

17.11

Reserves

No Mineral Ore Reserves have been reported to date as the preliminary economic assessment was based on Inferred resources and the level of study detail does not allow Reserves to be declared at this time.

18.0

OTHER RELEVANT DATA AND INFORMATION

Lycopodium is not aware of any other relevant data and information that is not already presented in this report.

19.0

ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES

19.1

Preamble

The sections below outline the basis of, and the key economic parameters derived from, the conceptual mine, process plant and infrastructure requirements for the Esaase Gold Project that were developed for the Preliminary Economic Assessment.

19.2

Open Pit Geotechnical Assessment

The Esaase gold deposit geometry is defined by a distinctive structural boundary that divides the more deformed, altered, and mineralized siltstone shale unit in the hanging wall from the more massively bedded siltstone in the footwall. All rocks in the siltstone / shale package are moderately to strongly folded and foliated, with the shale generally displaying better development of foliation than the siltstone.

The zones of rock types that are relevant for this study are those of the weathered material, of the hanging wall, and of the footwall.

The groundwater observations indicate that the water levels near surface are between 0.1 to 105 metres below surface with an average “below surface” value of 22.4 metres. An unconfined aquifer will be involved. Hence it is probable that high pore pressures will exist within the open pit walls.

Two sources of information formed the basis of the geotechnical data: namely information derived from resource focussed drilling (e.g. RQD and weathering) and information derived from six specific geotechnical drill holes.

The geometric slope recommendations for pit design are listed in Table 19.1.

Table 19.1      Slope Geometry Recommendations

Zone	Batter Angle [Degrees]	Batter Height [m]	Berm Width [m]	Inter Ramp Slope Angle [Degrees]
Weathered material Fresh rock – Foot wall Fresh rock – Hanging wall	56 70 75	10 20 20	3.0 8.5 8.5	46 52 55

The above slope recommendations assume drained slopes. Draining operations should start ahead of the mining operation with vertical well pumping. Such an approach must be implemented to assist the mining operations to be based on drained material as well as forming a reduction of the water table for future slopes. Once mining operations start, horizontal dewatering drilling would also be necessary to drain the slope walls to limit the slope’s pore pressures.

19.3

Proposed Mining Operations

19.3.1

Mining Method and Equipment

It was assumed that mining by conventional open pit method, drill and blast followed by load and haul, would be employed at the project.

Depending on the mine production rate, the mining fleet would most likely consist of between 120 t and 250 t sized hydraulic excavators and off highway dump trucks with a capacity of between 90 t to 150 t.

For the purpose the Study it was assumed that the main production fleet would comprise a 200 t class excavator and 90 t rated dump trucks.

19.3.2

Open Pit Design

The Whittle Four-X optimisation software was used for the pit optimisation work.

The resource model used in this document formed the basis of the pit optimisations. The pit optimisations were carried out based on the total Resource, including Inferred Resources and were carried out for three processing rates, namely 3.0 Mtpa, 5.0 Mtpa and 7.5 Mtpa.

For the purpose of the pit optimisation work five incremental metal parcels were calculated, based on five grade ranges, as set out below:

·

0.3 g/t – 0.4 g/t Au

·

0.4 g/t – 0.5 g/t Au

·

0.5 g/t – 0.6 g/t Au

·

0.6 g/t – 0.7 g/t Au

·

0.7 g/t – 0.8 g/t Au

·

0.8 g/t – 1.0 g/t Au

·

> 1.0 g/t Au.

In addition, the individual metal parcels were further categorised by weathering type as set out below:

·

Strongly weathered.

·

Transition plus Fresh.

The pit optimization financial analysis was carried out using the following base assumptions and parameters:

•

Mill feed production

3 Mtpa, 5 Mtpa or 7.5 Mtpa.

•

Mill limiting

sufficient waste is removed each period to enable the respective ore production rate to be maintained.

•

Discount rate

10% Real.

The costs associated with the mining operations were based on owner mining and were developed from Coffey Mining’s database and a number of similar sized current project estimates.

Processing costs for a carbon in leach (CIL) operation and the general and administration costs were provided by Lycopodium via Keegan.

The base case revenue parameters provided by Keegan and used for the Whittle optimisation were as follows:

•

Gold price

:US$850/oz

•

Royalty

3%

Based on the economic input parameters the cutoff is approximately 0.4 g/t Au.

At a 5.0 Mtpa processing rate the base case optimum pit shell based on the maximum undiscounted operating cashflow reached a depth of 70mRL or approximately 330 m below surface on the high wall side (south) of the pit. The pit shell contains 50.5 Mt of mill feed at 1.3 g/t for some 2.0 Moz of recovered gold. Some 181 Mt of waste are contained within the pit with a stripping ratio of 3.6:1. The undiscounted operating cashflow, exclusive of capital and start up costs, is US$538M. The Worst Case discounted cashflow is US$301 M, whilst the Best Case discounted cashflow is US$379M and the average discounted operating cashflow is US$340.

To average cash operating cost, including royalty, was estimated at $578/oz.

No detailed pit design work was undertaken for the Study. The optimum pit shell of the 5.0 Mtpa pit optimisation scenario, based on the total resource, including Inferred Resources, was used as the basis for further mine planning work.

19.3.3

Waste Storage Area

No detailed waste dump designs were undertaken. However, in order to assess the impact of waste dump capacity on the overall site layout, basic dump outlines were produced at an overall slope angle of 20 degrees.

19.3.4

Grade Control

It is envisaged that grade control will be carried out by reverse circulation (RC) drilling. RC grade control was assumed to be carried out on an 8 m x 6 m drill pattern, with samples collected every 2.5 m intervals. An allowance of 20% was built in for additional drilling and assaying to take into account the need to accurately define the edges of the ore.

RC drilling cost was estimated at $20/m, whilst assay cost was estimated at $4.35 per assay.

19.3.5

Production Schedules and Blending

The schedule has been developed on an annual basis. The schedule is based on bench by bench mining of the quantities calculated within the individual pit stages.

It is estimated that commissioning of the process plant commences after approximately three to six months of pre-strip, when a reasonable mill feed stockpile has been established.

The life of mine mining schedule associated with the adopted processing scenario is shown in Table 19.2.

