Technical Report for the First Disclosure of a …s1.q4cdn.com/789791377/files/doc_downloads/Wassa/wassa...Oxide Ore 3.38 1.16 Fresh Ore 6.91 1.69 Total Ore 10.29 1.51 Total Waste

Technical Report for the First Disclosure of a Mineral Reserve Estimate for a Material Property

Wassa Mine, South West Ghana

prepared for

Golden Star Resources Ltd. Denver, Colorado

submitted by

David Alexander C.Eng Bogoso Gold Limited

Bogoso, Ghana

1st August 2003

Report No. 010803/DA

Report No. 010803/DA 1st August 2003

TABLE OF CONTENTS

1 Summary .............................................................................................................................................1

2 Introduction and Terms of Reference.............................................................................................3

3 Disclaimer............................................................................................................................................4

4 Property Description and Location.................................................................................................4

5 Accessibility, Climate, Local Resources, Infrastructure and Physiography ............................6

6 History .................................................................................................................................................6

7 Geological Setting...............................................................................................................................7

7.1 Deposit Type................................................................................................................................12

7.2 Stratigraphy.................................................................................................................................12

7.3 Alteration.....................................................................................................................................13

7.4 Structural Geology......................................................................................................................13

7.5 Mineralization.............................................................................................................................14

8 Exploration........................................................................................................................................16

8.1 Drilling ........................................................................................................................................16

8.2 Sampling Method and Approach ...............................................................................................18

8.3 Sample Preparation, Analyses and Security .............................................................................19

8.4 Data Verification ........................................................................................................................21

9 Mineral Resource and Reserve Estimates....................................................................................22

9.1 Mineral Resource Estimate........................................................................................................22

9.2 Resource Estimation Parameters...............................................................................................24

9.3 Mineral Reserve Estimate ..........................................................................................................32

9.4 Evaluation of Optimum Pit Shell. ..............................................................................................36

9.5 Mine Design ................................................................................................................................36

9.6 Mineral Reserves ........................................................................................................................38

10 Mining Operations.........................................................................................................................39

10.1 Mining Operations....................................................................................................................39


10.2 Production Schedule.................................................................................................................40

11 Mineral Processing and Metallurgical Testwork......................................................................42

11.1 Metallurgical Testwork ............................................................................................................42

11.2 Process Plant Design ...............................................................................................................46

11.3 Processing Operations .............................................................................................................51

12 Infrastructure .................................................................................................................................52

12.1 Tailings Storage Facility..........................................................................................................52

12.2 Power Supply ............................................................................................................................53

12.3 Existing Infrastructure .............................................................................................................54

13 Environmental ................................................................................................................................55

14 Economic Analysis .........................................................................................................................57

14.1 Capital Cost Estimates .............................................................................................................57

14.2 Operating Cost Estimates ........................................................................................................59

14.3 Taxes and other payments........................................................................................................60

14.4 Economic Analysis....................................................................................................................61

14.5 Mine Life ...................................................................................................................................64

15 Interpretations and Conclusions .................................................................................................65

16 Recommendations..........................................................................................................................65

17 References .......................................................................................................................................66

18 Plans .................................................................................................................................................67

19 Date...................................................................................................................................................67

Golden Star Resources Page 1 of 67


Technical Report for the First Disclosure of a Mineral Reserve Estimate for a Material Property

Wassa Mine, South West Ghana

1 Summary

The Wassa mine, located in the Western Region of Ghana, is an open pit gold mining operation that started production in 1998. It is situated in an area of historic small scale mining, while large scale commercial operations started with the formation of a joint venture between the Mining Lease holders, Wassa Mineral Resources Limited, and the Irish companies Glencar Exploration Limited and Moydow Limited. The joint venture company, named Satellite Goldfields Limited (“SGL”), started an exploration drive across the Wassa Property in 1993, which culminated in a positive Feasibility Study completed in early 1998. Proven and Probable Mineral Reserves were estimated at 17.6 million tonnes at a grade of 1.7 g/t, with a contained gold content of 932,000 ounces. The study envisaged a 3 million tonne per annum heap leach operation, with annual production of around 100,000 oz over a 7 year mine life. Debt financing of $42.5 million was arranged through Standard Bank London Limited and the Commonwealth Development Corporation. Operations started in late 1998, but it soon became apparent that the gold recoveries predicted in the feasibility study could not be achieved. SGL tried and experimented with numerous methods to improve recoveries from the heap leach pads, but were unable to raise production to economically viable levels. In mid-2001, the secured senior creditors exercised security over the project, and put it up for sale. At this stage, a total of 8.33 million tonnes of ore had been mined, with a contained gold content of 444,000 ounces. Golden Star Resources Ltd. (“GSR”), amongst others, carried out a due diligence on the property, which included a preliminary drilling program to test the quoted resources. In late 2002, a final agreement was negotiated with Standard Bank to acquire the assets of SGL, which included the mining lease and all infrastructure. GSR immediately commenced an intensive exploration program, for the purposes of extending the resource, as well as obtaining representative samples for metallurgical testwork. In total, the GSR exploration program has drilled 41,071 meters, comprising 9,356m of diamond drilling, 25,360m of reverse circulation drilling, and 6,355m of reverse circulation drilling on the heap leach pads. Consultants Steffen Robertson and Kirsten were retained to carry out the geostatistical evaluation and resource estimation in March 2003, while GSR geologists continued to refine the geological model. Currently estimated Mineral Resources are defined as:



Category Cut Off

Grade (g/t) Tonnes Grade

g/t Ounces

Indicated Oxide 0.50 4,549,615 1.13 164,763 Indicated Fresh 0.60 11,542,802 1.56 579,467 Indicated Total 16,092,417 1.44 744,230 Inferred Oxide 0.50 2,114,965 1.02 69,664 Inferred Fresh 0.60 23,267,677 1.15 862,370 Inferred Total 25,382,642 1.14 932,034 Indicated HL 0.40 5,177,105 0.75 124,836

Metallurgical testwork was carried out at specialist laboratories in South Africa and Australia, and results indicated that the mineralized material from the deposit was highly suitable for conventional Carbon-in-Leach (“CIL”) processing. GSR embarked upon a full Feasibility Study, which was managed by Metallurgical Design and Management (Pty) Ltd. (“MDM”), who prepared the process flow sheet and carried out the process plant design. Open pit optimizations and mine design were carried out by GSR, with review by SRK, using a base case gold price of $300 per ounce. In conformity with the requirements of National Instrument 43-101, only the Indicated Resources were taken into account when carrying out the pit optimizations. The Mineral Reserves within the designed pit are estimated as:

Rock Type Tonnes x 106

Grade g/t

Oxide Ore 3.38 1.16 Fresh Ore 6.91 1.69 Total Ore 10.29 1.51 Total Waste 27.63 Total Mined 37.92 Strip Ratio 2.69

The feasibility study presumes the construction of a 3.5 million tonne per annum CIL plant, located adjacent to the site of the gold recovery facilities from the former heap leaching operation. The existing infrastructure will be used wherever possibly, specifically the four stage crushing circuits. Additional infrastructure required for the project includes the construction of a tailings storage facility to store the CIL tails, and the installation of a 35km 161 kV power transmission line to connect the Wassa site to the national power grid.



Construction of the process plant has started, and the two ball mills are already on site. Production is scheduled to commence in early 2004, with the first year of operations being spent reprocessing material from the heap leach pads. This is an inexpensive option for start up, and has the added benefit of releasing area for the site of the new tailings storage dam. Approximately 75,000 ounces of gold is expected to be produced, at a cash cost of $211 per ounce. Mining of ore from the open pits will start in 2005, at an annual rate of 3.5 million tonnes per annum. Ore will be blended into the process plant at a ratio of approximately one third oxides, two thirds fresh material, to produce over 150,000 ounces per annum at a cash cost of under $200 per ounce. Overall mine life, including the first year of heap leach processing, is around 4 years. The feasibility study has demonstrated that the Wassa mine is economically viable, with a positive Internal Rate of Return at gold prices above $275 per ounce. At a gold price of $325 per ounce, the project has an IRR of $27 million, with an after-tax undiscounted Net Present Value of $22 million. This is inclusive of a 3% government royalty on revenue and $3.5 million for post mining reclamation. The project has been based solely on the Indicated Resources as currently defined. There are significant Inferred Resources below planned pit bottom, and GSR has outlined a drilling program for late 2003 to test these Inferred Resources, to improve confidence in them such that they can be brought into the Indicated Resources category, and hence potentially be converted into Probable Mineral Reserves. This drilling program will be completed by end-2003, and a new Mineral Reserve estimate will be prepared in early 2004.

2 Introduction and Terms of Reference

Golden Star Resources Ltd. has made a first disclosure of a Mineral Reserve Estimate for its Wassa Mine in south west Ghana on July 10, 2003. The company is required under Canada’s National Instrument 43-101 to file a Technical Report relating to this disclosure within 30 days of the disclosure. Information used in the preparation of this Technical Report has been obtained from a Feasibility Study prepared by GSR on the property, and which includes independent reports prepared by specialist consultants in various technical fields. The sections dealing with Mineral Resources have been extracted from an independent review and Qualifying Report carried out by Associated Mining Consultants Ltd. (“AMCL”). This qualifying report can be accessed in full on Sedar (www.sedar.com) under the Golden Star Resources directory. AMCL also carried out a review of this technical report prior to filing. The author of this report is an employee of Bogoso Gold Limited, a 90% owned subsidiary of GSR, and has been intimately involved in the preparation of the Minerals Reserves estimates for the Wassa Mine. All costs are quoted in United States dollars.



3 Disclaimer

The author of this report has been intimately involved in the Wassa Feasibility Study, and this report is based largely on the findings of that study. GSR has relied heavily on specialist external consultants, recognized in their field, to oversee and review the significant aspects of the study, and use has been made of their technical reports. These reports are referenced in section 16. The author has not personally read the reports pertaining to the metallurgical testwork programs, but from discussions held with colleagues on the feasibility study team who attended much of the testwork programs, the author is satisfied that the reports do reflect a representative picture of the metallurgical characteristics of the Wassa ores.

4 Property Description and Location

The Wassa mine is located in the Mpohor Wassa East District, in the Western Region of Ghana. It is 80 km north of Cape Coast and 150 km west of the capital Accra. It is also 70 km by good gravel road from GSR’s mine and process plant site at Bogoso. The property lies between Latitudes 5º 25’ and 5º 30’ North, and between Longitudes 1º 42’ and 1º 46’ East The Wassa mining lease (No. 2033/1944) lies within the Subri-Akyempim Concession, and covers an area of 57 km2 (5,700 ha) with the southern portion being located within the Subri Forest Reserve. There are five recognized mineral prospects located within forest reserves, with Wassa being the only one for which there is a formal mining lease (as opposed to an exploration license). Despite the existence of the mining lease, prior approval is required for any activity within the forest reserve. The issue of activities within these forest reserves is being addressed at a national level, and indications are that restrictions are being eased. None of the currently defined Wassa Mineral Reserves are within the Forest Reserve, and are therefore unaffected by any such restrictions. The owner of the Wassa mining lease is a Ghanaian registered company Wexford Goldfields Limited (“WGL”), in which GSR has a 90% interest. In common with other mining companies in Ghana, WGL will pay a royalty to the Government of Ghana at a rate of between 3% and 12% on all gold sales revenues. The royalty rate applicable is defined in the Minerals Royalties Regulations 1998, and is related to the “Operating Margin” defined therein. Provided that the Operating Margin is less than 30%, the standard rate of 3% does not increase. For every 1% increase of the Operating Margin above 30%, the royalty rate rises by 0.225% up to a maximum of 12% Upon acquisition of the property from the lenders, two additional royalties were put in place as part of the acquisition consideration. The first royalty varies with the gold price and is to be paid quarterly. The amount of royalty is determined by multiplying the produced ounces from Wassa for each preceding quarter by US$7.00. The royalty rate is subject to a US$1.00 increase for every US$10.00 increase in the average market gold price over the period above US$280/oz, up to a maximum of US$15.00 per ounce reached at a gold price of US$350.00 per ounce.



The second royalty is also payable quarterly on the same produced ounces at US$8.00 per ounce, up to a cumulative maximum of US$5.5 million. On 24th July 2003, GSR announced that it had reached agreement with Standard Bank London Limited (“SBLL”) to buy back these royalties for a sum of US$11.5 million. In addition, this purchase price buys back a US$1.9 million loan facility that was in place with the bank, repayable over 5 years at an interest rate of LIBOR + 2.5%, and also releases any securities that the bank has over the Wassa Mine. This represents a 20% discount of the current face value of the debt. The agreement is subject to typical approvals and conditions.



5 Accessibility, Climate, Local Resources, Infrastructure and Physiography

The project area is characterized by gently rolling hills incised by an extensive drainage network. The area is relatively wet, with many low lying swampy areas. Extensive subsistence farming occurs throughout the area, with plantain, cassava, pineapple, maize, and cocoyams being the principle crops. Some small scale cultivation of commercial crops is also carried out, with cocoa, teak, coconuts and oil palm being the most common. With the exception of the forest reserve, there is little primary forest remaining, the area being mainly secondary regrowth. The Wassa project area falls within the semi-equatorial climatic zone of Ghana. The climate is characterized by seasonal weather patterns, involving a double wet season in April to June (major) and October to November (minor), and a main dry season between December and March. Average annual rainfall is 2,030mm per annum.

6 History

The Wassa area has witnessed several eras of local small scale (galamsey) and colonial mining activity from the beginning of the century, and mining of vein structures are evident from the numerous pits and adits covering the Wassa lease area. In recent times, the property was operated since 1988 as a small scale gravity circuit by a Ghanaian company, Wassa Mineral Resources Limited. In 1993, Wassa Mineral Resources were looking for a capital partner to further develop the mining lease, and invited the Irish companies Glencar Exploration Limited (“Glencar”) and Moydow Ltd to visit the concession. Following this visit, Satellite Goldfields Limited (“SGL”) was formed between Wassa Mineral Resources, Glencar and Moydow Ltd. The mining lease, which is valid for a 30 year period expiring in 2022, was assigned by Wassa Mineral Resources Limited to SGL. Extensive satellite imagery and geophysical interpretations were carried out, which identified a strong > 1 g/t gold target. Exploration drilling commenced in February 1994, and by March 1997, 58,709m of reverse circulation and diamond drilling had been completed. In September 1997, consulting engineers Pincock, Allen and Holt had completed a feasibility study, which determined a proven and probable mineable reserve of 17.6 million tonnes at 1.7 g/t, for a total of 932,000 contained ounces. Construction of the Wassa Mine was initiated in September 1998 after Glencar secured a US$42.5 million debt-financing package from a consortium of banks and institutions, primarily SBLL and the Commonwealth Development Corporation (“CDC”). The Wassa Mine was developed as a 3mtpa open pit heap leach operation with a forecast life of mine gold production of approximately 100,000 ounces per annum. The first ore from the pit was mined in October 1998. After approximately one years production it became evident that the predicted heap leach gold recovery of 85% in the oxide ore, could not be achieved, mainly due to the high clay content of the ore. After a number of attempts to improve on the recovery, including increased agglomeration and doubling the



leach solution application rate, it was concluded that the achievable gold recovery on oxide ore by heap leach, was between to 55% - 60%. Recoveries within the fresh material were better, but still below predicted levels. The combined effect of the lower than planned gold recovery and lull in the gold price at the time resulted in the company not being able to service its debt to the banks. In early 2001 the banks together with Glencar decided to sell the project to recover some of the accumulated debt. Mining was stopped at the end of October 2001 and irrigation of the heap leach with cyanide solution continued until March 2002, after which rinsing of the heaps with barren solution continued until August 2002. As at October 2002, 8.3mt of ore had been stacked at an average grade of 1.66g/t amounting to 444,5k oz gold. Actual gold production on the property to date has been 256,8k oz. Based on this the heap should still contain 188k oz of gold at average grade of 0.71 g/t. When the secured senior creditors exercised security over the project in 2001, they put the project up for sale with the issuance of an Information Memorandum dated May 2001. GSR was invited amongst other parties to conduct a due diligence on the operation, and in November, negotiations were started to acquire the Wassa assets. As part of a final due diligence on the resources, GSR undertook a structural evaluation and drilling program between December 2001 and April 2002. Upon completion of the acquisition of Wassa Mine by GSR, a further exploration program was undertaken. Both these exploration programs form part of a Feasibility Study that was completed in July 2003 which demonstrates the economic viability of reopening and expanding the existing open pits, and processing the ore through a conventional Carbon-in-Leach (CIL) circuit. Current ownership of the Wassa Mine is through a Ghanaian registered company, Wexford Goldfields Limited. Golden Star Resources, through a 100% interest in Caystar Holdings and its subsidiary Wasford Holdings (both Cayman Island corporations), holds a 90% interest in Wexford Goldfields Limited, with the remaining 10% being held by the Government of Ghana.

7 Geological Setting

The Wassa gold mine is located on the southeastern limb of the Ashanti Trend, a prominent north-east trending greenstone belt that extends for over 240 km within the Man Shield of the West African Craton. The belt takes the form of a synclinorium developed within lower Proterozoic sedimentary and volcaniclastic rocks of the Birimian Supergroup and the Tarkwaian Group, which are important hosts for gold. Figures 3.0 and 3.1 (Map Pocket at rear) are illustrations of the general geological setting of the Ashanti Trend in southwest Ghana and the Wassa Mine. Figures 3.2-3.4 illustrate the regional setting based on satellite imagery and the general geological setting of the mine in plan and section. Gold mineralization in the Birimian rocks of Ghana are concentrated along four, parallel corridors (greenstone belts), 10 km-15 km in width and several hundred kilometres in length. A fifth belt of gold mineralization is represented by the Tarkawaian rocks. Regional scale deformation was dominated by folding and thrusting focused at the boundary of the volcanic belts and sedimentary basins. The Ashanti Trend is one of four sub-parallel greenstone belts and hosts the principal deposits of the Ghanaian gold belt. The trend is closely aligned with a major thrust fault (later reactivated with sinistral



wrench movement) which separates the meta-sedimentary and meta-volcanic units of the Birimian and the clastic rocks of the Tarkwaian. The northwestern limb of the Ashanti Trend parallels a regional shear zone (the Ashanti shear zone). The Birimian comprises an assemblage of turbiditic sedimentary (phyllites, schists and greywackes), and volcaniclastic rocks deposited in shallow marine basins (Lower Birimian) separated by a sub-parallel series of north-east trending volcanic belts (Upper Birimian). The transition between volcanic belts and sedimentary basins is marked by chemical sediments including cherts, manganese and carbon-rich sediments. Conventional thinking, based on relative stratigraphic position, suggested that the volcanic suites overlay the Lower Birimian. Recent radiometric work suggests that, in fact, that the volcanic rocks are 50 Ma-60 Ma older than the sedimentary sequences and may have been thrust faulted into their present position. It is most likely that the Upper and Lower Birimian represent coeval, lateral facies, equivalents separated by transition zones containing chemical sediments. A thermal deformation event resulted in regional metamorphism of the Birimian and Tarkwaian rocks. The Eburnean tectono-thermal event has been interpreted as occurring as a period of Birimian volcanic eruption, intrusion of granitoids and a period of metamorphism, uplift and erosion (Eburnean 1; 2240 Ma-2150 Ma) followed by regional metamorphism of both the Birimian Supergroup and Tarkwa Group rocks with further intrusive activity (Eburnean 2; 2150 Ma-2130 Ma).Emplacement of sub-volcanic plutons such as occurred in the Eburnean tectono-thermal event around 2.1 billion years ago may have contributed to the formation of late, discordant epigenetic veins, vein systems and stockworks in the Birimian. Within the Birimian and Tarkwaian, NW-SE (principal maximum stress ±100°) compression or shortening is related to a single, long-lived progressive deformation event resulting in north-east tending westerly dipping (30°-60°) thrusts such as the Axim-Konongo thrust along which occur the Prestea deposits. These faults were subsequently re-activated on a local scale as sinistral strike-slip faults which control the presence and distribution of the mineralization at Prestea. Regionally, up to five phases of deformation has produced polyphase fold patterns in the region. Deformation involved Birimian, Tarkwaian and early intrusives resulting in a thrusted and folded orogenic belt. Imbricate slices of metavolcanics occur within the central structural corridor. A lower stress regime further away form the volcanic-sedimentary contact has resulted in broad open folding dominating the Tarkwa Group. The thrusts were reactivated during a change in the principal stress direction resulting in sinistral wrench faulting. Dextral relaxation structures have been noted. Gold appears to have been emplaced during a period of sinistral movement. Tarkwaian fluvial molasse sediments, principally, conglomerates (which host major gold deposits elsewhere in Ghana) are likely derived from erosion of the Birimian rocks which resulted in palaeo- placers similar to the Witwatersrand Basin of South Africa. Recent radiometric data suggests Tarkwaian deposition in an inter-montane graben was separated from the last Birimian volcaniclastic episode by a short period of extensional tectonism and block-faulting.



Gold deposits within the Birimian include Prestea (Golden Star Resources and Prestea Gold Resources), Bogoso (Golden Star Resources), Wassa, Obuasi (Ashanti), Ayanfuri, Amansie, Yamfo and Konongo. Gold deposits within the Tarkwaian include Teberebie, Abosso, Tarkwa, Iduapriem and Akyempim

234

SUBRI

(Toi)

Toe

MP11

G-27

Apia-Essuman

Subri-Manda

Biter Camp

Subiri

MP12

Bonsa

177

Subi

ri

G-27

subi

ri

Subiriq

79G-27

176G-27

PeterCamp

EstateRubber

Subri

78G-27

Essienkyem

Nsadweso

Odumase

Asumenan

Kubekro

235G-27

Akyempim

236

Kobeda

Dwaben

175

Nankadam

Nankadam

Apetetwum

NewtownAccra

FOREST

RESERVE

Bet

Sawbimawoba

Adeiye

Ampoma

80

CompoundAteiku

Apetetwumso

HaltSaponso

MP36

81

174

Adaase

B LOCK

EAST

BEN

Saponso

Adahasu

Daman

Bepontam

SuponAkokruaa

Atire(Ahire)

Onyamebekyere

Amponsakrom

5,000 10,000 15,000

5,000

10,000

15,000

20,000

Wassa Pits

MINE LEASE

EXPLORATION LICENCE

0 2000METRES

WEAK LINEAMENT

PIT FOOTPRINT

STRONG LINEAMENT

legend



7500 N

8000 N

8500 N

9000 N

8500E

9000E

9500E

10000E

MAIN 2

STARTER

MAIN 1 - B SHOOT

MAIN 1 - F SHOOT

DEADMANS HILL

MIDDLE ZONE

40

75

80 SOUTH EAST

419 ZONE

A

A A

0 500metres

250

Siliceous Phyllites

Fine Grained Felsites

Greenstone

Predominantly Chloritic Phyliites with thin interfoliated felsites

Medium to Coarse Grained Felsites (Rhyolites)Siliceous PhyllitesChloritic Siliceous PhyllitesBMU

legend



ATOPO SURFACE

OXIDE-FRESH INTERFACE

MAIN 2 MAIN 1 - B SHOOT SOUTH EAST

Siliceous Phyllites

Fine Grained Felsites

Greenstone

Predominantly Chloritic Phyliites with thin interfoliated felsites

Medium to Coarse Grained Felsites (Rhyolites)Siliceous PhyllitesChloritic Siliceous PhyllitesBMU

legend



7.1 Deposit Type

The deposit is classified as a classic greenstone-hosted quartz-carbonate vein lode gold deposit. The current exploration model has focused on structural controls on the emplacement of epithermal gold mineralization possibly related to the nearby Cape Coast or Dixcove granitic intrusions. This would suggest a "one-off" hydrothermal event. The gold mineralization appears to be locally remobilized into discontinuous, boudinaged quartz veins within a volcano-sedimentary package as a result of regional metamorphism to greenschist and lower amphibolite facies. Subsequent deformation has extensively, and preferentially, disrupted the thin sequence. It would appear that gold has not been moved far from its primary depositional source which may account for the overall low grade of the deposit. The veins contain coarse, particulate gold (a so-called nuggety deposit) with erratic distributions.

7.2 Stratigraphy

Megascopic structure in the Wassa mine area is interpreted to be an a synformal anticline based on the interpretation that mafic lithologies in the core may represent older rocks. The dominant rock types are altered meta-basalts and meta-phyllites with local chemical sediments as well as a feldspar porphyry. A mafic volcanic package in the core is surrounded by a more felsic volcanic unit. Age relationships amongst the various units have not been determined. From the core outwards the field classification of these lithologies is as follows: 7.2.1 Mafic Volcanic Unit: Referred to as "basalts" by Wassa exploration staff. These are mafic volcanic flows where the primary volcanic textures have been strongly overprinted by later hydrothermal alteration. A schist which mostly consists of second generation minerals, chlorite, sericite and carbonate (dolomitic to possibly ankeritic). Locally, silicified, fine grained and relatively massive light grey to medium green. (Light green Figure 3.2). This unit is host to mineralized quartz vein shoots in F shoot and B shoot, Starter Pit, 242 zone and Main 2 North zone. This mafic unit is cross-cut by probable mafic syn-volcanic dikes and sills which are locally referred to as "diorite". (purple colour on Figure 3.2). Coarse grained, equigranular, with chlorite and sericite derived from the hydrothermal alteration of amphiboles and plagioclase. Texturally this unit is very different from the enclosing more massive mafic volcanics and is most likely a gabbro or norite. 7.2.2 Banded Magnetic Units These units is intercalated within the mafic volcanics and consists of banded magnetite with thin detrital or chemical sediments. Unit is generally 2 m to 5 m wide. At least 2 separate units have been



identified, which are discontinuous along strike (grey colour on Figure 3.2). One is more extensively developed. 7.2.3 Phyllite Metasediments intercalated within the mafic volcanic units. Quartz rich, varying from wacke to arenite depending on the chlorite content. (yellow colour on Figure 3.2). 7.2.4 Felsic Volcanics Felsic volcanic flow package lies further from the axial plane and consists of a felsic flow with primary textures overprinted. The composition is less chloritic and more quartz rich than the mafic volcanic unit but is also primarily composed of carbonate and sericite. (light brown on Figure 3.2). 7.2.5 Felsic Porphyry A felsic quartz feldspar porphyry (quartz rich matrix) which appears to have intruded the felsic volcanics (nowhere has field mapping shown this unit to be in contact with the mafic volcanics) and is most likely a second syn-volcanic intrusive (light blue on Figure 3.2). This is the host to the South East ore deposit. 7.2.6 Graphitic Mudstone Thin graphitic units are intercalated within the felsic volcanics which may represent detrital or chemical sediments emplaced during a hiatus in the felsic volcanic event.

7.3 Alteration

Pervasive alteration has destroyed the primary textures of all rock types at Wassa. Chlorite and sericite are common with extensive silica and dolomitic(ankeritic) alteration. Tourmalinazation of certain quartz veins has occurred.

7.4 Structural Geology

Three phases of ductile deformation (shortening) are interpreted to occur locally at Wassa. Regionally, up to five events have been observed within the Ashanti trend at Obuasi. 7.4.1 D1 Event A primary D1 ductile deformation event has been inferred from observed F2 re-folding of an S1 foliation. Alteration of the rocks is generally so pervasive that no primary textures are readily discernable and it is impossible to determine whether the S1 foliation is bedding parallel. No F1 folding has been observed.



7.4.2 D2 Event F2 folds are generally rare. The best preserved examples occur in the Mid-East and Deadman’s Hill areas. They are isoclinal and generally asymmetric. The only symmetric F2 fold observed to date is at the west end of the Mid-East area. The symmetric fold is an M type fold plunging at approximately 45° to the south. The presence of this fold together with nearby asymmetric S type and Z type F2 folds suggests that the stratigraphic sequence of the Mid-East area (consisting of four sub-parallel horizons of mafic volcanic rocks interlayered with felsic rocks) may be structurally repeated by isoclinal F2 folding. The Mid-east ore zone may itself lie along the axial trace of this F2 fold. Mapping by the mine geological staff has led to the clear definition of the orientation of a penetrative S2 foliation. This follows an arcuate pattern around the current mine site and is particularly observed in the Starter Pit. 7.4.3 D3 Event A late, post-Tarkwaian thermal event resulted in metamorphism to low amphibolite facies. Predominantly compressional shorting resulted in thrusting and the development of megascopic F3 folds. These folds are reflected in the map pattern of the mine site. F3 folds are asymmetric and mostly reclined to the west in the direction of the Tarkwaian belt. F3 folds can be seen as large westerly cascading folds. Most small-scale F3 folds have generally steeply dipping to sub-vertical eastern limbs with the western limb more shallowly dipping at approximately 30° to the east- southeast. F3 folding is the most obvious fold style at Wassa affecting all lithologies and kinking the S2 foliation throughout the mine site. The F3 fold is apparent in the predominant synform whose axial trace follows the curve of B Shoot pit towards Deadman Hill. The intersection between an F3 axial planar S3 crenulation cleavage and S2 foliation has produced a very clearly developed intersection lineation that tends to plunge generally southwards with a rake of approximately 30°. 7.4.4 D4(?) Event A possible D4 brittle deformation event may have resulted in late dip-slip faults observed in B Shoot and Starter Pit cross-cutting the F2 and F3 folds and the QV3 veins.

7.5 Mineralization

Gold mineralization at Wassa occurs in separate zones of discontinuous quartz vein material located in the core of a south-westerly plunging fold structure. Three generations of quartz veining have been observed, referred to as QV1, QV2, and QV3. All veins are predominantly quartz-carbonate (dolomite ± ankerite) and occur in close proximity to each other. QV1 are characterized by the presence of rare, small to large blocky tourmaline masses and small blebs. QV1 veins contain massive, fibrous, translucent to smoky quartz. Affected by F2 folding, the veins have



been fragmented and in many cases boudinaged during the D2 event. This vein sequence is gold mineralized and contains both framboidal and euhedral pyrite crystals which have been stretched or partially transposed along S2 foliation. This suggests that the mineralizing event at Wassa is early, possibly synchronous with the D1 event. Bardoux (2002) constructed a preliminary structural analysis of the Wassa area. In early, draft versions of his report, he concurred with this interpretation of the timing of the mineralizing event but has subsequently modified his opinion to suggest that the mineralizing event occurred in D3 time. This is based on the belief that the three veins systems are related to one hydrothermal event, and since the QV3 veins are observed to cross-cut F3 folds, the mineralizing event must be synchronous with the D3 event. AMCL concurs with the current interpretation of the Wassa exploration staff that there are two to three generations of quartz veining. The QV3 veins must be later than the F3 folding since they cross-cut these folds but dilation to permit fluid flow would only occur during relaxation from the compressional event. QV2 veins are similar texturally to the QV1 veins but do not appear to have been affected by F2 folding and do not appear to contain tourmaline. QV1 and QV2 veins occur n close proximity to each other and may be difficult to distinguish in the field. The QV2 veins may contain traces of remobilized gold but are not the main source of gold at Wassa. Late QV3 are texturally quite different with a milky, pink colored tinge and are observed to cross-cut both QV1 veins and F3 folds. Only low grade (= 1 g/t) gold mineralisation has been detected in these veins. The veins are not continuous, occurring as disconnected fragments or swarms of quartz-carbonate vein remnants, boudines and blocks. Ore bodies are related to vein fragment spatial densities and the presence of sulphides, predominantly pyrite, with rare chalcopyrite. In veins such as these, with an extended structural history, gold-rich elements are frequently disrupted or dissected by later structures and as a result, present unique sampling problems due to discontinuous gold distributions. The structures that disrupt the vein continuity may have also disrupted the smaller internal elements of the vein. This implies difficulty in estimating the grade of the resource with any degree of precision. Gold content variograms prepared by SRK suggest a lack of continuity in that element. The mineralized zone containing the quartz does, however, show geological continuity at depth and long strike and it possible to interpret these zones as solids in the modeling software. Continuity, both grade and geological is a critical factor in resource estimation. In coarse gold deposits, geological continuity will be much greater than grade continuity. The geological interpretation of vein geology including mineralogy and structural history is key to understanding the grade distribution for mine planning. Geological complexity in coarse gold deposits is active at three levels:

• Internal distribution in the host structure is controlled by vein thickness and textural development of vein filling components giving an irregular gold distribution pattern;

• primary dimensions and shape of the host veins which is controlled by fracture patterns developed in the host lithology; and,



• relationship between the carrying structure and the vein. Mineral and textural assemblages are an important guide to correlation. Timing of various structures within the deformation sequence may also aid correlation. Geological mapping is key and will aid in resolving the timing of the QV emplacement which will in turn assist in understanding the grade distribution. A zone of oxidation occurs from the tops of low hills down to the valley floor. A thin transition zone overlies the sulfide or fresh mineralization. Historically, gold production has largely been from the oxide ores. The above section has been extracted from the Qualifying Report No. 03PM67, dated April 2003, prepared by consultants AMCL. This report can be viewed on Sedar (www.sedar.com) under the Golden Star Resources directory.