Table 19.2     Summary Mining Schedule 6.5 Mtpa Oxide, 5 Mtpa Thereafter, No Stockpiling

Period	Total (Mt)	Waste (Mt)	Strip Ratio (w:o)	Mill Feed Mined		Mill Feed On Stockpile (Cum.)		Mill Feed Processed
				Tonnes (Mt)	Grade (g/t]	Tonnes (Mt)	Grade (g/t	Tonnes (Mt)	Grade Ig/t]	Ox/Sul Ratio I%]
Pre-production	3.4	2.8	5.1	0.6	1.39	0.8	1.42	0.0	0.00
Year 1	22.6	16.8	2.9	5.8	1.38	0.0	1.17	6.4	1.38	100%
Year 2	25.5	18.1	2.4	7.5	1.33	1.6	1.35	6.5	1.32	100%
Year 3	26.4	21.4	4.2	5.0	1.37	1.0	1.38	5.0	1.37	24%
Year 4	26.3	22.1	5.2	4.3	1.33	0.3	1.38	5.0	1.33	21%
Year 5	26.8	21.8	4.3	5.0	1.30	0.3	1.38	5.0	1.30	21%
Year 6	26.9	21.9	4.4	5.0	1.24	0.3	1.38	5.0	1.24	6%
Year 7	26.3	21.3	4.3	5.0	1.19	0.3	1.38	5.0	1.19	4%
Year 8	24.6	19.6	3.9	5.0	1.26	0.3	1.38	5.0	1.26	1%
Year 9	15.3	10.5	2.2	4.8	1.17	0.0	0.00	5.0	1.18	0%
Year 10	4.9	2.4	1.0	2.5	1.47	0.0	0.00	2.5	1.47	0%
Total	229.1	178.6	3.5	50.5	1.3			50.5	1.3	33%

19.4

Hydrology and Hydrogeology

Groundwater is likely to occur in three distinct (but not necessarily separate) aquifer types:

Perched water in soil intersticies within alluvium, Bonte River mining tailings, and residual soils. This aquifer will generally be shallow with limited water quantities, but is also the aquifer likely to respond most to direct rainfall infiltration. This aquifer possibly presents its greatest risk to the pit where an existing stream bed crosses the pit envelope (approximately 724000N) and old tailings / Bonte River alluvium are at their thickest.

Base of weathering rock stress release zone (groundwater associated with a matrix of interconnected stress release joints). This zone will extend across the whole of the pit envelope, but permeabilities can be expected to be comparatively low

Significant fault systems or other structural discontinuities. High instantaneous yields are possible from such systems, however stored water in such systems can be quite small – with the result that they can be rapidly dewatered.

Recharge to the aquifers will occur as direct rainfall recharge, and also from permanent water bodies providing a constant source of water – a constant head “boundary”. The crossing of the northern extension of the ore body by Bonte River is one likely constant head boundary – particularly given the likelihood that water bearing fracture systems will be exposed in the base of the River. The presence of the stream crossing the pit at about 724000N, the high groundwater level and the close proximity of Bonte River to the west of this location are all factors in a second area of potential constant head. These can be substantially addressed by engineering works as part of pit development.

Based on an interpretation of pit groundwater contours it is believed that a flow system exists with groundwater probably discharging from east to Bonte River to the north and to the west of the pit. Furthermore, a groundwater divide exists within the pit. The steepness of the groundwater contours, the small area in which recharge is likely to be able to occur, and the absence of evidence of significant spring discharge in Bonte River (the river does cease flowing in dry seasonal conditions) all suggest natural groundwater throughflow in this area is likely to be small.

19.4.1

Open Pit Dewatering

Preliminary conclusions made by independent consultant Knight Piésold Pty Ltd included the following points:

•

Hydraulic conductivities “k” for the matrix are expected to range from 0.05 m/d to 1 m/d, with k of certain faults and fractured zones ranging from 5 m/d to 15 m/d.

•

Recharge in the region is expected to be between 8% and 11% of mean annual precipitation.

•

Groundwater inflow into the pit is expected to be in the region of 250L/s to 450L/s.

Based on Coffey Mining’s experience elsewhere in Ghana (including the back analysis of groundwater flows into two existing pits in Birimian sediments during pit dewatering, as well as interpreting the results of numerous small scale pumping tests in surrounding areas to those pits), permeabilities may be at the lower end of the range (about 0.1 m/d) for the rock weathering profile, and substantially lower in fresher rock.

On the steep sloped terrain at Esaase with low permeability weathering products reducing infiltration rates, Coffey Mining would expect groundwater recharge to be substantially less than 8% to 11% (possibly an order of magnitude less). Higher recharge is possible over the high water table flood terraces associated with Bonte River.

In view of the above, Coffey Mining believes the Knight Piésold estimates of groundwater inflow are likely to be pessimistic, and therefore represent an “upper bound”.

In the absence of major hydraulic recharge boundaries, groundwater inflows for the base case pit optimisation shell may be as low as 50L/s – and much of the seepage may not be recoverable due to evaporation off the pit walls (an average daily evaporation rate of 4 mm/day if applied over the full area of a pit envelope of 1 km², would be theoretically more than 40L/s). The “lower bound” recoverable seepage rate from the pit, taking into account dust suppression requirements, could be zero. As a result Coffey Mining recommends that Keegan plan for an alternative water supply (Ofin River) for standby during dry season conditions.

Coffey Mining has noted two potential recharge zones. If the Bonte River to the north of the pit is feeding a significant fracture system that could provide a recharge mechanism to the pit, then dewatering bores would be targeted to intersect the fracture system and pump the inflowing water back into the Bonte River. This becomes more difficult the closer the pit is to the Bonte River, and problematic if the pit was to extend as far as the Bonte River. Alternatively, the creation of a lined section of the river could be considered to reduce any such infiltration should it become a significant problem.

The stream crossing the pit at 724000N, the naturally high water table and the associated saturated sediments / tailings represents the second potential recharge area. Coffey Mining anticipates a requirement for stream management including diversionary works will be essential for full pit development, and consideration may also need to be given to managing an eastwards groundwater seepage front through the surficial tailings and alluvium material from the Bonte River as it passes to the west of the pit.

There is a considerable incentive as pit design proceeds, to address both these potential hydrogeological features through engineering design.