8 Exploration

8.1 Drilling

Exploration work carried by Satellite Goldfields between 1992 and 1997 delineated a resource defined by 86,980 m of reverse circulation drilling (1,072 holes) and 12,700 m of diamond drilling (136 holes). Pit delineation drilling has been carried out both in and around the Wassa pit area. A total of 34,250m of drilling has been carried out in 523 holes since the feasibility study was completed in September 1997. The current phase of exploration was initiated in November, 2001 with an extensive reverse circulation (12,200 m in 101 holes) and diamond (7,900 m in 96 holes) drilling carried out between the second half of 2002 and early 2003. The current program used two drilling contractors, Geodrill Limited and Pontil Minerex Limited with two Universal Drill Rigs, a KL900, a KL600 and a Longyear 38. This can be summarized as follows:

• November 2001: Geological compilation, pit mapping • December 2001: Structural geology evaluated • January 2002: Preliminary solids modeling and resource estimate • February 2002: Further structural evaluation as well as a preliminary visit by SRK

to advise on resource modeling • March/April: Phase 1 Drilling, summarized as follows:



Table 8.1: Phase 1 Drilling

Reverse Circulation Drilling

Reverse Circulation Collars

Diamond Drilling (Diamond Tails) Zone No. Holes Length

(m) Length (m) No. Holes Length (m)

Main 2 North Starter Pit B-Shoot

F-Shoot Dead Man’s Hill 7 595.0 151.7 14 1,254.0

Phase I*

Sub-Total 7 595.0 151.7 14 1,254.0

*As part of the Phase 1 activities 43 rotary air blast holes (1,092 m) were also completed.

• May, 2002: Completion of preliminary "in-house" resource estimate • May-August, 2002: No activities on site while acquisition agreement is renegotiated • August 20, 2002-

November 1, 2002: Golden Star staff return to site and commence a Phase 2 drilling program summarized as follows:


Reverse Circulation Drilling

Reverse Circulation Collars

Diamond Drilling (Diamond Tails) Zone

No. Holes Length (m) Length (m) No. Holes Length (m) 242 8 759 196.3 4 325.8 Main 2 South 1 110.0 Main 2 North 1 92 7 740.5 419 1 110 111 3 333.4 South East 8 808 271.7 9 728.8 Mid-East 1 106.8 128.2 3 266.3 Starter Pit 8 660.2 B-Shoot 115.2 15 1,444.6 F-Shoot 8 734 113.4 5 391.9 Dead Man’s Hill 11 952 244.4 3 298.1

Phase 2 1,2,3,4

Sub-Total 38 3,561.2 1,180.2 58 5,300.0

1: As part of phase 2 drilling 250 rotary air blast totaling 6,500 m were completed. 2: 321 4 m deep auger holes totaling 990 m were completed for geochemical sampling. 3: 825 1 m deep geochemical test holes were completed at Juabem and Nsadweso. 4: 467 heap leach pad evaluation holes totaling 6,355 m were completed.



• October-January: A Phase 3 drilling program was completed as follows:


Reverse Circulation Drilling Reverse Circulation

Collars

Diamond Drilling (Diamond Tails)

Zone No. Holes Length (m) Length (m) No. Holes Length (m)

242 15 1,025 317.0 DT 193.2 Main 2 South Main 2 North 11 924 119.9 2 117.4 419 21 1,099 633.6 DT 803.9 South East 52 3,442 1,375.5 DT 694.3 Mid-East 8 276 310.3 DT 159.0 Starter Pit 15 650 182.6 DT 198.8 B-Shoot 20 1,020 427.6 2 344.1 F-Shoot 15 971 516.6 DT 72.1 Dead Man’s Hill 17 1,634

Phase 3 1

Sub-Total 174 11,041 3,883.1 4 2,583

1 65 reverse circulation condemnation drill holes totaling 4,946 m were completed as part of this phase. A summary of GSR exploration on the property to date is as follows: Table 8.4: Exploration Summary Exploration Type No. Holes Meters Reverse Circulation Drilling 173 15151 Reverse Circulation Collars 72 5263 Diamond Drilling (Diamond Tails) 122 9356 Reverse Circulation drilling (Heap Leach Pads) 467 6355 Reverse Circulation drilling (Condemnation) 65 4946 Rotary Air Blast 18710 Soil 765 825 Deep Auger 825 1000 Trench 1 21

8.2 Sampling Method and Approach

Primary sampling is dependent upon the drilling method used to obtain the sample. Reverse circulation (RC) cuttings are sampled on a 1 m interval. 3 m composites were then prepared for assays. Rotary air blast (RAB) holes are also sampled on 1 m intervals with these composited to 3 m for the initial



analysis. If the results indicate a grade in excess of 0.20 g/t then individual 1 m splits are re-analyzed and numerically (not analytically) composited. The NQ (47 mm) diamond drill core is split in half at the mine site and sampled on a lithological basis. The core is crushed at the laboratory in a jaw crusher. The total number of samples collected by Golden Star as part of their due diligence and exploration is summarized as follows: Table 8.5: Sample Summary

RC/RCC Drilling 1m 3m Total

Diamond Drilling (Diamond Tails)

Replicates Standards

242 605 787 1392 415 96 95 Main 2 South 47 3 3 Main 2 North 384 382 766 734 79 77 419 1,099 651 1,750 1,046 147 139 South East 2,340 2,020 4,360 1,299 305 298 Mid-East 298 211 509 316 33 31 Starter Pit 370 296 296 826 32 67 B-Shoot 27 576 503 1,447 118 120 F-Shoot 515 696 1,211 481 94 97 Dead Man’s Hill 616 1,202 1,818 643 138 140 Total 6,254 6,721 12,605 7,254 1,045 1,067 RAB 2,719 7,773 898 867 Heap Leach Pads (3m & 6m)

1,450 107 106 Condemnation 1,021

Soil 1,650 Deep Auger 1,364 All samples were sent Transworld Laboratories Ghana Limited or the mine laboratory at Bogoso/Prestea.

8.3 Sample Preparation, Analyses and Security

8.3.1 Sample Preparation Reverse circulation samples are passed through a three stage cascade riffle splitter at the drill site to reduce the mass of the sample to approximately 2 kg. Two RC drill diameters were used in the exploration program, 5¼" and 4.5", producing 38 kg and 28 kg of sample for each 1 m interval, respectively. 3 m composites were prepared by spear sampling the coarse reject bags from the appropriate 1 m interval samples. The 2 kg samples are pulverized. Rotary air blast (RAB) samples are treated in a similar fashion. The diamond drill core is logged at the Wassa core facility. Sample intervals selected based on lithology and hence may vary in length. The core is sawn at the facility and then shipped to the



laboratory where the ½ NQ diamond core is jaw crushed. A 2 kg sample is split from the jaw-crushed core and then pulverized. The coarse reject remnant (if any-sample intervals of 0.5 m would be about 2 kg and hence pulverized in its' entirety) is returned to the Wassa core facility. All samples are collected in clear plastic sample bags with redundant labelling. Samples are shipped by truck to the laboratory where they are logged in to a secure storage area. 8.3.2 Analyses The majority of samples were assayed at a commercial laboratory facility, Transworld Laboratories Ghana Limited in Tarkwa. Some samples were shipped to the Golden Star mine laboratory at Bogoso/Prestea. In previous visits to Ghana AMCL staff have visited both the Transworld laboratory in Tarkwa and the Bogoso Gold Limited laboratory at the Bogoso/Prestea mine site. As well, we have reviewed in detail, internal QA/QC procedures. The internal laboratory procedures are those you would expect to find in place at any reputable facility. The samples are not subject to any specific security procedures. RC samples are stored at the drill site for some time. Core is transported to a secure facility at the mine offices and laboratory samples are stored in a secure facility. Sample pulps are prepared at the laboratory where a 50 g charge is extracted for fire assay with instrumental finish with atomic absorption spectrometry. The Bogoso/Prestea laboratory re-assays any result >2 g/t from pulps. It is unclear whether Transworld follows the same procedure or whether any checks on high-grade samples using a gravimetric finish are conducted. 8.3.3 Standards Program The Wassa exploration staff have a detailed program for internal data verification which was developed largely for the Bogoso/Prestea mine site. Commercial (Australian Gannet) and in-house standards are used at both the in-house laboratory at the Bogoso Gold Limited Bogoso/Prestea mine site and inserted into sample sets sent to the commercial laboratory in Tarkwa. Both rejects and replicates are used in assessing analytical data quality. Two control samples (1 replicate and 1 standard) were submitted with every 20 samples. Blanks are not run. 8.3.4 Replicate Program Reverse circulation replicates for 1 m sample intervals were prepared by re-splitting the coarse reject at the drill site and inserting the replicate into the sample stream. The replicates for the 3 m composite samples were obtained re-sampling the coarse rejects with the spear sampler. Every eighteenth sample is replicated. Pulp reject samples are returned to Wassa where, 1 in 10 are renumbered and re-submitted to the laboratory for analysis. These are referred to as RC rejects.



Rotary air blast samples are stated to be subject to similar QA/QC procedures as the RC samples although pulp samples are not resubmitted. For the NQ diamond core the remaining ½ core is quartered and the replicate sample submitted to the lab with a new number. This does not represent a true replicate as the sample is obtained from a spatially different piece of core. The replicate sample size is smaller, based on ¼ core rather than the original sample size of ½ core. As well, coarse jaw-crushed reject material is returned to the core storage facility where 1 in 10 random samples are selected, re-bagged and re-submitted to the laboratory with a new number. Although diamond drill core pulps are available from the laboratory they are not utilized as part of the QA/QC program. A large amount of analytical quality control data has been collected over the course of the current exploration campaign. Unfortunately the data has been reviewed only briefly and irregularly to try to identify trends which could suggest any potential laboratory issues. Regular examination of the data with statistical analysis may have identified some sampling issues which are addressed in the following section. To be fair to the Wassa exploration staff, recent exploration and due diligence programs have been completed under time constraints which made detailed analysis of the QC data difficult. As well, an independent consultant reviewed the QA/QC program in February and did not appear to suggest any major changes in data collection procedures were required.

8.4 Data Verification

SRK completed a review of the Wassa analytical quality assurance procedures and quality control data (QA/QC) as part of the resource estimation report. This included a small test program to confirm suspected sampling problems. AMCL discussed aspects of the QA/QC procedures and data with Wassa mine staff and SRK. As we substantially concur with the conclusions and recommendations reached by SRK, no additional analytical data verification was deemed necessary by AMCL. AMCL did examine, in detail, core from each of the resource areas and compared the lithological logs with sample intervals and analytical results. Analytical results matched to the sample intervals and appeared appropriate for the lithology described. The field data itself was in good shape with all requested records readily retrieved by the exploration staff. The following discussion summarizes the SRK findings. SRK has identified issues with the reproducibility of sample analytical data. The reproducibility issues (including differences in pulp splits) are consistent with the presence of coarse particulate gold. The presence of coarse (nuggety) gold and the generally low-grade nature of the deposit exacerbates problems in approximating the real grade of the deposit. A significant difference between the precision of the analytical data from the reverse circulation samples and diamond drill core samples has been identified with the diamond drill core samples having



a much lower correlation coefficient. In the opinion of SRK the difference in precision is directly attributable to the riffle splitting of the jaw-crushed diamond core samples prior to pulverization. SRK also suggests that the small size of the reverse circulation samples (2 kg) may result in larger grade variance within this population than is evident from the data. Sampling theory would suggest that a systematic understatement of the grade may be possible under these conditions. SRK recommends ½ NQ diamond core samples should be pulverized in their entirety prior to splitting. Single splitters with repeat splitting should be adopted rather than three stage splitters. In coarse-gold deposits traditional fire assays using small charge sizes (25 g-100 g) consistently understate assays using large charge sizes (>1 kg-3 kg), metallic screen fire assays or bulk leach extractable gold (BLEG) analyses. SRK has recommended the use of an Australian proprietary leach accelerant (Leachwell®) in conjunction with BLEG analyses. AMCL is not familiar with the use of this particular additive, however, tests using metallic screen fire assays and BLEG analyses (with and without the accelerant) are strongly recommended. In AMCL's opinion, bulk sampling may be the most appropriate methodology for estimating the grade of individual ore shoots. Bulk samples should be restricted to a single geological and grade domain. Bulk samples can be readily treated through a pilot plant facility. (See Section 11.1 Metallurgical Testwork). The above section has been extracted from the Qualifying Report No. 03PM67, dated April 2003, prepared by consultants AMCL. This report can be viewed on Sedar (www.sedar.com) under the Golden Star Resources directory.

9 Mineral Resource and Reserve Estimates

9.1 Mineral Resource Estimate

In a report dated February 2003, Steffen, Robertson and Kirsten (South Africa) (Pty) Ltd. reported the following resource estimates: Table 9.1: Oxide Resource Estimate

Gold Content1 Mining Zone Resource

Classification Cut-off

Grade (g/t) Tonnes Grade (g/t) kg Ounces (troy) Indicated 0.5 921,176 1.08 993 31,919 South East Inferred 0.5 1,135,092 1.12 1,270 40,845 Indicated 0.5 727,105 0.99 719 23,125 F-Shoot Inferred 0.5 1,089,179 0.86 936 30,086 Indicated 0.5 159,069 1.33 212 6,800 B-Shoot Inferred 0.5 128,560 1.27 163 5,241



Indicated 0.5 1,028,261 1.20 1,234 39,659 Dead Man’s Hill Inferred 0.5 944,987 0.98 926 29,766

Indicated 0.5 642,670 1.39 900 28,933 242 Inferred 0.5 616,987 1.09 674 21,666 Indicated 0.5 - - - - Mid-East Inferred 0.5 1,291,821 1.33 1,723 55,398 Indicated 0.5 3,485,281 1.16 4,057 130,437 Totals Inferred 0.5 5,206,626 1.09 5,692 183,001

1 The contained gold contents may vary slightly from those reported by SRK due to rounding. Table 9.2: Fresh Resource Estimate

Gold Content1 Mining Zone Resource

Classification Cut-off

Grade (g/t) Tonnes Grade (g/t) kg Ounces (troy) Indicated 0.64 575,751 1.28 737 23,696 South East Inferred 0.61 9,374,735 1.06 9,894 318,107 Indicated 0.58 1,449,741 1.54 2,232 71,758 F-Shoot Inferred 0.61 3,799,632 1.33 5,040 162,029 Indicated 0.68 1,841,533 1.81 3,330 107,054 B-Shoot Inferred 0.51 2,973,577 1.66 4,947 159,054 Indicated 0.61 477,095 0.99 472 15,177 Dead Man’s

Hill Inferred 0.61 2,815,074 0.95 2,670 85,828 Indicated 0.61 3,329,583 1.90 6,317 203,112 242 Inferred 0.61 7,661,333 1.22 9,323 299,741 Indicated 0.61 - - - - Mid-East Inferred 0.61 1,107,691 1.03 1,144 36,785 Indicated 0.61 7,673,704 1.71 13,088 420,797 Totals Inferred 0.61 27,732,043 1.19 33,017 1,061,542

1 The contained gold contents may vary slightly from those reported by SRK due to rounding. The "fresh" resource consists of non-refractory sulphide. The total oxide and fresh mineral resource in the indicated category is:

11,158,985 tonnes @ 1.54 g/t for a contained gold content of 17,145 kg or 551,233 ounces. This corresponds to the "mineralized material" of American usage. The total oxide and fresh mineral resource in the inferred category is:

32,938,669 tonnes @ 1.18 g/t



for a contained gold content of 38,709 kg or 1,244,523 ounces. No measured mineral resources have yet been identified. AMCL would caution that this resource estimation should be considered as a "global estimate". The volume or tonnage has been determined to a reasonable level of confidence, however, sampling problems and erratic grade distribution would suggest that the grade has not. In this type of deposit, lacking grade continuity, AMCL would prefer to see a possible grade range as well as the grade estimate, however, we are of the opinion that the estimate is conservative and that the reported grade may represent a minimum value. Cut-off grades were determined using estimated mining costs and recoveries, preliminary metallurgical recoveries and processing costs, estimated general and administrative costs and a US$325/oz gold price. AMCL has determined that the resource estimate conforms to the terminology for resource and reserve estimation developed by the Canadian Institute of Mining, Metallurgy and Petroleum as required under NI 43-101. 90 % of the contained gold resource is attributable to Golden Star Resources Ltd. with the remainder attributable to the Government of Ghana.