It should be noted that the quantities of water that are able to seep into the pit from both sources, will depend on the width or thickness of water bearing systems within the pit, and the overall permeability of those systems. The currently proposed test regime is targeted at providing a greater insight to permeabilities across the pit through an assessment of bore yields, confirmation of hydraulic gradients, hydrogeochemical assessment, and also monitoring bore responses to major rainfall and river flow events.

Based on the above Coffey Mining believes that the water inflow range provided by Knight Pies°Id is the upper bound.

Based on the very limited data Coffey Mining believes that the water inflow is likely to be in the range of 100L/s+501/s and 100L/s was adopted as the basis for the study.

19.4.2

Water Supply

An adequate supply of process water is anticipated to be available.

Make-up water supply for the processing plant will be provided from one or more of three sources:

•

Pit dewatering

•

Local bores

•

The nearby Ofin River.

Water supply will be the subject of further investigation and an options study during the Pre-feasibility Study.

19.5

Tailings Storage Facility Designs and Management

A preliminary site selection and conceptual design and costing for the tailings storage facility (TSF) at Esaase was undertaken by Coffey Mining. An assessment of four preliminary sites was made.

The design of the tailings storage facility (TSF) options for the Project was aimed at:

•

Optimising tailings storage capacity and maximising tailings density;

•

and Reducing environmental and societal impact.

For the purposes of the study a site located north of the proposed mine plant at the head of the Bronte river was selected. A perimeter embankment at 372mRL provides containment of a final capacity of 59 Mt. Total catchment area is approximately 523 ha and the final tailings area is approximately 210 ha.

19.6

Infrastructure and Services

19.6.1

Introduction

The scope of infrastructure and services required for the project was developed from a "desk top" evaluation of project requirements in consultation with Keegan and other consultants, principally Coffey Mining.

With the exception of the tailings storage facility, design of buildings and infrastructure has been limited to the development of a list providing only a brief description of the scope of each building / item on the basis that the capital cost for these items will be derived by the use of factors from the Lycopodium estimating database.

19.6.2

Power Supply

A preliminary approach was made to the Volta River Authority (VRA) by Lycopodium, on behalf of Keegan, to confirm the conceptual approach to the provision of grid power by the VRA.

A total power capacity requirement of 25 MW has been used for this study. The power requirement has been estimated from a process plant power requirement based on preliminary comminution testwork data and an allowance for infrastructure.

The proposed power supply is based on a 161 kV, 100 MVA capacity power line from the existing VRA substation 40 km south east at Obuasi, and a step-down substation at Esaase. No allowance has been made for redundant systems.

19.6.3

Site Buildings and Roads

A summary list of buildings, infrastructure, roads and miscellaneous items was prepared to cover the scope of facilities required to support the operation.

Site buildings allowed in the capital estimate are as follows:

Steel Framed Buildings

·

Plant workshop / warehouse.

·

Reagent storage shed.

Mine Facilities

•

Mine change room / meal room.

•

Heavy and light vehicle workshop.

•

Truck washdown bay. Explosives magazine.

•

Fuel storage and dispensing facility.

Plant and Administration Buildings

·

Administration office incorporating mine offices and clinic.

·

Community relations building.

·

Assay laboratory.

·

Plant metallurgical laboratory.

·

Plant gatehouse / change rooms.

·

Plant office.

·

Plant meals.

·

Plant control room.

·

Switchrooms and substations.

Accommodation and Other Buildings

·

6 off senior staff 3 bedroom houses.

·

3 blocks of 8 senior single quarters with en-suite bathrooms.

·

Senior staff wet / dry mess.

·

Senior staff laundry.

·

Village administration office.

Temporary Buildings

10 dormitories for construction personnel.

·

Construction mess.

·

Construction laundry.

19.6.4 Miscellaneous Facilities

Allowance has been made in the capital estimate for:

·

Haul roads (in the mining estimate).

·

Revegetation nursery and rehabilitation area.

·

A sewerage treatment plant servicing the plant, offices and accommodation.

·

Upgrading of the site access road where necessary.

·

Roads and tracks servicing the plant and TSF.

·

Landfill sites for domestic and hazardous waste.

·

Tailings and decant return pipelines.

·

Mobile equipment.

19.7

Environmental / Permitting Process

This section provides an overview of the Ghanaian environmental legislation and guidelines applicable to the project and summarises the project permitting process.

19.7.1

Ghanaian Legislation and Guidelines Environmental and Social

Key environmental and social legislation in Ghana is the Environmental Protection Agency Act 1994 (Act 490) and the Environmental Assessment Regulations 1999.

The Environmental Protection Agency Act 1994 establishes Ghana's Environmental Protection Agency (EPA) and defines the functions of the EPA, including, but not limited to, the following:

·

Prescribing standards and guidelines relating to the pollution of air, water and land.

·

Ensuring compliance with environmental impact assessment procedures in the planning and execution of development projects.

Any undertaking that has the potential to have an adverse impact on the environment can be required by the EPA to submit an Environmental Impact Statement (EIS) under Part II of the Environmental Protection Agency Act 1994. Where this is the case, no other government department can issue any licence, permit or approval until notified by the EPA that the EIA for the undertaking has been submitted and approved. The EIS covers both the biophysical and the socio-economic aspects and impacts of the project.

If the undertaking poses a serious threat to the environment or to public health then the EPA may serve an enforcement notice, specifying the offending activity, the steps required to be taken to prevent harm to the environment or to public health, and the time in which these steps must be taken. Any person breaching an enforcement notice commits an offence and can be fined or imprisoned.

The Environmental Assessment Regulations 1999 support the Environmental Protection Agency Act 1994 and describe the process of environmental assessment in Ghana.

Submission of an EIS is mandatory for any mining project where the mining lease covers a total area in excess of 10 hectares (25 acres). The regulations outline the environmental and social aspects that must be addressed in an EIS. This includes addressing the possible direct and indirect environmental impacts of the proposed undertaking during the pre-construction, construction, operation, decommissioning (i.e. mine closure) and post-decommissioning phases.

An Environmental Scoping Document must first be prepared and approved by the EPA prior to submitting an Environmental Impact Statement. As the name suggests, the purpose of the scoping document is to determine an agreed scope of works for the EIS and must include a draft terms of reference.

The regulations also prescribe a number of activities, which must be carried out once an environmental permit is obtained. These activities include:

·

Submit and have approved an Environmental Management Plan within 18 months of commencement of operations. The Environmental Management Plan must be revised and reapproved every three years thereafter.

·

Submit an Annual Environmental Report 12 months after the commencement of operation and every 12 months thereafter.

·

Obtain an Environmental Certificate from the EPA within 24 months of commencement of operations. This certificate is issued subject to:

·

evidence of actual commencement of operations

·

acquisition of other permits and approvals, where applicable

·

evidence of compliance with mitigation commitments indicated in the Environmental Impact Statement or Preliminary Environmental Report

·

submission of the first Annual Environmental Report.

Minerals and Mining

The primary legislation governing mining in Ghana is the Minerals and Mining Law 1986 (P.N.D.C.L. 153) and the associated Minerals and Mining (Amendment) Act 1994 (Act 475). An amendment to this act is pending: the Minerals and Mining Act of 2006 (Act 703) has been released as a draft, but is not yet promulgated.

In order to obtain the right to mine, a mining lease must be granted under the Minerals and Mining Law 1986. A mining lease cannot be granted unless the following requirements are satisfied:

·

The proposed program of mining ensures the most efficient, beneficial and timely use of the mineral resources concerned.

·

The proposed mining operations take proper account of environmental and safety factors.

·

The area of land over which the mining lease is sought is not in excess of the area reasonably required to carry out the proposed program of mining.

·

The applicant is not in default.

Once granted, the holder of a mining lease must notify the Provisional National Defence Council Secretary for Lands and Natural Resources of any amendments to the program of mining operations. Unless the Secretary rejects these amendments within three months of notification, the amendments will have affect after this time. Under Section 72 of the Minerals and Mining Law 1986, the holder of a mineral right must have due regards to the effects of mineral operations on the environment and must take whatever steps necessary to prevent pollution of the environment as a result of mineral operations.

In addition, the Minerals and Mining Law 1986 requires applicants for a mining lease to provide a detailed program for the recruitment and training of Ghanaian personnel and requires mineral operations to give preference to materials and products made in Ghana and service agencies located in Ghana and owned by Ghanaians, companies or partnerships registered under the Companies Code Act 1963 (Act 179) or the Incorporated Private Partnerships Act 1962 (Act 152), or a public corporation.

The Minerals and Mining Law 1986 (Part IX) also sets out the surface rights of any lawful occupier of the land over which mineral rights are held. Mineral rights include the right to reconnoitre, prospect for minerals or to mine minerals and must be exercised in a manner that will minimise the impacts to any lawful occupier of the land. The lawful occupier of any land within the area subject to a mineral right shall retain:

·

The right to graze livestock on or cultivate the land, provided grazing or cultivation does not interfere with mineral operations (applies to both mining and exploration).

·

The owner or lawful occupier of the land within a mining area must not erect any building or structure without the consent of the holder of the mining lease. The holder of the mining lease must give consent where a request is reasonable (applies to mining operations only).

·

The right to compensation for disturbance of other rights and for damage done to the surface of the land, buildings, works or improvements, livestock, crops or trees (applies to both mining and exploration).

Compensation

Acquisition and access to land in Ghana for development activities, including mining, may be undertaken either through the states power of eminent domain or by private treaty. The taking of land in either way requires the payment of due compensation or, where people are displaced, they may be resettled depending on the circumstances of the displaced persons. The regulatory oversight of private sector land acquisition and resettlement related to mining activities and actions is governed by the Constitution and two legislative acts:

·

The 1992 Constitution of Ghana, ensures protection of private property and establishes requirements for resettlement in the event of displacement from State acquisition (Article 20 [1, 2 and 3]).

·

The State Lands Act of 1962 (Act 125) and its subsequent amendment, State Lands (Amendment) Act 2000 (Act 586), mandates compensation payment for displaced persons and sets procedures for public land acquisitions.

·

The Minerals and Mining Law 1986 (Part IX) as discussed immediately above.

·

The Minerals and Mining Act of 2006 (Act 703), currently only a draft and not yet promulgated, vests all mineral rights in land to the State and entitles land owners or occupiers to the right for compensation. In particular, the Minerals and Mining Act 2006 (Act 703), Section 74 [1], requires compensation for:

·

deprivation of the use or a particular use of the natural surface of the land or part of the land

·

loss of or damage to immovable property

·

in the case of land under cultivation, loss of earnings or sustenance suffered by the owner or lawful occupier, having due regard to the nature of their interest in the land

·

loss of expected income, depending on the nature of crops on the land and their life expectancy.

The latter three items have been the norm since the previous law was introduced in 1986, whereas an approach to the first has not been established or tested. Newmont's Akyem and Ahafo expansion projects and Adamus' N'zema Project are anticipated to be formulating an approach to this. Example compensation rates have been obtained for the Ahafo and Akyem projects. Compensation rates for other projects will be sought.

In addition, there is a requirement for the determination of the amount of compensation to be by agreement between the parties (Section 73 [3]).

Mining and Environmental Guidelines

Ghana's Mining and Environmental Guidelines were developed by the Minerals Commission and the then Environmental Protection Council (now EPA) in May 1994, in consultation with government agencies, mining companies, nongovernmental organisations (NG0s), universities, research institutions and the general public.

The Mining and Environmental Guidelines cover both exploration and mining activities as described below.

Application to Explore

The Mining and Environmental Guidelines require exploration companies to submit a set of forms, entitled Environmental Overview, to the Minerals Commission, prior to the commencement of field exploration.

The guidelines specify requirements for exploration liaison, including the requirement to meet appropriate land owners, or their representatives, at least 14 days in advance of exploration, to outline the nature and timing of the proposed exploration program and the likely environmental effects of each phase, proposed compensation arrangements, general opportunities for casual employment and the supply of goods and services.

In addition, the company must meet with the appropriate land owners at least seven days prior to the commencement of each phase of exploration work to describe and/or determine the following:

·

The impending phase of work.

·

The area of land that will be explored.

·

The expected effects on vegetation, land use, air quality, water quality and noise levels.

·

Applicable compensation rates and when compensation will be determined and paid.

·

The number of casual labourers required and the need for goods and services.

·

Location of any sacred groves, burial grounds or other fetish lands.

·

Location of any area of environmental sensitivity to local people.