9.2 Resource Estimation Parameters

The resource estimate is based on a geostatistical model developed in Gemcom® software. The following discussion is summarized from SRK Report No. 315917, Mineral Resource Estimate for the Wassa Deposit, Ghana, dated February, 2003. Golden Star geologists have modeled quartz-rich zones as continuous undulating, tabular bodies. Gold is restricted to one or more generations of quartz veins. Grade data on which the model is based are obtained from reverse circulation, rotary air blast and diamond drilling. The different generations of QV1, QV2 and QV3 quartz veins can not be reliably distinguished in the drill samples. The widths of the tabular bodies (6 m-12 m) are considered to represent open-pit mineable units. Wireframe modeling of the lodes has been completed with hard boundaries between mineralized and barren rock. The wireframes were completed by using plans and sections to trace 2D polylines which were then stitched together to form 3D solid body models. Solids models were completed for all six zones with surfaces created for topography and the oxide/fresh interface. The geological solids were used to create the rock type model and filter samples for statistical analysis. The rock type model was used to control the grade interpolation. The geological solid were also used to create a partial or percent model which allowed partial block models to be estimated, improving volume estimates. A block size of 25 m x 12.5 m x 6 m was used. The deposit has been drilled at 25 m x 25 m to 25 m x 50 m spacing, although not completely filled in. SRK has expressed the concern that the tabular lodes are small and while selective mining may be possible, dilution may be greater than anticipated. The SGL grade control data would suggest that there are no hard chemical edges to the mineralized bodies.



AMCL considers that, in fact, careful visual and chemical grade control may permit selective mining producing fewer tonnes at a higher grade. In higher grade areas such as B-Shoot and the 242 zone, in pit mapping should be done to determine whether higher-grade zones can be discriminated. Within the wireframe models the grade samples were composited to 3 m within the high-grade quartz-rich bodies. Only samples that occur within the wireframes have been used for grade interpolation. The weathered/fresh rock contact is relatively sharp with the saprolite hosted oxide separated within the wireframes. Univariate statistics were compiled for each mineralized zone by drill hole type. The relative standard deviation of the date is quite different for each of the drilling method populations. Outlier values which affect the data have been identified but not removed. Several different types of variography were conducted on the exploration data. Raw variograms did not exhibit any grade continuity within the individual tabular bodies. Pairwise-relative variograms and variograms of the log of the gold grade were also modeled. These showed significantly improved structures, however, SRK rejected the use of the pairwise variography as not representative of a true spatial covariance. The exploration data are highly skewed and the variograms of the logarithms of the gold grade were rejected on this basis. SRK opted to apply a gaussian transformation to the gold grades to derive a gaussian equivalent value. The relationship between the untransformed grades and the gaussian equivalent values were modeled using a series of orthogonal polynomials which allowed variograms of the gaussian equivalent values to be back-transformed into real sample space. The data was then re-modeled to develop a variogram that best fit the variance of the data in untransformed space. These variograms were subject to ordinary kriging estimation processes to derive sample weighting and block estimates for material that intersected the wireframe models established for each domain. The sample search area was tested for each domain on several representative blocks. Variograms for this deposit have high relative nugget effects and short ranges relative to the dimensions of the mineralized bodies. In SRK's opinion the gold distribution is partly random and partly spatially controlled by the distribution and orientation of quartz veins. The high nugget effect is related to the coarse particulate nature of the gold, structural disturbance and to sampling issues. Blocks with two or more drill hole intercepts, generally, indicate high confidence estimates but these blocks are discrete and rarely connected. Proximity to data (<15 m) does not guarantee high- quality block estimates. Simple kriging was used as a check on the use of the mean grade in the resource estimation process. This suggested a strong reliance on the mean grade with 50 % of the weighting applied to the mean. Block estimates derived from the kriging indicate that the block estimates are not precise and should not be relied on for detailed mine planning.



Two searches were conducted on each block. In the first search the size included as much local data as possible whilst in the second the search tried to generate grade values for all the unestimated blocks using an extended search range. In terms of resource estimation SRK has concluded that sampling problems, high relative nugget effects and generally short projection ranges preclude any of the resource estimate being classified as a measured resource. The wireframe models prepared by Wassa geologists do not represent a unique interpretation of the geology. Other interpretations may honor the data. The distinction between indicated and inferred resources was based on cross-sections where the point was identified that drill hole data could not confirm the continuity of the mineralized zone. A wireframe was created for each domain which enclosed the volume considered to be an indicated resource. SRK determined that blocks that intersected the wireframe with >90 % of the block volume falling inside the wireframe were considered to be in the indicated resource category. Golden Star geologists determined that blocks that intersected the wireframe with >50 % of the block volume falling inside the wireframe should be considered to be in the indicated resource category. Golden Star would place a greater proportion of the resource in the indicated category. In AMCL's opinion the SRK definition is conservative. SRK suggests that the correlation between block grades and drill hole grades within the blocks are low which means that a direct comparison of block grades with the drill hole grades is not valid. Given the data coverage, some level of smoothing is to be anticipated within the estimation with the theoretical block variance being greater than the variance the estimated block grades. In all of the mining zones except Mid-east this is the case. This results in the Mid-east resource estimate being classified in the inferred category. Uniform relative densities of 2.7 t/m3 for all fresh (sulphide) domains and 1.8 t/m3 for all oxide domains were used. In AMCL's opinion these are reasonable. 9.2.1 Comments on Resource Estimation Procedures AMCL has reviewed the basis of the SRK resource estimation. The classification and reporting of resources and reserves from coarse gold-bearing veins is a difficult task. Both geological and grade continuity must be clearly established. The use of a gaussian transform does not appear to be fully warranted. While the data does have a reasonable spread which is mainly attributable to the high-grade outliers, AMCL was able to produce variograms for some of the deposits at Wassa with only moderate nugget effects and reasonable (albeit short) ranges. AMCL used Isaaks & Co. Sage® 2001 variogram modelling software to create variograms for the South East mineralized zone. The figures on the following pages are the direct output from the software.



Figure 9.0: Isaaks Output



Figure 9.1: Isaaks Output



The variograms created in Sage® were exported in a Gemcom® compatible format as follows:

Figure 9.2: South East Zone Down Hole Variogram A large body of literature suggests that inverse distance weighting methods (often to the power cubed)using elliptical search areas where the long axis is parallel to the ore shoots are appropriate methods for interpolating grades for this type of deposit. The visual validation of the model is stated by SRK to be invalid. This raises some concerns as kriging is a distance weighting method of grade estimation so if the visual comparison of drill hole grades and estimated block grades is invalid then either the model is over smoothed or the model is not selective enough compared to the input values. This situation is directly attributable to the large block size relative to the drill hole spacing. This may have a direct impact on the estimation of grade and is the reason for AMCL's conclusion that the resource estimation should be considered as a "global estimate". The degree of over smoothing and relatively large block size would suggest that the model may not be appropriate for production planning.



No grade cutting or capping has been carried out within the model. In AMCL's opinion this is appropriate for this type of deposit as the local, high-grades represent an important part of the grade distribution. Additional drilling may allow some inferred resources to be reclassified in the indicated category but it is unlikely that much of the Wassa resource can ever be classified as measured. We would stress that this does not mean that the gold is not present, merely randomly distributed. 9.2.2 Reconciliation of the SRK Resource Estimate With Golden Star's Most Recent Published

Resource Estimate Golden Star Resources Ltd.'s most recent published resource estimate for Wassa was released in a United States Form 10-K filing dated March 31, 2003 and in a press-release (PR03-11) dated April 30, 2003. This resource estimate was based on the work completed by SRK and reviewed in this report. Table 9.3: 10-K Published Resource Estimate

Measured Resources

Indicated Resources

Measured + Indicated Resources

Contained Gold

Inferred Resources

Material

Tonnes (‘000)

Grade (g/t)

Tonnes (‘000)

Grade (g/t)

Tonnes (‘000)

Grade (g/t) (‘000 oz) Tonnes

(‘000) Grade (g/t)

Oxide - - 3,305 1.13 3,305 1.13 120 3,319 1.07

Fresh - - 9,287 1.64 9,287 1.64 489 25,523 1.16

Heap Leach Pads - - 5,177 0.75 5,177 0.75 126 - -

Total - - 17,770 1.29 17,770 1.29 735 28,843 1.15

90% Attribution - - 15,993 1.29 15,993 1.29 661 25,958 1.15

This resource estimate is based on a different definition of "indicated resources" than used in the SRK report (see discussion on page 32). Based on a visual inspection of cross-sections where drill hole data could not be used to establish the continuity of the mineralized zone, a wireframe was created for each domain which was considered to be in the indicated resource category. Blocks which intersect the wireframe model with >50 % of the volume of the block falling within the wireframe are classified in the indicated category. Both SRK and AMCL concur that the classification of resources is subjective and have no objection to the classification used and reported by Golden Star. Cut-off grades were 0.4 g/t and 0.6 g/t, respectively for oxide and fresh material determined using processing costs of US$5.66/t-US$6.40/t, process recoveries of 92 %-93 % and overall mining recoveries of 95 % at a gold price of US$325/oz. These are similar to the current study. AMCL notes that it would report very slightly different numbers for the contained ounces of gold. This is most likely due to rounding differences within the various calculations.



The SRK oxide resource statement provides an estimate of measured and indicated resources (mineralized material of American usage) of 3,486 Mt @ 1.16 g/t for a contained gold content of 130,437 oz.. The Golden Star published resource statement reports an estimate of:

3,305 Mt @ 1.13 g/t for a contained gold content of 120,000 oz.. This is a marginal (non-material) reduction in tonnage (5.3 %) and grade (2.6 %) from that reported by SRK. The SRK fresh (non-refractory sulphide) resource statement provides an estimate of measured and indicated resources (mineralized material of American usage) of 7,674 Mt @ 1.71 g/t for a contained gold content of 420,797 oz. The Golden Star published resource statement reports an estimate of:

9,287 Mt @ 1.64 g/t for a contained gold content of 489,000 ounces. Golden Star has reported an approximately 17.3 % greater tonnage at a 4.2 % lower grade for an overall 14 % increase in overall gold content. It must be stressed that this is only a reclassification of some additional material in the indicated category. The resources in the inferred category have been concomitantly decreased. The total oxide and fresh mineral resource in the indicated category is:

12,592,595 tonnes @ 1.50 g/t for a contained gold content of 609,086 ounces. This corresponds to the "mineralized material" of American usage. The total oxide and fresh mineral resource in the inferred category is:

28,842,514 tonnes @ 1.15 g/t A substantial low-grade resource has been identified in the heap leach pads and reported in the 10- K filing. A resource classified in the indicated category is estimated at 5.177 Mt @ 0.75 g/t for a contained gold content of 126,000 oz. AMCL has briefly reviewed the drilling and sampling procedures and are satisfied that this resource does exist. The above section has been extracted from the Qualifying Report No. 03PM67, dated April 2003, prepared by consultants AMCL. This report can be viewed on Sedar (www.sedar.com) under the Golden Star Resources directory.



GSR geologists have carried out additional work on the resource wireframe model as additional data has become available, and have also remodeled the Mid-East zone, resulting in the following resource estimation:

Category COG Tonnes Grade g/t

Ounces

Indicated Oxide 0.50 4,549,615 1.13 164,763 Indicated Fresh 0.60 11,542,802 1.56 579,467 Indicated Total 16,092,417 1.44 744,230 Inferred Oxide 0.50 2,114,965 1.02 69,664 Inferred Fresh 0.60 23,267,677 1.15 862,370 Inferred Total 25,382,642 1.14 932,034

This results in the reclassification of approximately 3.5 million tonnes of material from the Inferred Resource category to the Indicated Category. SRK have concurred that the above approach is equally valid, and have indicated that the above classifications can be accepted. In addition to the above resources, a substantial resource exists within the heap leach pads processed by the previous operator. This resource, which can be classified in the Indicated Resource category, is shown below:

Category COG Tonnes Grade g/t

Ounces

Indicated HL 0.40 5,177,105 0.75 124,836

The above resources were estimated by Mr. S. Mitchell Wasel, an employee of GSR. Mr. Wasel is a qualified geologist with over 15 years experience, and has been working as Exploration Manager at Bogoso Gold Limited (a 90% owned subsidiary of GSR) since 1999. He is a Member of the Australasian Institute of Mining and Metallurgy. Caution: The terms Measured Resources and Indicated Resources conform to the requirements of the Canadian regulatory authorities, but are not recognized by the US Securities and Exchange Commission. For such usage, they should be referred to as ‘Mineralized Material.’

9.3 Mineral Reserve Estimate

Pit optimizations were carried out using the NPVScheduler® suite of programs from Earthworks Corporation. The grade resource model used in the optimizations was that prepared by SRK in March 2003, but making use of the revised resource classification categories as prepared by GSR (see Section 9.2 above). In the optimization process, only the Indicated Resource category of mineralized material was used in the evaluation process, with all Inferred Resource material being treated as waste rock.



9.3.1 Pit Optimization Parameters The parameters used in the pit optimization process are shown in the table below. All costs are expressed in US dollars.

Oxides Fresh Waste Mining ($/tonne) US$0.75 US$0.98 Ore Mining ($/tonne) US$1.10 US$1.19 Waste Haulage Distance 1.0 km Extra Ore Haulage Distance 0.3 km Incremental Haulage cost US$0.12 per tonne per km Incremental Vertical cost US$0.012 per 6m bench Rehabilitation cost US$0.05 per tonne fresh waste US$0.05 per tonne oxide waste Dilution 6% at zero grade Mine Recovery Factor 98% Oxides Fresh Process Costs US$3.56 US$4.29 Process Recoveries 93% 92% G & A Costs US$3.262 million per annum Slope Angles Footwall Hanging Wall Oxide Fresh Oxide Fresh All Pits 55° 55° 55° 55° Gold Price US$300/oz Royalty 3% Discount Rate 10%



9.3.2 Optimization Cost Parameters The mining cost parameters used in the optimization have been derived from a combination of costs currently and historically achieved at the GSR’s ongoing mining operations at Bogoso and Prestea, and from first principals, modified where appropriate to take account of the specifics of the Wassa operations. The costs were divided up into Fixed Costs, which are relatively independent on tonnage moved, and Variable Costs, which vary in proportion to the total tonnage moved. Further division has been made into oxide mining costs, and fresh mining costs, sub-divided into ore mining costs and waste mining costs. These costs were subsequently applied to an indicative mining schedule, to derive weight average costs suitable for optimization purposes. Incremental costs were allowed for increases in pit depth below a reference level, to account for increases in cycle times with depth, and hence increases in costs. Process unit costs were obtained from the metallurgical testwork program carried out on the oxide and fresh material by LRA and by Ammtec, as detailed in Section 9. Recoveries were similarly based on the results of the testwork program. General and Administration costs comprise all those costs that do not fall within the mining or operating ambit, and would generally be fixed costs, being relatively independent on tonnages mined or processed, unless there were to be a significant change in production levels. 9.3.3 Geotechnics Geotechnical investigations had been carried out by the previous operator prior to the start of mining, using Golder Associates as specialist consultants. Upon acquiring the project, GSR commissioned SRK to conduct a review of this previous work, and to propose new pit slope angles. SRK spent 7 days on site, and were able to utilize the previous Golder Associates data, and were also able to take direct measurements from exposed faces in the open pits, as well as from fresh core samples from the ongoing diamond drilling program. A geotechnical report No. U2145 was issued in March 2003. A summary of the SRK recommendations is shown below:

Pits Wall Material type Slope Angle

Oxide 55° 419 SE B Shoot F Shoot

All Walls Fresh See Chart below

Oxide 50° North West Fresh 50° Oxide 55°

Main 2 North 242 South East Fresh See Chart below

Oxide 55° Starter Pit All Walls Fresh See Chart below



Oxide (slopes up to 50° high) 55° Oxide (slopes more than 50° high) 50°

Mid East Mid West Deadman’s Hill

All Walls Fresh See Chart below

SlopeHeight/Slope Angle relationship for Fresh Rock Units

20

30

40

50

60

70

80

90

100

110

120

130

140

150

160

50 51 52 53 54 55 56 57 58 59 60 61 62 63 64 65

Overall Slope angle in Fresh Rock (°)

For pit optimization purposes, it was deemed appropriate to utilize a common angle of 55° for all slopes, in oxides and in fresh, on the understanding that after inclusion of ramps in the pit design phase, the inter-ramp angles in the fresh material could be considerably steepened up. 9.3.4 Dilution A dilution of 6% was applied to all ore blocks, to account for waste material mined with the ore and sent to the process plant. This factor of 6% was derived from analysis of the grade control data collected during the previous phase of operations, reconciled against process plant feed. The dilution tonnage was ascribed a zero grade, although in practice, there would likely be some grade attached to waste material flanking the ore zones. An ore recovery factor of 98% was also applied, to cater for 2% of ore grade material that may be sent through to the waste rock dumps.