·

Provision of an address and telephone number of a locally based company representative.

The guidelines also detail environmental and consultation requirements for access management, drilling and excavations and abandonment of exploration sites. Exploration site rehabilitation must be fully completed within three months of cessation of exploration.

Application to Mine

The Mining and Environmental Guidelines stipulate that pursuant to the Environmental Protection Agency Act, 1994, any company proposing to develop a mining project that covers a land surface area of greater than 10 hectares must submit an EIS to the EPA for approval. Concurrently, copies of the EIS must be submitted to the Minerals Commission and the Mines Department. The guidelines specify the aspects of the project that must be considered in an EIS and the minimum information required.

In addition, mining companies must submit an Environmental Action Plan (EAP) as part of the EIS. The EAP outlines specific management and mitigation measures that a mining company will put in place to prevent environmental harm and must be updated every two years.

Ongoing reporting requirements include submission of an Annual Environmental Report at the end of each calendar year to the EPA, the Minerals Commission and the Mines Department. Mining companies must also submit copies of all environmental audit reports to these same parties. These audit reports are kept confidential and are not in the public domain.

Other EPA Guidelines

The EPA has published three environmental guidelines applicable to mining activities, which prescribe maximum permissible levels for various residential and industrial settings:

·

Sector Specific Effluent Quality Guidelines for Discharges into Natural Water Bodies.

·

Environmental Quality Guidelines for Ambient Air.

·

Environmental Quality Guidelines for Ambient Noise.

The EPA has also published two guidelines relevant to social and environmental impact assessment:

·

Environmental Assessment in Ghana, A Guide (1996).

·

Environmental Impact Assessment Procedures (1995). 19.7.2 Permitting Process

The primary permits required for the development of the Esaase Gold Project are an Environmental Permit issued by the EPA, and an amendment to the existing mining leases issued by the Minister on the advice of the Minerals Commission. Various other permits from other government departments will also be required (e.g. water abstraction permits), but are relatively secondary in nature.

The processes outlined below may be refined pending further discussions with the EPA, Minerals Commission and other regulatory stakeholders. The processes may also be subject to change and/or require reiterative steps, pending the findings of the impact assessment process and stakeholder consultation and negotiation.

Environmental Permit

As per the discussion of the Environmental Assessment Regulations 1999, the Esaase Gold Project will require an EIS.

Based on consultation to date with the EPA, the following permitting process has been defined:

·

Keegan submits an Environmental Assessment Registration Form (EARF) to the EPA, which contains a short description of the project and proponent.

·

The EPA reviews the EARF and issues a Screening Report within 25 days stating that the project requires an EIS.

·

Keegan submits an Environmental Scoping Document (ESD) for the project, the purpose of which is to determine an agreed scope of works for the subsequent EIS.

·

The EPA reviews the ESD and, within 25 days, advises whether it is acceptable.

·

Once the ESD is accepted, Keegan notifies the relevant Ministries, publicly advertises the proposed undertaking, and makes copies of the ESD available for inspection by the general public in the project locality.

·

Dependent on the significance of any potential social or environmental issues identified in the ESD, the EPA may choose to hold a Public Hearing at this point or not.

·

Keegan submits the Draft EIS, which includes a Conceptual Mine Closure Plan and, if necessary, a Resettlement Action Plan (RAP).

·

The EPA publishes a notice for 21 days and makes copies of the EIS publicly available. Public comments are received.

·

The EPA holds a public hearing to solicit feedback and concerns, which is conducted by an appointed panel of stakeholders. The panel provides written comments and recommendations within 15 days.

·

The EPA reviews the draft EIS after receipt of the public hearing recommendations and advises whether it is acceptable.

·

Additional studies or consultations with stakeholders may be necessary before the EPA is satisfied to approve the EIS and issue an Environmental Permit. Similarly, if material changes are made to the project design during the Definitive Feasibility Study, additional studies or consultations may need to be undertaken and the EIS revised.

·

Following approval of the EIS, a Security Agreement may be negotiated to set out the level of project royalties, taxes and other fiscal conditions for a period of 15 years. The Security Agreement would also include agreed / legal reclamation criteria, and set out the reclamation liability and reclamation bond conditions.

·

Once the EPA has approved the EIS, any other pending approvals by other government departments can be issued. Examples of this may include granting of amendments to the existing mining leases by the Minerals Commission and granting of water abstraction permits by the Water Resources Commission.

The project permitting schedule is shown in the context of the overall project development schedule in Figure 19.1 below. It is anticipated that it will take until 1 Q12 to receive EPA approval, which will be necessary prior to starting site works. It is important to note that the permitting schedule as shown is dependent on timely availability of the various engineering studies that will form the “project description”, upon which the impact assessment will be undertaken and mitigation and management measures developed. Although the baseline surveys and stakeholder consultation are proceeding in parallel with current engineering work, the EIS document can not be finalised (or submitted) until the project description is fully defined by the future Pre-Feasibility (PFS) and Definitive Feasibility (DFS) studies. Therefore, preparation of the EIS and subsequent EPA approval are anticipated to be on the critical path for project development.

19.8

Project Development Schedule

While the Preliminary Economic Assessment provides a preliminary assessment of the likely economics of the project, prior to committing to detail design, procurement and construction, the concepts outlined will need to be further developed and backed by auditable design data.

A more comprehensive program of metallurgical testwork is required to provide a firm basis of design for the process plant. This is planned for 2010.

Site visits and a more detailed analysis of needs are required to scope the infrastructure and services requirements of the project.

Baseline social and environmental surveys must be completed, environmental impact assessments submitted and approved and a range of project permits obtained.

Additional and ongoing work is requirements to progress the Coffey Mining scope through pre and definitive feasibility studies.

A high level preliminary schedule has been developed (refer Figure 19.1) providing an indication of the likely time scale required for project development. The schedule assumes that development will be based on the current mineral resources with no delays for additional exploration or definition drilling although it does not preclude these happening in parallel with the scheduled activities.

The schedule predicts first gold production at the beginning of 2Q14.

Project development is broken down into the following combination of activities and phases:

·

Sample selection and pre-feasibility study level testwork program.

·

Pre-feasibility study to identify the optimum project scope and scale.

·

Definitive feasibility study to optimise the scope and prepare ±15% capital and operating cost estimates.