9.3.5 Gold Price A gold price of $300 per ounce was used in the optimizing process. This was tempered by a government royalty of 3% of revenue.

9.4 Evaluation of Optimum Pit Shell.

The optimum pit shell from the pit optimization process was evaluated against the SRK grade model, and returned the following tonnages and grades.

$300 Rock Type Category Tonnes X 106

Grade g/t

Indicated 3.73 1.21 Inferred 0.14 1.24 Oxide Mineralized

Material TOTAL 3.87 1.21 Indicated 7.27 1.76 Inferred 0.23 1.51 Fresh Mineralized

Material TOTAL 7.50 1.75

Total Ore 11.37 1.57 Oxide Waste 11.62 Fresh Waste 17.01 Total Waste 28.63 Total Material 40.00 Strip Ratio 2.52

This is all the material contained within the pit envelope, and should not be considered as Reserves, as it does not incorporate pit design works, nor factors such as dilution and ore loss. It will be noted that 0.37 million tonnes of mineralized material within the pit is from the Inferred Resource category, representing approximately 3% of the total mineralized material.

9.5 Mine Design

The optimum pit envelope obtained as above was subsequently engineered to produce a practical, workable pit. This involved the inserting of access ramps, the steepening of inter-ramp angles to the geotechnical design recommendations, the installation of appropriate safety berms etc. Any synergies between the various parts of the deposit were investigated, such as the sharing of access ramps between adjacent pits, thereby improving upon waste stripping requirements. In addition, several small oxide



pits as defined by the pit optimization process were excluded from further consideration, due to their small size, or for their possible interference with existing infrastructure around the mine site. Following the mine design process, the in-pit resource tonnages and grades can be shown in the table below. Cut off grades used to determine ore tonnages were 0.53 g/t and 0.63 g/t for oxide material and fresh material respectively.

$300 Designed Pit – In-Pit Resources Rock Type Units Main

Pit DMH Mid

East South East

419 Total

Oxide Grade tonnes 106 0.62 1.01 0.33 0.86 0.46 3.28 Oxide Grade g/t 1.36 1.19 1.42 1.07 1.15 1.20 Fresh Ore tonnes 106 4.38 0.23 0.07 0.97 1.07 6.71 Fresh Grade g/t 1.92 1.15 1.32 1.30 1.66 1.75 Total Ore tonnes 106 5.00 1.24 0.40 1.83 1.53 10.00 Total Grade g/t 1.85 1.18 1.40 1.19 1.51 1.57 Waste tonnes 106 15.67 1.93 0.92 4.05 5.35 27.92 Total tonnes 106 20.67 3.17 1.31 5.88 6.88 37.92 Strip Ratio 3.13 1.56 2.30 2.22 3.50 2.7

It will be noted from the above that there has been a 12% reduction in ore tonnage between the optimum pit envelope (see section 9.3) and the sum of the individual designed pits. This is an ore loss that has been occasioned by the remnants of ore left in the bottom of the pits, as minimum mining widths are reached, and also the pockets of oxide ore in small optimized pit envelopes that have not been included as designed pits. There has also been a small increase in waste stripping, which, together with the reduction in ore, has led to an overall increase in stripping ratio (waste : ore) from 2.5 to 2.7. This is due principally to the insertion of ramps, which have resulted in a flattening back of some of the slopes as the ramps spiral to the pit bottom. During this next phase of the project, the designed pits will be critically analyzed to identify where refinements can be made to the ramp systems, which will have the effect of reducing the overall slope angles and hence reduce waste stripping requirements. This will also incorporate a possible change to 12% gradient ramps should it be deemed appropriate. The above tonnages are do not include the effects of dilution and ore loss, and should not be treated as Reserves.



9.6 Mineral Reserves

During the pit optimization process, a dilution factor of 6% at zero grade was applied to all mineralized blocks. This has the effect during the optimization process of reducing the economic value of the block, and hence has an impact on final pit size. In this conversion of in-pit Resources to Mineral Reserves, a dilution factor of 5% has been applied, at a grade of 0.25 g/t. During the previous mining operation, the primary excavator used was an O&K RH 120 tonne machine, using a 13 m3 bucket with a width of 3.7m. GSR plan to use a smaller excavator with a 6m3 bucket, with a bucket width of 2.1m. It is believed that this smaller bucket size will have a significant impact on the degree of selectivity on mining the ore blocks, and as a result, the dilution factor has been subjectively reduced to 5%. The dilution grade of 0.25 g/t is based on that experienced by the previous operator. The same mine recovery factor of 98% was applied to the in-pit resource tonnages to derive the Mineral Reserve estimate inside the design pits.

$300 Pit – Mineral Reserves Rock Type Units Main

Pit DMH Mid

East South East

419 Total

Oxide Grade tonnes 106 0.64 1.04 0.34 0.88 0.47 3.38 Oxide Grade g/t 1.31 1.15 1.36 1.03 1.11 1.16 Fresh Ore tonnes 106 4.51 0.23 0.07 0.99 1.10 6.91 Fresh Grade g/t 1.84 1.11 1.27 1.25 1.59 1.69 Total Ore tonnes 106 5.15 1.27 0.41 1.88 1.57 10.29 Total Grade g/t 1.77 1.14 1.35 1.15 1.45 1.51 Total Waste tonnes 106 15.52 1.89 0.90 4.00 5.31 27.63 Total Mined tonnes 106 20.67 3.17 1.31 5.88 6.88 37.92

Strip Ratio 3.02 1.48 2.21 2.13 3.38 2.69

The categorization of Proven and Probable Reserves in this study has used the generally accepted guidelines defined within the Canadian NI 43-101. Proven Reserves are those in pit mineralized materials for which there is the same level of geological understanding as the Measured Resources, and which, as a result of this feasibility study, can be



deemed to be able to be economically mined and processed, under the defined economic conditions. There are no Proven Reserves within the pit designs for the Wassa Mine. None of the Resources are yet known with sufficient demonstrated confidence to be able to categorize them as Measured Resources, and hence there can be no classification into Proven Reserves. Probable Reserves are those materials for which the level of geological understanding is the same as the category of Indicated Resources, but which can also be deemed to be able to be economically mined and processed. While it is possible to categories Measured Resources as Probable Reserves, due, for example, to uncertainties surrounding recoveries, it is not possible to categories Indicated Resources as Proven Reserves. This Mineral Reserve Statement has made the assumption that all Indicated Resources are sufficiently understood in terms of mining and processing that they can be directly classified as Probable Reserves. The pit optimizations were based on the Indicated Resources only, with Inferred Resources being treated as waste. However, within the designed pits, there is a small proportion of inferred category material, as mentioned in section 9.4. This makes up approximately 3% of the total resource, and is not considered significant. The decision has been taken to incorporate this inferred tonnage as part of the Probable Reserve tonnage. These Mineral Reserves are included within the Mineral Resource estimate as summarized in Section 9.2 The above reserves were estimated by Mr. David Alexander. Mr. Alexander is a qualified mining engineer with over 20 years experience in the mining industry, and has been working as Projects Planning Manager at Bogoso Gold Limited (a 90% owned subsidiary of GSR) since 2000. He is a Member of the Institution of Mining and Metallurgy, and is a Chartered Engineer under the auspices of the Engineering Council of UK.

10 Mining Operations

10.1 Mining Operations

The Wassa deposit was developed by the previous operator as an open pit mining operation, and this continues to be the most practical method of exploiting the resource. The style of mineralisation and the low grade of the ore precludes any form of underground development, or any form of hydraulic mining. The previous operator mined the deposits for an annual average production of 3 million tonnes per annum. Their mining fleet comprised a 260 tonne primary excavator, and four 90 tonne capacity haul trucks. This was a relatively inflexible mining fleet, in that blending from the pits was difficult to achieve, due to the difficulties in moving such size of excavator from section to section. GSR has therefore planned on using three 110 tonne excavators, loading into nine 90 tonne capacity haul trucks. By so doing, GSR will be able to schedule the feed to the process plant from various pits concurrently, in order to achieve a consistent blend of oxide and fresh material, while still maintaining



waste stripping. The large fleet is necessary due to an increased mining rate, and the increased depth of the open pits. It is recognized that the excavators will be markedly undertrucked, and discussions are ongoing with equipment suppliers regarding more suitable matching of excavators with trucks. The reclamation of the Phase 2 heap leach material during 2004 will be done using two Caterpillar 988G front end loaders. Mining operations, which are scheduled to commence in December 2004, will be carried out by conventional backhoe excavator and truck methods, on an ‘owner operator’ basis. It is planned to use three Liebherr 984C excavators, loading into a fleet (at maximum size) of nine Caterpillar 777D 90 tonne capacity haul trucks. It should be noted that these equipment types and specifications are for estimation purposes only, and that no commitment has been made to any particular supplier. Productivities have been derived from first principles, and from work studies performed on the Liebherr 984 excavators operating at GSR’s Bogoso/Prestea operation. Cycle times for the haul trucks have been estimated using the Caterpillar FPC (Fleet Productivity and Costing) package. Drilling will be carried out using Tamrock Pantera 1500 drill rigs, using 102mm diameter blast holes over six meter benches. Powder factors for oxides are based on those currently prevailing at Bogoso/Prestea, while for the fresh rock, the powder factor has been based on the experience gained at the nearby Abosso mine. Mining will be carried out on a 24 hour per day, 7 day per week basis, with operators working to a four shift rotating roster. Maintenance work on the mining equipment will be carried out by an in-house maintenance department, which will utilise the heavy equipment yard as established by the former mining contractors. This comprises four fully equipped maintenance bays, and is adjacent to a fuel farm. Major repairs and component rebuilds will be carried out at the specialist workshops located at the Bogoso minesite.

10.2 Production Schedule

Production through the CIL plant is expected to start in 1st Quarter 2004. The initial feed to the mills will be material reclaimed from the Phase 2 heap leach pad, prepared by the previous operator. A reserve of 4.4 million tonnes has been estimated for this area, at a grade of 0.75 g/t. While this grade is low, operating costs will be low, due to the fact that the material has already been mined, and has been crushed to –6mm in size. Due to the fact that the leach pad material has already been crushed, it is expected that 4.0 million tonnes of this can be processed in the first year of operations (2004), with the balance in the first months of 2005. At this point, it will become necessary for the feed to be supplemented by ore from the open pits. A nominal blend of 66% fresh to 34% oxide has been allowed for in the mining schedule from the open pits. Any significant increase in fresh material percentage would result in reductions to the milling rate, to a maximum of 2.9 million tonnes per annum for 100% fresh ore



It is recognized that it may not always be possible to achieve this blend direct from the open pits, and so stockpiles will be maintained on the crusher pad to ensure a consistent flow. Approximately 40% of ore is expected to be direct tipped into the crusher rock box, and 60% to be reclaimed from the stockpiles. Operations at Wassa ceased while there was still ore exposed in the pit bottoms. Work can therefore resume there with minimal pre-production work. An amount of 1 million tonnes of oxide pre-stripping has been allowed for in late 2004, to ensure sufficient longer term exposures. A summary of the production schedule is shown on the following page:

Mining

Units 2004 2005 2006 2007 Totals

Oxide Ore tonnes x 103 1380 1210 790 3380 Oxide Grade g/t 1.13 1.19 1.12 1.15 Oxide Waste tonnes x 103 1000 4140 2250 3110 10500 Fresh Ore tonnes x 103 2120 2290 2500 6910 Fresh Grade g/t 1.56 1.85 1.63 1.68 Fresh Waste tonnes x 103 6480 4540 6110 17130 Total Ore tonnes x 103 3500 3500 3290 10290 Total Grade g/t 1.39 1.62 1.51 1.51 Total Waste tonnes x 103 1000 14120 10290 12510 37920 Strip Ratio 3.0 1.9 2.8 2.7 Processing Oxide Ore tonnes x 103 1380 1210 790 3380 Oxide Grade g/t 1.13 1.18 1.15 1.15 Recovery % 93% 93% 93% 93% Fresh Ore tonnes x 103 2120 2290 2500 6910 Fresh Grade g/t 1.56 1.85 1.63 1.68 Recovery % 92% 92% 92% 92% Leach Pad Ore tonnes x 103 4000 400 4400 Leach Pad Grade g/t 0.70 0.70 0.70 Recovery % 86% 86% 86% Total Ore tonnes x 103 4000 3900 3500 3290 14690 Total Grade g/t 0.70 1.32 1.62 1.51 1.27 Recovery % 86% 92.0% 92.3% 92.2% 91.2% Gold Produced ounces 74,945 150,001 166,946 149,924 541,815



11 Mineral Processing and Metallurgical Testwork

11.1 Metallurgical Testwork

Extensive metallurgical testwork was carried out on the Wassa ores, in order to fully ascertain their processability through a conventional Carbon-in-Leach process route. Sample selection and preparation was carried by GSR geologists, while the testwork was conducted by Lakefield Research Africa (“LRA”) of Johannesburg, South Africa, and Ammtec Metallurgical Laboratories (“Ammtec”) of Perth Australia.

11.1.1 Sample Selection

The metallurgical samples were primarily selected from diamond drill (DD) core and from reverse circulation drill (RC) chippings, collected during the exploration drilling programs of 2002 and early 2003. During this program, a total of 20,100m were drilled over 197 holes, comprising 12,200m of RC chippings from 101 holes, and 7,900m of DD core from 96 holes. A small proportion of the samples came from available core from previous exploration drilling programs. Diamond drilling was generally carried out to NQ diameter (47 mm), and totaled 7,900 m over 96 holes. All core was logged in detail, taking into account such features as recovery percentages, degree of weathering, lithologies, alteration, and vein densities. Structural features were also described and recorded for geotechnical purposes. The core was photographed, prior to being split using diamond saw. Half of the core was taken for Au assay purposes, while the balance was kept in sequentially marked wooden core boxes. Those cores selected for metallurgical testwork purposes were subsequently quartered, before being washed, dried and weighed. The remaining quarter core was left in the core box as permanent reference. Reverse circulation samples were collected using a face sampling bit, and collected at 1m intervals. For assay purposes, the 1m samples were passed through a 3-stage splitter, with each split being bagged and labeled separately. The bags are stored in sequential order, enabling easier access for selection. The selection of the appropriate samples for metallurgical testwork was done with reference to an indicative optimum pit shell at a gold price of $300/oz. By using this shell, the geologists were able to target those drill intersections that fell within the pits, thereby increasing the representivity of the testwork program. In total, seven areas were identified as being of value to treat as separate areas, being dependent on such aspects as geographic separation, or through significant changes in geological or structural controls. These seven areas were subsequently sub-divided by elevations, into three broad horizons, representing increasing depth from surface. These elevations were 1042m – 976m RL, 976m – 958m RL, and 958m – 880m RL (general ground level is 1006m RL). In total therefore, 21 variability samples were collected for fresh material, while 6 variability samples were collected for oxides (mining of oxides by the previous operator had depleted several of the seven identified areas of any significant oxide resource). The minimum weight required for testwork on each variability sample was 32kg, and



included one to two meters of low grade material from the flanks of the orebody to provide a representation of mining dilution. In addition to these variability samples, sufficient excess material was selected from each of the seven identified areas, to enable representative bulk composite samples to be prepared in the testwork laboratories. These samples, were combined in proportion to the percentage that the particular area contributed to the overall plant feed for each of the oxide and fresh ore respectively, and the composites would be used to determine the plant process flow, overall recoveries, appropriate reagent suites, and operating costs. The minimum weight required for the bulk sample testwork program was 90kg. The leach pad material stacked and processed by the previous operators still represents a significant resource, and an intense deep augering program was carried out on the pads to evaluate them. The individual auger samples from each hole were composited into a single sample using an electric mixer, before being split through the three way splitter. For bulk compositing purposes, the leach pads were divided into Phase I and phase II regions, with each phase sub-divided into strips of approximate equally tonnage. A total of 8 such composite samples were taken, at 55kg each. These were subsequently combined to provide two representative bulk samples for the leach pad material, one each for Phase 1 and for Phase 2. In total, the following samples were dispatched by air freight to LRA:

Fresh material 1,080 kg Oxide material 390 kg Leach pad material 440 kg

The total quantity of material sent to South Africa for the testwork program, was in excess of requirements, and there provides flexibility should additional test be required over and above those initially planned.

11.1.2 Metallurgical Testwork Results

The metallurgical samples were delivered to LRA in Johannesburg for a detailed metallurgical study of both bulk blend samples and individual (pit by pit) variability samples. From the variability samples and bulk composite samples, LRA performed the task of compiling them into representative samples for testwork. This was done to provide a degree of independence in sample selection. For each variability sample, a weight of 32kg was subsplit from the sample provided. For the bulk samples required for testwork, namely Oxide, Fresh, Heap Leach 1 and Heap Leach 2, each were composited by LRA personnel using the resource data to achieve the correct proportionality of each sample within the overall bulk sample. These bulk composites were used to determine overall plant process flow, overall recoveries and operating costs and to define the optimum operational parameters. A total of 102 kg was split out per bulk sample.



The test program highlighted a number of ‘spotty gold’ difficulties with the Wassa ore in terms of severe inhomogeneity of fire assay grade, even for example across spacing as small as an inch apart, between two neighboring quarters of a piece of 48 mm diameter diamond drill core. The LRA assay procedures appeared particularly susceptible to the spotty effect making repeatability very poor and a concerning level of non-random bias was evident in the kinetic and variability phases of the program. It was decided to use LRA’s results from the bulk sample testing to determine the optimum process route, reagent consumptions and leach parameters. LRA’s testwork was also used to produce a best indication of expected recoveries, but it was considered that the consistent lack of reconciliation from the test accounting and poor repeatability meant that a bankable level of confidence could not be derived for gold recoveries. Consequently, variability sub-samples were spilt out and shipped to Ammtec. The Ammtec procedure uses separate tails assays on five different size fractions to determine the overall tails value and this procedure mitigates against the spotty gold assay effect and produced bankable quality reconciliation. The various leach kinetics curves were not impaired by widely scattered data points, as had been the case with the LRA’s work. The LRA bulk sample data described a fresh ore of medium hardness and abrasivity (14.8 kWh/t BWI and 0.29 abrasion index). Oxides are typical for the area and are generally too soft for impact hardness or bond work index testing. Bulk samples have been subjected to the “La Plante Gravity Recoverable Gold” (GRG) test using 40 kg of sample in sequential grind vs. recovery tests through a 7” Knelson centrifugal concentrator. First pass recoveries in excess of 60% have been indicated for both fresh and oxide samples. Gold recovery is predominantly in the -75µm to +38µm size fraction, although gold deportment occurs across a wide size range. Leach performance as predicted by LRA is practically insensitive to grind size. Coarser grinds tend to produce greater tails variability as a result of occasional unliberated gold and / or coarse gold particles that have excessive leach times due to mass transfer restrictions. Sensitivity to cyanide concentration is also greatly increased for coarser fractions (<65% -75 µm). A final grind of 75% passing 75 µm produces consistent tails values and far greater insensitivity to variations in reagent concentration and this grind size would be ideal for a CIL circuit that had no upstream gravity recovery. CIL results of gravity tails material indicates that equivalent or better final tails values results can be obtained at a much coarser grind size of 55% passing 75 µm. Given the good GRG results, there is a clear benefit from including a gravity circuit. Extremely fast kinetics of up to 75% dissolution within the first 10 minutes, indicates strong compatibility with the proposed pipe reactor flowsheet. There is a benefit from adding oxygen and the first few hours of leach show improved kinetics over air-sparged tests. Mild preg-robbing tendencies were observed, particularly in the oxide ores, and this dictated a CIL rather than a CIP flowsheet. Optimum leach times were difficult to determine from the various LRA’s kinetics curves which at times appeared to conflict with each other. A leach time of 20–24 hours appeared the most suitable



compromise of the various data sequences. Oxide bulk sample testing indicated a 93% recovery and fresh ore indicated 92% recovery potential. The Ammtec variability results showed greater reproducibility and the predicted whole ore oxide ore recovery is 95.5% while the predicted fresh ore recovery is 92.3%. Optimum leach time was confirmed at 24 hours. The average of the 7 fresh ore variability samples (pit by pit) should provide very similar results to the average of the 3 RL variability samples, since each of these represents two different ways of categorizing the entire fresh ore resource. The LRA work provided no correlation between the 2 sets of data which was another evidence of poor repeatability whereas the Ammtec work summarised in the table below shows excellent correlation on head and tails grades, recoveries as well as cyanide consumptions. Only the lime consumption of a single test proved anomalous.

Sample CN

Added (kg/t)

CN Used (kg/t)

Lime Used (kg/t)

Calc. Head (g/t)

Tail

(g/t)

Recovery

24 hr (%)

Recovery

48 hr (%)

Resource

(%) Fresh Ore

B Shoot 0.35 0.09 0.78 1.69 0.089 94.62 94.75 27.12 F Shoot 0.35 0.09 0.83 2.6 0.157 92.53 93.96 16.37 M2N Starter 1 0.35 0.05 0.84 1.69 0.066 95.22 96.13 18.9 M2N Starter 2 0.35 0.04 3.86 1.33 0.096 90.39 92.86 18.9 Mid East 0.35 0.24 0.96 2.99 0.333 88.77 88.88 0.9 Dead Man’s Hill 0.35 0.07 0.95 1.18 0.200 83.19 83.09 5.6 South East 0.35 0.04 0.97 1.64 0.143 90.53 91.28 13.8 Weighted average by pit 0.35 0.07 1.39 1.75 0.110 92.35 93.33 100.0 RL 958 - 880 (deep) 0.35 0.07 0.67 2.35 0.107 94.13 95.47 - RL 958 - 976 (mid) 0.35 0.10 1.34 1.93 0.119 93.47 93.83 - RL 1042 - 976 (shallow) 0.35 0.07 0.76 1.04 0.085 89.42 91.81 - Average by RL 0.35 0.08 0.92 1.77 0.100 92.34 93.70 -

Oxide Ore F Shoot 0.35 0.07 1.92 1.09 0.022 97.84 97.95 13.7 M2N Starter 0.35 0.11 1.07 0.55 0.026 92.76 95.24 13.3 Dead Man’s Hill 1 0.35 0.16 2.48 1.11 0.042 96.14 96.25 16.0 Dead Man’s Hill 2 0.35 0.13 2.62 1.64 0.079 92.76 95.25 15.9 South East 1 0.35 0.17 2.54 1.24 0.012 98.61 99.03 15.0 South East 2 0.35 0.11 2.3 1.49 0.028 95.57 98.15 15.0 Weighted average by pit 0.35 0.13 2.19 1.20 0.040 95.50 96.86 100.0

Heap Leach 2 P2Z1 0.35 0.05 1.32 0.56 0.045 89.69 92.10 - P2Z2 0.35 0.08 1.24 0.53 0.044 89.46 91.86 -



P2Z3 0.35 0.05 0.97 0.49 0.033 93.6 93.31 - P2Z4 0.35 0.07 0.98 0.71 0.035 94.47 95.07 - Average 0.35 0.06 1.13 0.57 0.040 91.81 93.09 - 30 Oxide:70 Fresh Blend 0.35 0.12 1.77 1.82 0.081 95.29 95.58 -

From the above Ammtec results, for fresh ore variabilities, pit by pit averaging gives a recovery of 92.35% while averaging by elevation gives 92.34% recovery. By comparison, the LRA data showed a wide variation in weighted averages for fresh ore pits from the last variabilities of 86.6% while the weighted average from the three variability RL tests was 75%. The LRA work also showed widespread discrepancies on recoveries, head grades and reagent consumptions. It was therefore decided not to use the LRA data for predictive recovery purposes.

11.2 Process Plant Design

The design of the process plant was based on the results of the metallurgical testwork program on the ore samples, as well as on strategic considerations by GSR, particularly with reference to annual throughput. The following basic design parameters were adopted:

Ore Blend: 1/3 Oxide, 2/3 Fresh Bond Ball Mill Work Indices: Fresh - 14.8 kWh/t

Oxide - 8 kWh/t Maximum Mill Power (Gross) 3 000 kW/mill (2 off) Annual Throughput 4,000,000 tonnes Availability 93% Operating Hours per Year 8147 Design Throughput 495 t/h Design Circulating Load ~350% Cyclone Overflow Solids Content ~42% Leach Residence 19.6 hours Recovery Overall 92.5%

11.2.1 Ore Characteristics for Plant Design Bulk Density - Oxide 1.5 t/m3 Bulk Density - Fresh 1.6 t/m3 Bulk Density - Blend 1.6 t/m3 Bulk Density - Phase 2 1.6 t/m³

Specific Gravity - Oxide 1.8 t/m3 Specific Gravity - Fresh 2.7 t/m3 Specific Gravity - Blend 2.7 t/m3 Specific Gravity - Phase 2 2.5 t/m³

Angle of repose 40º Angle of draw 60º Rod Mill Work Index - Blend 15.3 kWh/t



Ball Mill Work Index - Blend 14.7 kWh/t Bond Abrasion Index – Blend 0.29 Moisture Content - Oxide 15% Moisture Content - Fresh 8% Moisture Content - Blend 10% Moisture Content - Phase 2 20%

Plant Feed Grade - Oxide 1.2 gAu/t Plant Feed Grade - Fresh 2.1 gAu/t Plant Feed Grade - Blend 1.6 gAu/t Plant Feed Grade - Phase 2 0.6 gAu/t Specific Gravity Dry Carbon - Design 0.800 t/m³ Specific Gravity Wet Carbon - Design 1.370 t/m³ Bulk Density Wet Carbon - Design 0.5 t/m³ 11.2.2 Operating Schedule Annual Treatment Rate - Normal 3 500 000 tonnes Normal Annual Treatment Rate - Design 4 000 000 tonnes Maximum Operating Days Per Year 365 11.2.3 Crushing Operating Hours Per Day 24 Normal Operating Hours Per Day 16 Minimum Available Hours Per Year 5 840 Minimum Utilization 85 % Utilized Hours Per Year 4 964 Minimum Required Throughput 806 tph Maximum 11.2.4 Milling & CIL Operating Hours Per Day 24 Available Hours Per Year 8 760 Utilization 93 % Utilized Hours Per Year 8 147 Required Throughput 490 tph Maximum 11.2.5 Crushing The existing four stage crushing plant is to be utilized without modification. ROM Feed F100 1 000 mm ROM Feed F80 450 mm Crusher Product P100 9 mm



Crusher Product P80 6 mm Mill Stockpile Live Capacity 4 400 t Mill Stockpile Total Capacity 18 900 t 11.2.6 Heap Leach Reclaim Reclaim Stockpile Live Capacity 730 t Reclaim Stockpile Total Capacity 6 600 t Scrubber/Trommel Discharge Slurry 50% solids Scrubber/Trommel speed (Normal) 55 % of critical (new liners) Scrubber/Trommel screen Aperture 13 x 17.5 mm Scrubber/Trommel screen Flux 450 m³/m²/h Scrubber/Trommel Screen Dimensions 1.941mø x 2.135 ml 11.2.7 Milling The plant comprises two identical lines of single stage ball mills in closed circuit with cyclones. Number of Mills 2 Milling Rate 246 Dry Tph Per Mill Of Fresh Feed Mill Diameter Inside Shell 5.06 m Mill Effective Grinding Length 6.71 m Mill Speed ~75 % Of Critical Installed Motor Power 3 000 kW Trommel Aperture 13 x 17.5 mm Mill Product P80 85 µm Ball charge - Maximum 36 % v/v Circulating Load - Design 300 % Cyclone Overflow Slurry - design 42 % Solids By Mass Mill Discharge Slurry - Design 70 % Solids By Mass Maximum Ball size 50 mm (65mm optional at coarse crush) Mill Lining type Rubber Trash Screen Type Horizontal Vibrating Deck & Aperture 0.7mm x 3.3mm Vibroplast Polyurethane Panel 11.2.8 Gravity Two centrifugal primary concentrators followed by a shaking table for redressing of primary concentrates. Scalping Screen Type Horizontal Vibrating Screen Deck & Aperture 2mm x 2mm Polyurethane Panel Screen Feed Dry Solids 133 tph Concentrator Feed 120 tph Concentration Cycle Time > 1 hour



Dry Concentrate Arisings ~ 41 kg/batch Primary Concentrate Arisings < 1 tpd 11.2.9 CIL Total Number Of Tanks 6 off Tank Nominal Volume 2 500 m3 Total residence time ~19.5 hours at 3.5 MTPA Intertank Screen Type Mechanically swept cylindrical with pumping mechanism Intertank Screen Flux 95 m³/m²/h Intertank Screen Aperture 800 µm Stainless Steel Wedge Wire Loaded Carbon Screen Type Horizontal Vibrating Deck & Aperture 0.7mm x 3.3mm Vibroplast Polyurethane Panel Carbon Advance Operating Schedule Consecutive Carbon Advance Period 4.5 Hours/Day Over 3 x 1.5 Hour Periods Carbon Concentration 10 kg/m³ Carbon Type 6 x 12 Mesh Coconut Loaded Carbon Grade - Design 1 350 gAu/tonne Carbon Eluted Carbon Grade - Design <100 gAu/tonne Carbon

11.2.10 Acid Wash CIL Carbon Arisings 336 Tonnes/Month Maximum Acid Washes/Month - Design 28 Acid Wash Cone internal angle ~60 º Acid Wash Batch Size 12 tonnes Carbon Acid Wash Batch (Bed) Volume 24 m³ Acid Wash Temperature Ambient Acid Wash Flow Rate 2 Bed volumes/hour Acid Wash Strength ~ 3 % HCl Acid Wash Period ~ 2 hours Acid Wash Column Material Fibre Reinforced Plastic 11.2.11 Elution & Electrowinning Elution Type Pressure Zadra Elution Column Material 3CR12 Elutions/Month - Design 28 Eluate Solution ~3 % NaOH, ~1 % NaCN (Optional) Elution Flow Rate 2 Bed Volumes/Hour Elution Tank Temperature 90 ºC Ambient Temperature 25 ºC Elution Column Temperature ~140 °C Elution Tank Preheat Time 7 Hours Elution Column Fill/Empty Time 1 Hour Elution Column Heat Up Time 3 Hours



Elution Time 12 to 16 Hours 11.2.12 Regeneration Kiln operating schedule 20 Hours/Batch Carbon Regeneration Rate 600 kg/Hour Carbon Conditioning Batch Size 1 t Carbon Conditioning Solids 20 % Carbon Conditioning Tank Volume 6 m³ 11.2.13 Reagents

11.2.13.1 Lime Available CaO in bulk lime 60 % Lime packaging Road Tanker Lime Addition 1.1 kg/t as100% CaO 11.2.13.2 Cyanide NaCN packaging 1 ton Bag/Box Briquette NaCN briquette strength 97 % NaCN NaCN briquette bulk density 1 010 kg/m³ NaCN strength as made up 20 % m/m NaCN NaCN solution RD at 20% NaCN 1.10 t/m³ NaCN briquette strength 97 % NaCN NaCN briquette bulk density 1 010 kg/m³ NaCN Total Addition 0.36 kg/t as 100% NaCN Mill NaCN Addition 0.24 kg/t as 100% NaCN CIL NaCN Addition 0.12 kg/t as 100% NaCN NaCN Storage Tanks Capacity > 24 Hours Each 11.2.13.3 Caustic Soda Caustic Packaging 50 kg/Bag 99% NaOH Pearls/Flake Caustic Solution Strength 20 % m/m NaOH Density (RD) of NaOH solution 1.24 t/m³ 11.2.13.4 Hydrochloric Acid Hydrochloric acid packaging 1 m³ Isotainer or 200 Liter drum Hydrochloric Acid Strength ~33 % HCl m/m Density of 33% HCl Solution 1.15 t/m³ 11.2.13.5 Oxygen Oxygen Addition to CIL feed slurry 80 ppm or ~ 50 kg/h



Oxygen Addition to CIL 78 ppm or ~ 33.5 kg/h Total Oxygen Requirement 2000 kg/day

11.2.14 Tailings Tails Screen Type Horizontal vibrating Deck & Aperture 0.7mm x 3.3mm Vibroplast Polyurethane Panel Tails Return Water - Maximum 90 % of Normal Water Fed To Dam

11.3 Processing Operations

The process plant will comprise the existing quaternary Svedala crushing plant, which will feed through a 5000 tonne live capacity stockpile into two reconditioned 3 mW 16.5’ x 22.73’ Allis Chalmers ball mills operating in independent and parallel closed circuits. In this way, any down time on one mill will not require shut down of the total plant. These two mills are already on site, with the motors, gearboxes and girth rings currently being refurbished in South Africa. Mill specific, single stage cluster cyclones will be used to classify the mill discharge, with the coarse underflow being returned to the ball mill feed spouts in closed circuit arrangement, while the overflows will gravitate to a single trash removal screen. Cyclone feed can be diverted to feed independent gravity preparation screens, the underflows of which will feed two Knelson 48“ concentrators. The gravity circuits are mill specific although concentrate from each is routed to a common Gemini table. The trash screen underflow will be dosed with cyanide and pumped through a 900m long pipeline to a new 6 stage CIL circuit, each stage comprising a 2 500 m3 leach tank. This separation of the mills from the rest of the process plant is occasioned by the existing infrastructure at the plant site area, with the crushers for the leach pad operation being located some 900m from the gold ADR recovery section (Adsorbtion-Desorbtion-Recovery), where the new CIL plant will be built. It was deemed preferable to site the two ball mills adjacent to the crushers, and pump the slurry, rather than to site the mills at the CIL plant, and convey the crushed ore from the crushers. This approach has added benefits, in that the 900m long pipeline, in which the cyanided ore slurry is pumped under pressure, acts in essence act as a pipe reactor, which calculations indicate has the same effect on the process as building an additional CIL tank. Each tank will be charged with oxygen to a dissolved oxygen level of 100ppm, with oxygen also being fed into the pipeline at 5 bar pressure. A new acid wash, elution (12 ton / batch) and regeneration facility will be located alongside the CIL train. The CIL train is likely to operate as a hybrid direct leach / carbon-in-leach circuit, although mechanicals to allow a full CIL train have been costed. The CIL circuit will provide 19.6 hours of retention time and extra civil plinths will be cast to allow the on-line construction of a further two tanks if this is required once operation is underway. Tower cranes will be erected at the mill and CIL locations to speed up construction and reduce crane hire charges during this phase. The cranes will remain on site to handle maintenance issues.