·

Baseline studies and permitting.

·

Finalisation of project financing.

·

Engineering, procurement and construction.

·

Commissioning and ramp up.

Figure 19.1                  Preliminary Project Development Schedule

\

19.9

Project Economics

19.9.1

Capital Cost Estimate

The capital cost estimate for the project is shown in Table 19.3.

Table 19.3           Summary Capital Estimate (US$, 4Q09, +30% / -15%)

Mining · Equipment · Pre-production · Mining Infrastructure	$41.7 M $7.5 M $6.5 M	$ 55.7 M
Mining
· Equipment	$41.7 M
· Pre-production	$7.5 M
· Mining Infrastructure	$6.5 M	$ 55.7M
Infrastructure (including roads, buildings, communications)		$ 14.6 M
Mineral Processing Plant		$ 91.8 M
Tailings		$ 9.4 M
Indirects		$ 46.0 M
Owners costs		$ 42.7 M
EPCM		$ 23.9 M
Subtotal		$ 284.2M
Contingency		$ 35.4 M
Subtotal (millions)		$319.5 M
Sustaining and Deferred capital		$24M
Total (Millions)		$343M

The estimate is in United States dollars and based on pricing in the fourth quarter of 2009. No allowance has been made for escalation.

The Esaase capital cost by area, excluding sustaining and deferred capital and various owners costs, is shown in Table 19.4 below.

Table 19.4     Capital Estimate by Area (US$, 4Q09, +30% / -15%)

Mining

The capital cost estimate for owner mining was based on estimates developed as of the 4th quarter of 2009.

The capital cost attributable to mining will primarily comprise the following items:

·

Mobile mine equipment.

·

Haul road construction.

·

Mine infrastructure:

·

workshop and warehouse

·

mobile equipment wash pad

·

tyre bay

·

fuel storage offices

·

accommodation.

·

Other:

·

Light vehicles Survey equipment

·

Computer hardware and software.

It is estimated that over the life of mine approximately $7.5 million of capital expenditure will be required to replace mine equipment that have reached their rated equipment life.

Process Plant and Infrastructure

The direct capital cost for the process plant and infrastructure was estimated by Lycopodium based on the developed scope of facilities.

The cost of the initial starter dam for the Tailings Storage Facility (TSF) was estimated at $6.7 million (pre-contingency). This amount excludes the waste haulage component as this is included in the mining costs.

Construction lndirects

Construction indirects were estimated based on recent similar projects in similar locations.

EPCM

A flat percentage rate of 15% for EPCM was applied against all direct cost items.

Contingency

A flat percentage allowance of 15% was applied to all items with the exception of owners costs, pre-production costs and spare parts allowance.

Escalation

No escalation has been allowed in the estimate.

Owners Costs

An allowance of $6 million was advised by Keegan to salary and other expenses related to maintaining an owners project management team. In addition, various pre-production expenses, an allowance for spare parts and mobile equipment outside the mine fleet were estimated.

Pre-production Costs

Pre-production costs were estimated for mining, plant operations and general administration and include:

·

Mine pre-stripping.

·

First fill and opening stocks of reagents and consumables.

·

Pre-production labour costs and recruitment expenses

·

Other non-construction costs incurred prior to commencement of operations.

Spare Parts

An allowance of $3 million has been made for initial operating and capital spare parts.

Plant Mobile Equipment

An allowance for plant mobile equipment was made based on a preliminary assessment of requirements.

Exclusions

The capital cost quoted excludes:

·

Sunk Costs

·

Escalation.

·

Future exploration / drilling costs.

·

Costs for testwork and engineering associated with the future PFS and DFS.

·

Costs for baseline studies, permitting and community programs prior to approval for construction.

·

Any head office or site costs incurred in Ghana prior to approval for construction.

19.9.2

Operating Cost Estimate

Operating costs have been determined based on the design treatment rate of 5.0 Mtpa of fresh and oxide material. Additionally, operating costs have been estimated for the treatment of 6.5 Mtpa of oxide material during the first two years of operation. The estimates have been based on a P80 grind size of 150 µm, operating 24 hours per day and 365 days per year with a milling circuit availability of 91.3%.

The mining-related operating cost estimates have been developed by Coffey Mining while the operating cost estimates for the process plant and administration have been developed by Lycopodium. The administration operating costs compiled by Lycopodium incorporate information from Coffey Mining and Keegan.

The operating costs have been compiled from a variety of sources including:

·

Mining and TSF operating costs by Coffey Mining using their in-house database.

·

Cyanide and lime consumption based on preliminary testwork.

·

Assumed reagent usage regimes for cyanide detoxification based on Lycopodium experience.

·

Budget quotations recently obtained from suppliers or Lycopodium’s database for prices of process consumables.

·

Grinding energy estimated from preliminary comminution testwork, with media consumption rates based on limited abrasion testwork data.

·

Calculations and estimates based on typical operating data for similar operations.

·

Lycopodium database of process and general administration costs for similar sized operations.

The following exchange rates have been used for the preparation of the operating cost estimate:

·

US$1.00 = A$1.20 (Australian Dollar).

·

1 Ghanaian Cedi = US$0.70.

With the proviso that the anticipated power tariff used for the estimate has been advised by Keegan to be US$0.08/kWh and reagent consumption rates have been based on limited testwork data, the process operating and administration cost estimates are considered to have an accuracy of –15% +20%. Coffey Mining quote an accuracy of ±40% for the mining operating costs. Costs are presented in United States dollars (US$) and are based on prices for the fourth quarter of 2009 (4Q09).

Operating Cost Summary

Mining operating costs are based on the life of mine average costs.

Plant operating costs have been developed from the conceptual plant design data in the process design criteria. Reagent and consumables costs are based on the limited testwork undertaken to date.

General and administration costs were developed from assumptions regarding the scope and nature of the proposed operation and incorporate a number of notional allowances where quotations or data was not available. No allowance has been made for off-site costs including head office costs and refining costs, other than labour costs for an Accra office.

A nominal allowance of US$2,000,000 per year has been included in the operating cost estimate to allow for site rehabilitation, closure and reclamation costs.

Table 19.5             Summary Operating Cost Estimate (US$, 4Q09)

Cost Centre	Design Basis 5.0 Mtpa		LOM Averaged
	Operating Cost (US$/t milled)	Operating Cost (US$/oz Au recovered)	Operating Cost (US$/t milled)	Operating Cost (US$/oz Au recovered)
Mine Operating Cost	8.91	224.82	8.83	224.80
Processing Cost	7.45	188.07	7.39	188.29
General and Administration Cost	1.17	29.62	1.15	29.25
Rehabilitation / Closure / Reclamation	0.40	10.09	0.40	10.09
Refining	0.16	4.00	0.16	4.00
Royalties	1.18	29.75	1.17	29.75
Total Operating Cost	19.28	486	19.09	486

Note:

1.

Process and administration costs developed by Lycopodium Minerals

2.

Mining and mining administration costs developed by Coffey Mining

3.

The LOM Averaged Operating Cost estimate have been based on a nominal throughput of 6.5 Mtpa during the first two years of operation, followed by a nominal throughput of 5.0 Mtpa, as used for the financial analysis

Anticipated site costs by expense type are shown in Table 19.6 below.

Table 19.6             Operating Cost Summary by Expense Type (5 Mtpa Basis) (US$, 4Q09)

Cost Centre	Averaged Annual Cost (US$/yr)	Operating Cost (US$/t milled)
Operating Consumables	39,843,519	7.97
Maintenance Materials and Contracts	23,485,990	4.70
Labour	7,737,507	1.55
Power	11,952,303	2.39
Other General and Administration Costs	4,672,703	0.93
Rehabilitation and Closure Cost	2,000,000	0.40
Total Operating Cost	89,692,022	17.94

Note:

1.

Excludes refining charges and royalty.

Anticipated manpower requirements are shown in Table 19.7 below.

Table 19.7         Summary Manning Estimate

Item	National Staff	Expatriates	Total
Mining Supervision	40	6	46
Mining Labour	104	–	104
Operations and Metallurgy	46	3	49
Process Plant Maintenance	27	3	30
Administration	120	2	122
Manning – Mine and Process	337	14	351
Mine fleet contract maintenance	70	8	78
Explosives supplier	5	–	5
Total Complement	412	22	434

19.9.3

Economic Analysis

A preliminary economic assessment of the Esaase Gold Project has been conducted using a simple cash flow model.

The economic evaluation of the Esaase Gold Project on a project basis has been based upon:

·

Capital cost estimates prepared by Lycopodium and Coffey Mining.

·

Mine schedule and mining operating cost estimates prepared by Coffey Mining. Process operating and administration cost estimates prepared by Lycopodium.

·

Sustaining capital costs for mining operation provided by Coffey Mining.

·

Owners capital costs based on estimates prepared by Lycopodium or provided by Keegan Resources.

·

Gold price provided by Keegan Resources.

·

Ghanaian corporate tax rate of 25%.

·

Total Royalties payable of 3.5% per mining lease. The cash flow analysis is based on full equity funding.

Table 19.6 provides a summary of the production information on which the cash flow model summarised in Table 19.7 has been based.

The mine life capital cost for the project at a nominal design throughput rate of 5.0 million tonnes per annum, is estimated at US$350.1 M.

At a gold price of US$850 per ounce, the project is estimated to have an after-tax IRR of 17.2% and a pay-back period of 3.34 years. At a discount rate of 5.0%, the after-tax NPV is estimated at US$168.1 million.

The project economics are summarised in Table 19.8.

Table 19.8            Project Production Summary

Mining Schedule Primary material mined Oxide material mined Waste mined	33.7 16.8 178.6	Mt Mt Mt
Total Material Mined	229.1	Mt
Total Mill Feed Processed	50.5	Mt
Mine Life	10	years
Contained Gold	2,105	koz Au
Recovered Gold	1,982	koz Au
Average Strip Ratio	3.5	(w : o)
Average Grade	1.30	g Au / t
Average Gold Recovery	94.0	%
Average Process Plant Throughput	5.04	Mtpa
Average Annual Gold Production	198	koz Au / y

Table 19.9                 Project Cash Flow Summary

Year 1 - 3 US$ Million		Project US$ Million	US$/oz Au Recovered	US$/t Milled	US$/t Mined
Mining Cost	149.7	474.2	239.29	9.40	2.34
Processing Cost	119.5	373.1	188.29	7.39	1.66
General and Administration Cost	8.8	29.2	14.75	0.58	0.13
Rehabilitation / Closure / Reclamation	6.0	20.0	10.09	0.40	0.09
Total Operating Cost	284.0	896.6	452	17.77	4.22
Smelting and Refining Cost	2.9	7.9	4.00	0.16	0.04
Royalties	21.7	59.0	29.75	1.17	0.26
Total Cash Cost	308.6	963.5	486	19.09	4.27
Revenue	620.3	1,704.3	850	33.38	7.46
Total Cash Cost	308.6	963.5	486	19.09	4.27
Operating Cash Flow (EBITDA)	311.7	740.8	364	14.28	3.19
Depreciation and Amortisation	322.1	359.9	182	7.13	1.59
Earnings Before Interest & Taxes (EBIT)	(10.4)	381.0	182	7.15	1.60
Interest	–	–	–	–	–
Gross Profit before Tax	(10.4)	381.0	182	7.15	1.60
Tax	–	85.2	14.75	0.58	0.13
Net Profit After Tax	(10.4)	295.7	167	6.57	1.47

Table 19.10             Project Financial Measures Summary

Mine life	10	Years
Revenue from Gold	1,684.3	US$ M
Direct cash cost (operating cost only)	452	US$ / oz Au
Total cash cost (including royalties, refining)	486	US$ / oz Au
Capital expenditure (excl working capital)	343.5	US$ M
Initial capital investment	325.5	US$ M
Plant and equipment salvage	20.0	US$ M
Pre-Tax Economics
Free cash flow after cost allocation (undiscounted)	397.3	US$ M
Internal rate of return (IRR)	19.5	%
Project NPV (discounted at 5.0%)	221.4	US$ M
Payback period	3.335	Years
After-Tax Economics
Free cash flow after cost allocation (undiscounted)	312.1	US$ M
Internal rate of return (IRR)	17.2	%
Project NPV (discounted at 5.0%)	168.1	US$ M
Payback period	3.34	years

The cash flow analysis is based on the following:

·

Annual tonnage, strip ratio and head grade have been based on the preliminary mining schedule.