Overall plant recovery is expected to be 92.5% for a blend of oxides and fresh ore, while treatment of heap leach material is expected to achieve 86% recovery for the Phase 2 material. Should Phase 1 leach material be processed, recoveries would be expected to be around 75%. For fresh ore, a bond work index of 14.8 kWh/t, cyanide consumption of 0.35 kg/t and lime consumption of 0.7 kg/t is expected. For oxide ore, a bond work index of 8 kWh/t, cyanide consumption of 0.35 kg/t and lime consumption of 1.44 kg/t is expected. It should be noted that the Ammtec lime consumptions from their testwork are higher than the LRA figures quoted here, however this is primarily because Ammtec uses a 60% activity lime whereas LRA uses a 95% active lime source. For the Phase 2 heap leach material, a bond work index of 11.48 kWh/t, cyanide consumption of 0.35 kg/t and lime consumption of 0.7 kg/t is expected. Reclamation from the pads will be carried out by front end loaders, which will feed into a hopper on the existing conveyor belt systems, which have already been turned around. The conveyors will feed into a scrubber, with the slurried material being gravity fed through a pie launder into the mill discharge sumps. The material will subsequently be cycled into the mills for grinding.

12 Infrastructure 12.1 Tailings Storage Facility

A significant change to the Wassa operation from its previous mode of operation is the need for a tailings storage facility (TSF) to accommodate the tailings from the CIL process. Knight Piésold consulting were engaged to carry out the carry out the design and engineering of the facility. Knight Piésold, in conjunction with Golden Star, initially carried out a conceptual TSF site options study. The study identified two potential sites, one located in the valley immediately west of and adjacent to the Phase 2 heap leach pad area and a second in a valley one kilometer southwest of the village of Akyempim. Conceptual designs for both sites were developed, and each site was evaluated in terms of development costs, operating costs, land take, rehabilitation considerations, flexibility of operation and ease of expansion. The study concluded that the Akyempim site provided the more cost effective storage with the lowest initial development cost. However, the Akyempim site was located in an area of previously undisturbed ground and therefore the facility was considered to be environmentally less favorable than the Phase 2 heap leach pad site, which would be mostly located on land already disturbed by the previous mining operation. The crop compensation, pumping and ultimate operating costs were also lower for the heap leach site. In addition, the Akyempim site was also located on an area identified as very prospective. After taking account of these considerations, the heap leach site was selected as the preferred site.



A geotechnical evaluation of the preferred site concluded that the embankments will be founded on competent residual soils and weathered rock and the site is therefore suitable for the proposed use. The design of the TSF is a conventional saddle dam arrangement, with a starter wall being constructed as the lower end on a valley, and gradually raised as the levels of tailings increases. Maximum benefit was taken of the local topography to minimize additional construction costs. This was particularly so with regards to the Phase 2 heap leach pad area. The processing plan envisages the rehandling and processing of the bulk of the Phase 2 pad material, and as this is removed to be processed, it opens up further space for the expansion of the TSF into that already disturbed area, thereby minimizing any additional greenfield land take. The TSF as currently designed has a storage capacity of approximately 21 million tonnes, deposited at an average annual rate of 3.5 million tonnes per annum at a deposition density of 45%. This is sufficient to accommodate all 14.7 million tonnes of material within the current mining and processing schedule, as well as providing surplus capacity for another two years of plant feed.

12.2 Power Supply

The electricity supply in Ghana is principally provided by the Volta River Authority (VRA) through a hydroelectric power station from the Akosombo Dam (at the bottom of the Volta Lake). Power provided from the dam supplies the national grid system, which links to the major settlements in the country such as Accra, Kumasi, Tamale and Takoradi. This supply is augmented with power from the Aboadze Thermal Power Station near Takoradi, as well as by imports from neighboring Côte d’Ivoire.

Power in the Western Region is provided from Takoradi to the larger settlements, e.g. Prestea, Bogoso, and Tarkwa, while electricity supply to smaller settlements remains very limited. The national grid has been extended to supply the Abosso gold mine (owned by Goldfields Ghana), approximately 20km distant, but there is currently no national grid supply to the Wassa project area.

The previous operators therefore generated their own power, using five diesel-fuelled Caterpillar 3512TA 1.25 mVA generator sets. This provided sufficient power for the crushers and the pumps to the leach pads, and the ADR facility, with power also being fed to the main accommodation camp and the nearby villages of Akyempim and Kubekro. However, this generator plant will be inadequate to supply the two ball mills and the rest of the CIL plant, and alternatives were investigated. An approach was made to Abosso, to continue their line through to Wassa, but this was rejected by Abosso. It is believed that they are operating to the maximum of the line, and there would be insufficient power to be passed through to Wassa. It was decided therefore to tap off the main 161 kV powerline that runs from Prestea through to Dunkwa, approximately 15km from the Bogoso sub-station and the BGL minesite. The route selected between the tap off point and the Wassa project is approximately 35km, and traverses a 2km stretch of the Bonsa Ben Forest reserve. The towers to be used are innovative to Ghana, in that they will be galvanized steel lattice pivot towers, rather than the more conventional free standing towers. These types of pylons are relatively common in



other parts on the world, including South Africa, from where they will be sourced. They have been tested for 220 kV conductors, and are 35m in height, and can be spaced approximately 300 to 400 meters apart. The advantage of these towers is that the height will assist in the clearing of vegetation and minimize maintenance, and their design allows modification of the line route to bypass obstacles like villages and bad topography (unlike the conventional free standing pillions that require a more direct route). The height of the towers will also reduce the number of towers required minimizing environmental impact. They are light, simple and quick to erect with minimum civil works required.

The conductor will be capable of transmitting 600 amps or just below 200 MVA and will be the Pelican type conforming to the normal VRA specification for high-tension transmission lines. Wassa will require only 57 amps or 16 MVA for the project, so the additional capacity will allow the VRA to expand the supply to other areas in the region.

At Wassa a full transformer bay will be installed which will consist of a 161 kV to 34.5 kV transformer rated at 25/33 MVA. This will subsequently be transformed down to 6.6 kV. The installation of the power line will be funded by GSR, but once installed, it will become the property of the VRA, as an asset to the nation. The project will obtain a rebate from VRA over the first three years of project life, as recompense.

12.3 Existing Infrastructure

The development of the Wassa project by the previous operators resulted in significant infrastructure, which has been acquired by GSR as part of the asset purchase agreement. Major items of this infrastructure which will be utilized by the current project include: i) Administration facilities: Administration and technical services offices, warehouses, light

vehicle workshops, and security buildings. ii) Process Plant: The Wassa ADR (Adsorbtion-Desorbtion-Recovery) recovery plant, together

with a four stage crusher circuit, overland conveyors, process water ponds, gold room, assay laboratory, plant workshops and offices

iii) Mining infrastructure: haulroads, ROM crusher pad iv) Fuel farm: There are two 500,000 litre fuel storage tanks on site, owned by a major

multinational oil company v) Accommodation: A Senior Staff camp, comprising 28 two and three bedroom houses, 26 single

quarters, a guesthouse, and a clubhouse and recreational facilities. In May 2003, GSR reached agreement with the mining contractors for the previous operator, to acquire from them a fully equipped 4 bay heavy equipment workshop and associated offices. Together with this purchase came 5 additional houses at the adjacent to the senior staff accommodation. There is therefore very little additional infrastructure that needs to be provided for the operation, with the exception of the tailings storage facility and the 161kV power line as described above.



13 Environmental During the initial feasibility study carried out by the previous operator into the Wassa property, a preliminary environmental baseline study over the project site was undertaken by SGS Laboratory Services Ghana Ltd (SGS). When the final project feasibility study was produced in September 1997, and was accepted by the joint venture partners, SGS were further retained to complete the formal Environmental Impact Assessment (EIA) and to produce the Environmental Impact Statement (EIS) for submission to the Ghanaian Environmental Protection Agency (EPA). This EIS was presented to the EPA in February 1998, and subsequently approved. The SGS EIS covered the then planned area of the project, including pits, waste dumps, roads etc. However concerns raised by a local community at Kubekro village resulted in SGL reaching a separate agreement with this community, enabling later access to the Deadman’s Hill pit area. A Costed Reclamation Plan was subsequently submitted to the EPA to cover the ongoing and planned operations in the area. When GSR acquired the property in late 2002, the EPA were approached to clarify the environmental legislative and reporting requirements involved in reopening the project. Although GSR initially believed that a supplement to the previously approved EIS would be sufficient for the project, the EPA has indicated that a full EIA and EIS will be required for the project. Their rationale is that while the pits and waste dumps would be within the same general areas as planned and permitted with the previous EIS, the use of a CIL plant would require the construction of a tailings storage facility, and the diversion of a section of the main Akyempim to Atieku public road, which would be a significant change to the previous project description. GSR has therefore retained independent environmental consultants, Scott Wilson Mining (SWM), to carry out the environmental assessments of the project, and to produce the relevant environmental impact reports. The first stage of the EIA process, as required by Ghanaian legislation, was the preparation of a Scoping Report. This preparation included extensive consultations with local and nearby communities, as well as with other interested and affected parties. A draft copy of the report was submitted to the EPA in early April 2003, with copies also being made available to the district and regional authorities, and to the general public. At the request of the EPA, a Public Forum was held in May 2003, where GSR was given the opportunity to formally present the project to local chiefs, dignitaries and the public. The aim of the public forum was to elicit from potentially affected parties any specific concerns that they may have with the project, and to ensure that these concerns are adequately addressed during the EIA process. Following the Public Forum, the EPA identified three key areas of concern:

a) Employment opportunities for locals in the project area b) The safety aspects of the new tailings storage facility and the impacts that the road diversion

around the TSF would have on pedestrians. c) The adequate payment of compensation for land required for operational purposes.



The EPA were satisfied with the conduct and outcome of the Public Forum, in that it raised no significant technical issues that could not be addressed, and specifically in that there were no objections of social or community grounds. As with any project in a populated area, the social considerations are a high priority to all stakeholders. Official feedback from the EPA on the Public Forum and the Scoping Report were received mid-May, with a final copy of the report being delivered mid-June. Following from the feedback on the Scoping Report, the full EIA was commenced, and it is expected that the EIS will be submitted to the EPA by end-August. Legislation allows the EPA up to 90 days to review and respond to the issues raised and addressed in the EIS, which is within the overall project timeframe. Based on the positive reception at the Public Forum, GSR believes that this time period can be significantly shortened. To specifically address the issues raised in the public forum:

i) GSR, in consultations with local chiefs, has complied a list of former employees at the Wassa project, together with other potential applicants, with the key issue being that they come from the local area. It is GSR policy, as evidenced at its Bogoso and Prestea mine sites, to employee local people where they have the appropriate skills and experience. Where specific technical skills may be required, then it is often necessary to search further afield.

ii) The TSF design has been carried out by reputable international consultants, who have been, and are still, the TSF consultants to the Bogoso/Prestea mine. GSR is addressing the perceived safety issue raised during the EIA process, by conducting ongoing consultations with nearby communities.

iii) Compensation payments are an issue that had on occasion been poorly handled by the previous operator, and consequently GSR inherited the mistrust engendered. GSR has attempted to be transparent in dealings with all parties potentially affected by compensation claims, which principally affects the persons carrying out farming in the site of the new TSF.

It should be noted that there are no villages or communities that need to be resettled or relocated in the establishment of this project. As a separate issue, the EPA requested a separate EIA be carried out for the power line route, and this exercise is currently underway.

Although the formal Environmental Impact Assessment is still underway, the Scoping Report for the Wassa project did not raise any contentious issues. The key factor in this project is that this has already been an operating mine site, approved and permitted by the EPA. The mining operations will be taking place within the areas previously defined as potential mining areas, and the waste dumps will be within areas previously defined for that purpose. In addition, the construction of the process plant and milling area are both within the fenced off site that encompasses the industrial and processing activities of the previous operator, and as such, there will be no new land take involved in setting up the mining and processing aspects of the operations.



The main significant aspect is the new tailings storage facility, and this has been designed to encompass as much as practical of the existing disturbed areas of the heap leach pads, thereby reducing the area of greenfield landtake. The EPA have been to the Wassa site several times since GSR commenced activities, most noticeably for the Public Forum, and there is a regular monthly inspection by the District Environmental Officer from Tarkwa. GSR are therefore confident that an Environmental Permit will be issued without any significant restrictions. During the preparation of the EIS, a Mine Closure Plan will be prepared, as well as a provisional Costed Reclamation Plan. This will be used as the basis for discussions with the EPA on the posting of suitable reclamation bonds. Past practice by GSR on the Prestea Concession has been for the company to pay a cash bond of a percentage of the full reclamation amount, and provide a bank guarantee for the balance, which complies with the EPA’s requirements.

14 Economic Analysis

14.1 Capital Cost Estimates

The annual capital expenditure estimated for the life of the project are shown in the table below:

Capital Expenditure (US$ million) By Major Item

2003 2004 2005 2006 2007

Community Development 0.09 0.07 0.05 0.02

Pre Development Cost 1.20

Mining Fleet & Development 1.39 14.05 1,06 0.10 0.05

Processing Plant

18.30 1.21

Tailings Dam Construction 0.81 0.61 0.24 0.43

Administration

0.64 0.13 0.13 0.09 0.01

EPA Bond 0.26 0.09

Contingency 0.50 0.50

Total 22.84 16.85 1.59 0.67 0.08

The capital costs estimates are based on recent quotations for new mining equipment, a formal quotation for refurbishing the mills and ancillary items, new plant equipment and materials, and a fixed price lump sum proposal for process plant construction. Tailings construction costs are based on an



independent verified design and construction estimate. Other capital costs were derived from quotations and estimates based on operational experience in Ghana. The capital expenditure for the project will be expended in two main phases during each of the first two years of project life. The first phase for the development of Wassa of $25.5 million will be 100% funded by GSR. The project is expected to be completed in the first quarter of 2004 and will be funded from existing cash resources. When mining commences in December 2004, it is envisaged that a fleet of three 110-tonne backhoe excavators and nine 95-tonne dump trucks will be used to mine approximately 35,000 tonnes per day of ore and other material. The mining fleet is expected to be Company owned, which will require an additional capital expenditure of approximately $14.2 million between 2004 and 2005. Vendor financing is currently available for 85% of the cost. Total capital expenditure during life of mine will be $42 million. A summary of capital expenditure for the process plant construction and for the mining equipment fleet is shown below. Capital costs have been based on the following exchange rates to the US dollar: South African Rand = 8.42

Euro = 1.13

Process Plant Capital Estimate

Major Items US$ 000,000 %

Mechanicals Electrical Structural steel & Plate Work Piping & Valves Earthworks & Civils Infrastructure (Power line) Transport EPCM - Engineering - Procurement - Construction Management

- Commissioning PC Sums Total

3.115 1.831 1.699 0.473 0.617 3.339 1.009

0.595 0.071 1.491 0.070 1,030

15.308

20 12 11 3 4 22 7

4

0.05 10

0.05 7

100

It is noted that the current exchange rate between the South African Rand and the US$ is at approximately 7.45 : 1. However, the contract to construct the process plant is on a lump sum turnkey basis, awarded to MDM under FIDIC (Federation Internationale des Ingenieurs-Conseils) conditions of contract, denominated in US dollars. GSR is thus not impacted by the variability of the exchange rates.