·

The mining, processing and administration costs estimated by Coffey Mining, Lycopodium and Keegan.

·

The overall recovery figures of 93.3% for oxide material as 94.4% for primary material have been based on preliminary testwork.

·

Capital costs estimated by Coffey Mining, Lycopodium and Keegan.

Depreciation

•

Provision has been made for depreciation at 80% write-off in the year of acquisition followed by 50% of the residual value in each subsequent year on all capital cost items, excluding working capital.

•

The calculation of depreciation in Ghana allows for an uplift of 5% in the residual capital value of the capital allowance for the asset in the second year. This means that the capital allowance is depreciated at 105% of the original value of the asset.

Corporate Tax

·

Provision has been made for company tax at 25% of gross profit.

·

Tax losses as carried forward over a period of up to 5 years from the year in which the loss was incurred.

·

It has been assumed that a tax loss of US$40 million from the sunk costs is available to be carried forward in Year -2, as advised by Keegan.

Gold Price

•

A gold price of US$850 per ounce has been assumed (as advised by Keegan).

•

A refinery gold payable rate of 99.99% has been assumed.

•

The refining charge of US$4.00 per ounce Au (allowance) is assumed to include the cost of transport of the gold from site to the refinery.

Royalties

·

Royalties have been allowed for at 3.50% of revenue.

General

•

The cash flow model is based on a fully equity funded project.

•

No provision has been made for interest.

•

No provision has been made for head office costs.

•

It has been assumed that the plant will have a salvage value of US$20 million, which will be recouped in the final year of production.

19.9.4

Sensitivity Analysis

Figure 19.2 shows the sensitivity response of the calculated after-tax IRR to variations in gold prices, project capital cost, processing cost, administration cost, mining cost and royalties.

Figure 19.3 shows the corresponding sensitivity of the project NPV discounted at 5%.

Figure 19.2       Sensitivity Response of After-Tax IRR to Variation in Gold Price, Capital Cost, Processing Cost, Administration Cost, Mining Cost and Royalties

Figure 19.3     Sensitivity Response of After-Tax NPV Discounted at 5% to Variation in Gold Price, Capital Cost, Processing Cost, Administration Cost, Mining Cost and Royalties

19.9.1

Sensitivity Response and Range of Variation

The basis of estimate used for the tornado sensitivity analysis is presented in Table 19.9. Figure 19.4 shows the change in project NPV for variation of each parameter within its expected analysis range.

Table 19.11        Sensitivity Analysis Inputs

	Analysis Range		Estimate Basis (Averaged for LOM)
	Negative Variation	Positive Variation
Project capital cost	-20 %	+40 %	US$ M	343.5
Strip ratio	-10 %	+20 %	—	3.54
Geological grade	-20 %	10 %	g Au/t ore	1.30
Recovery	_3 %	+2 %	ok	94.0
Gold price	-15 %	+40 %	US$/oz	850
Royalties	-10 %	+50 %	%	3.5
Oxide material mining cost (waste and mill feed)	-10 %	+40 %	US$/t ore	7.92
Primary material mining cost (waste and mill feed)	-10 %	+40 %	US$/t ore	8.48
Processing cost	-10 %	+40 %	US$/t ore	7.39
Administration cost	-10 %	+40 %	US$/t ore	0.87

The sensitivity diagram in Figures 19.2 and 19.3 take into account both the expected range of variation in each of the parameters and the sensitivity to that parameter. The project presents as being most sensitive to gold price and geological grade, with similar sensitivities towards project capital cost and processing cost. The project seems to be relatively insensitive towards the expected variations in recovery, royalties and administration cost.

Figure 19.4     Tornado-type Sensitivity Diagram for Variable Ranges of Sensitivity

20.0

INTERPRETATIONS AND CONCLUSIONS

Drilling and studies completed to date have defined an Inferred and Indicated Mineral Resource at Esaase.

Preliminary testwork indicates that the mineralisation within the notional pit shell at Esaase is amenable to treatment by conventional grinding, gravity concentration and carbon in leach treatment to recover the contained gold.

A study of possible development scenarios has been undertaken culminating in the development of order of magnitude project capital and operating costs suitable for a preliminary economic assessment to be made.

The preliminary economic assessment indicates that, at the current level of project definition, project economics are sufficiently attractive to warrant further work to better define the scope, scale and economics of a possible future operation.

21.0

RECOMMENDATIONS

It is recommended that Keegan continue activities targeted at completing a pre-feasibility study by early 2011 as well as continuing environmental baseline monitoring and associated activities related to statutory permitting and approvals

21.1

Proposed Work Program to Advance Project to Pre-feasibility Level

A broad timeline for the activities required to advance the project through future studies, engineering, permitting, construction and commissioning is provided in section 19.8 of this document. The schedule anticipates completion of a pre-feasibility study by the end of 2010.

Preliminary budget expenditures to advance to the completion of a pre-feasibility study by the end of 2010 are as follows:

Exploration: US$13,400,000

Pre-feasibility mining study: US$250,000

Associated geotechnical and other investigations: US$150,000

Metallurgical testwork: US$350,000

Project definition and pre-feasibility engineering, preparation of study report: US$500,000

Owners costs associated with progressing site investigations and permitting: US$1,500,000

Other owners costs and Esaase site costs: US$ 1,100,000

Total: US$17,250,000

22.0

REFERENCES

Esaase Gold Deposit Resource Estimation, Coffey Mining Pty Ltd, April 24 2009.

Esaase Gold Project Preliminary Economic Assessment 1526-STY-001, Lycopodium Minerals Pty Ltd, May 2010.

23.0

DATE AND SIGNATURE PAGE

Name	Brian R Wolfe
Degree and Professional Association	BSc Hons(Geol), MAusIMM Post Grad Cert (Geostats)
Position	Specialist Consultant - Resources
Signature	Signed
Date	19 May 2010

Name	Harry Warries
Degree and Professional Association	Masters Degree in Mining, MAusIMM, SME
Position	Manager Mining
Signature	Signed
Date	19 May 2010

Name	Aidan Ryan
Degree and Professional Association	Bsc (Eng) Hons Metallurgy, MBA, MAusIMM, Registered Professional Engineer Queensland
Position	Principal Metallurgist
Signature	Signed
Date	19 May 2010