Mining Equipment Capital Estimate

Major Items US$ ‘000,000 % 9 * Cat 777 D Haul Trucks 3 * Liebherr 984C Excavators 3 * Tamrock Pantera 1500 Drill Rigs 2 * Cat 988G Wheel Loader 3 * Cat Track Dozers 1 * Cat Wheel Dozer 1 * Cat Grader Other Equipment Light Vehicles Total

6.84 2.40 1.35 0.94 1.44 0.34 0.27 1.20 0.30

15.08

45 16 9 6

10 2 2 8 2

100

It should be noted that all the above mining equipment has been based on supplier quotations for new items, and these costs have been used in the economic evaluation of the project. However, during the phase when the leach pads are being processed, prior to the start of mining operations, a detailed study will be carried out into the alternative use of contract mining. Should this demonstrate significant benefits, and should this approach be adopted, then the capital outlined above will be reduced, while unit operating costs will commensurately rise.

14.2 Operating Cost Estimates

Operating costs were determined through a combination of first principles using a detailed itemized budget, and from experience gained at GSR operations at Bogoso/Prestea. All costs were reviewed and verified by independent consultants, including MDM for process costs, SRK for mining costs, and through consultation with equipment suppliers including the Caterpillar agents, Unatrac Ltd. The average cash cost, excluding royalties, over the life of the project is $7.08 per tonne of ore processed or $192 per once of gold produced. Operating cost remains relatively constant throughout the project life at around $200 per ounce except for the third year when the cost reduce to $165 per ounce.

By Major Section $/t of Ore $/ounce Mining 2.72 74 Processing 3.58 97 G&A 0.88 23 Power Rebate -0.10 -2 Cash Cost 7.08 192 Government Royalty 0.36 10 Total Cash Cost 7.44 202



The operating costs broken down into each major component of the three main areas of mining and processing and G&A are:

By Major Section $/t of Ore Processed $/t of Material Mined Mining Manpower Drilling Blasting Loading Hauling Support Fleet Services Fixed Cost Grade Control Wall Control Total Mining

Processing Manpower Power Reagents Maintenance Services Total Processing Power Rebate

G&A Manpower Services Environmental Total G&A

Total Operating Cost

0.37 0.21 0.53 0.31 0.64 0.32 0.06 0.09 0.17 0.02 2.72 0.28 0.98 1.56 0.50 0.26 3.58

-0.10 0.15 0.66 0.07 0.88 7.08

0.14 0.08 0.21 0.12 0.25 0.12 0.02 0.04 0.07

0.01 1.06

14.3 Taxes and other payments

Mining projects in Ghana are subject to the following key fiscal and tax considerations:

i) Royalty of 3% increasing to up a cap of 12% subject to certain operating return tests, ii) Corporate tax rate of 32.5%,



iii) Capital allowance of 75% in year of capital expenditure with balance of capital expenditure subject to a 50% capital allowance in subsequent years. In addition, a bonus Investment Allowance of 5% is earned in year of capital expenditure,

iv) Dividends to shareholders are taxed at 15%, and v) No import duties or sales tax is imposed on any items purchased for the mining activity

which are included on a “Mining List” agreed between the Government and mining industry. Items not on the mining list incur import duties and sales taxes at a rate of 15%.

14.4 Economic Analysis

Wassa will commence operations in 2004 by processing material on the previous operator’s heap leach pads. The early reprocessing of this material will make the leach pad area available for tailings containment. Production from the heap leach material in 2004 will be about 75,000 ounces of gold at an estimated cash cost of $211 per ounce. Open pit mining development will commence in December 2004, with open pit ore mining and processing starting in early 2004. This will result in an increase gold production to more than 140,000 ounces of gold per year, at an average cash cost of less than $200 per ounce. The Feasibility Study completed in June demonstrates that the Wassa project is economically viable with a positive Internal Rate of Return at gold prices above $275 per ounce. Based solely on the current Probable Mineral Reserves, and an average gold price over the life of the mine of $325 per ounce, the after-tax undiscounted Net Present Value for the project is estimated to be $22 million with an IRR of 27%. The NPV and IRR analysis is inclusive of a 3% government royalty and an allowance of $3.5 million for post-mining reclamation. The analysis assumes Company-owned mining equipment, with 85% of the capital cost of the mining fleet being debt financed. Property acquisition costs are excluded from the analysis. Golden Star plans to fund the Wassa construction cost from its existing cash reserves. The economic evaluation is highly sensitive to the gold price, as is shown in the table below:

Gold Price in $/ounce Parameter Units $300 $325 $350 $400

Gold Production NPV (Undiscounted) IRR Project Payback Cash Cost Capital Cost Project Life

Ounces $ million % Years $/ounce $ million Years

545,309

12 16 3

201 42 4

545,309

22 27 2.5 201 42 4

545,309

30 38 2.2 201 42 4

545,309

48 58 1.7 201 42 4



A sensitivity analysis conducted on the effects of changes in gold price, recovery, capital cost and operating cost concluded that the project is equally and most sensitive to changes in recovery, grade and gold price. The project is less sensitive to changes in operating cost and least sensitive to changes in capital cost. The tables below depicts the change in the IRR and NPV with change in gold price, capital costs, and operating costs.

Gold Price - 10 % 0 % + 10 % NPV $8.9m $22.1m $33.3m IRR 11 % 27 % 41 %

Capital Expenditure

- 10 % 0 % + 10 % NPV $24.5m $22.1m $19.8m IRR 34 % 27 % 22 %

Operating Expenditure

- 10 % 0 % + 10 % NPV $29.0m $22.1m $15.4m IRR 36 % 27 % 18 %

A full economic spreadsheet for the project is shown in the table overleaf.





14.5 Mine Life

The mine life as defined by the Probable Reserves is currently four years, inclusive of the material from the Phase 2 heap leach pad. It is recognized that this is a relatively short life, with all its associated risks. However, significant exploration work has been carried out in and around the existing mine workings, and a large Inferred Resource has been identified. For planning purposes and to guide additional drilling programs in the mineralized areas, GSR has carried out pit optimizations, which include both the Indicated and the Inferred Resources.

Units $275/oz $300/oz $325/oz

Mt 10.1 11.9 13.2 g/t Au 1.65 1.59 1.54

Indicated Mineral Resource

within Optimized Pit Ounces Au 536,000 588,000 654,000 Mt 1.8 3.0 4.3 g/t Au 1.72 1.65 1.56

Inferred Mineral Resource

within Optimized Pit Ounces Au 96,000 159,000 215,000

Other Material Mt 34.0 46.4 58.0 Strip Ratio 2.8:1 3.1:1 3.3:1

It can be seen that the conversion of the Inferred Resources into an Indicated Category would have a major impact on the available Mineral Reserves, with consequent increases in mine life and improvements in overall project economics. The Inferred Resources inside the $325/oz optimized pit envelope will be the subject of an infill drilling program in late 2003. This will be carried out specifically to improve the confidence in the resources to enable them to be moved from the Inferred Category into the Indicated Category. A sum of $900,000 has been allocated for this exercise. There are no plans at this stage, prior to in-pit grade control, to further drill the resource to prove up any Measured Resources. Caution: The term "inferred mineral resource" is used in conformity with Canadian regulatory authorities. It is not recognized by the US Securities and Exchange Commission. "Inferred mineral resources" have a great amount of uncertainty as to their existence, and great uncertainty as to their economic and legal feasibility. It cannot be assumed that all or any part of the inferred mineral resources will ever be upgraded to a higher category. The exploration potential of the Wassa Lease is high, despite the challenges imposed by the superimposed folding pattern of the ore bodies as observed in the current Wassa pits. However a better understanding of the geology in conjunction with existing geochemistry data has generated additional targets south of the main Wassa deposit, which are currently at various stages of exploration. South-



Akyempim, Ballyebo, Nsa and Bawdia Bosso are a few of the more advanced targets that have been outlined by soil geochemistry and tested with Rotary Air Blast (RAB) drilling. These targets require further RAB and RC drilling but initial results indicate there are primary mineralized systems underlying the gold in soil anomalies. Additional gold in soil anomalies exist in the southern portion of the mining lease. These have not yet been thoroughly evaluated as the targets lie within the Subri Forest Reserve. As it has done at four other mining properties in Ghana (Chirano, Kubi, Mampon and Akyim), the Government has indicated a willingness to allow exploitation in the forest reserve in the southern part of the Wassa concession, subject to the usual, and more stringent, environmental conditions. Wassa is unique amongst these five sites within a forest reserve in that there is an existing Mining Lease over the concession, rather than an Exploration License. Work programs and budgets have been put together to evaluate the subsurface extent of these untested anomalies with work scheduled to start in the second half of 2003.

15 Interpretations and Conclusions

The Wassa Project has been fully detailed in a Feasibility Study carried out by GSR under the auspices of the consulting firm Metallurgical Design and Management (“MDM”). Much of the work that has gone into the study was carried out by employees of Bogoso Gold Limited, a 90% owned subsidiary of GSR. Reputable external consultants have been used to provide review of the work carried out by these employees, while specialist consultants have been used directly for specific aspects of the study. The Feasibility Study has demonstrated the economic viability of the Wassa Project, with a positive Internal Rate of Return being achieved at a gold price of $275 per ounce. At a gold price of $325 per ounce, the IRR is 27%, with an after-tax NPV of $22 million

16 Recommendations

The Wassa project has been demonstrated through the Feasibility Study to be viable, with a planned life of four years, based solely on the Indicated Resources portion of the total Wassa Mineral Resources. However, there is the potential to convert a significant proportion of the Inferred Resources into the Indicated Resources category through further exploration drilling. This would provide a larger resource base on which carry out pit optimizations and designs, would it is likely that the Mineral Reserves will increase as a result, thereby increasing the mine life. GSR therefore plans to spend approximately $900,000 on a drilling program in late 2003 specifically to improve the confidence in the resources, and a new Mineral Reserve Estimate will be prepared early 2004. The mining of the Wassa deposit has currently been assumed to be carried out by the mine itself, using mine owned equipment. Quotations have been obtained for all major items of mining plant. In view of the potentially short life of the operation, GSR plans to carry out an in depth exercise to evaluate the merits (or otherwise) of establishing a contract mining scenario in place of ‘owner operator’. This will be done during the processing phase of the heap leach material, prior to the start of mining operations in late 2004.



17 References June 2003 Gold Recovery Testwork conducted on samples from the Wassa Project for Bogoso

Gold Limited. Report No. A8782, Ammtec Metallurgical Laboratories, Perth, Australia June 2003 Laboratory Testwork to evaluate gold bearing ore samples from Bogoso Gold

Limited. Report No. Met 03/G91, Lakefield Research laboratories (SGS), Johannesburg, South Africa

Claffey, D.: June 2003 Wassa Mine Feasibility Study Tailings Storage Facility Feasibility Design Report.

Report No. 15801R2, Knight Piésold Ghana Limited, Accra, Ghana. Marshall, N.: March 2003 Wassa Open Pit Geotechnical Review. Report No. U2145, Steffen Robertson and

Kirsten Consulting, UK McCandlish, K.: April 2003 Qualifying Report for the First Disclosure of a Resource Estimate on a Material

Property, Wassa Mine, Southwest Ghana. Report No. 03PM67, Associated Mining Consultants Ltd., Calgary, Canada

Skelton, R., et. al: July 2003: Review of the Mining Section of the Base Case $300 Indicated Resource for the Wassa

Feasibility Study. Report No. U1940, Steffen Robertson and Kirsten Consulting, UK Tarling, J.: June 2003 Wassa Project Environmental Scoping Report. Report No. 11749, Scott Wilson Mining,

UK



18 Plans

The following plans are attached: Fig 3.0: Wassa Local Geology Superimposed on Regional Geology of South West Ghana Fig 3.1: Wassa Local Geology Drill Hole Location Figure 1: Wassa Project Site Outline – The Wassa mining lease boundary and the project location Figure 2: Wassa Project Existing Mine Infrastructure – The current pit, waste dump and leach pad

outlines Figure 3: Wassa Project Plan of New Mine Development – The planned expansions to the open pits

and waste dumps, and the planned new tailings storage facility

19 Date This Technical Report was issued on 1st August 2003

BLEKROMONCESSION

ASEMPACONCESSION

AMENFI 2CONCESSION

CO.ON

INTER AFRIQUEAPPLICATION

PAMPECONCESSION

PRESTEACONCESSION

BOGOSOCONCESSION

AMENFI 1CONCESSION

KOBRACONCESSION

CADEMMACONCESSION

INTER AFRIQUEAPPLICATION

FLAGBASE WESTCONCESSION

FLAGBASE WESTCONCESSION

RIYADH EASTCONCESSION

RIYADHCONCESSION

FLAGBASE 2CONCESSION

FLAGBASE EASTCONCESSION

AXM

INCO

NCES

SIO

N

NETASCONCESSION

FOREMOST "A" CONCESSION

ASHEBA-KANYAKAWCONCESSION

AXMINCONCESSION

BRO

TET

CO

NC

ESSI

ON

ST J

UD

E C

ON

CES

SIO

N

FOR

EMO

ST "B

" CO

NC

ESSI

ON

FairstarExplorationsIncorporated

FairstarExplorationsIncorporated

WEXFORD GOLDFIELDSAPPLICATION II

WEX

FOR

D G

OLD

FIEL

DS

APPL

ICAT

ION

I

WEXFORD GOLDFIELDSLIMITED

WASSA PROPERTIES

ASHANTI - A

KROPONG GOLD BELT

2100

00 m

E

2100

00 m

E

1800

00 m

E

1500

00 m

E

130000 mN

100000 mN 100000 mN

130000 mN

1800

00 m

E

40000 mN 40000 mN

70000 mN 70000 mN

1500

00 m

E

0 5 10

kilometers

Scale: 1:125000

Drawing:

Projection: Ghana National Metric (Clarke 1880/War Office 1924)

Date:10/2/2003

Author: J. Amekudi

Office: Wassa

Quartzites, grits/ mature quartzose oligomictic conglomerates

Sandstones, quartzites and grits with phyllites (Dompim Phyllite)

Dolerite and gabbro sills

Dolerite Dykes

Sedimentary cover

GEOLOGICAL LEGEND FOR SOUTH WEST GHANA

Gabbro and diorite (possibly of Dixcove Suite affinity)

Hornblende bearing granite, granodirite, monzonite, porphyry, diorite and aplite

Biotite and/or muscovite bearing granite,granodiorite, pegmatite and aplite

Upper Birimian

Lower Birimian

Kawere Group

Birimian System

Basic to intermediate metavolcanics, volcaniclastics, greywackes and phyllites

Phyllite and chloritoid phyllite with subsidiary arenaceous beds

Phyllites, schists, tuffs and greywackes

Sandstones, quartzites, grits and immature polymictic conglomerate

Huni Sandstone

Tarkwa Phyllite

Banket Series

Tarkwaian System

Intrusives

Dixcove Suite Granitoids

Cape Coast Suite Granitoids

PHAN

ERO

ZOIC

PRO

TER

OZO

IC

WASSA GOLDMINES

WEXFORD GOLDFIELDS LTD.

FIG 3.0

Wassa Local GeologySuperimposed on

Regional Geology of SW GhanaWith Mineral Occurranceand Soil Geochem Image

Crusher Area

00 mN

4000

0 m

E

20000 mN

4000

0 m

E

21500 mN00 mN

4090

0 m

E

4070

0 m

E

4050

0 m

E

4010

0 m

E

4030

0 m

E

3930

0 m

E

3950

0 m

E

3970

0 m

E

3990

0 m

E

ME

ST

M2N

FS

WASTE DUMP 2

ME

DMH

WASTE DUMP 3

BS

242

M2S

SE

419

BSWASTE DUMP 1

DMRB

DMRB101

DMRB103

DMRB059

DMRB055

DMRB100

DMRB052

DMRB051

DMRB050

DMRB058

DMRB104

DMRB057

DMRB069

DMRB056

DMRB083

DMRB047

DMRB054

DMRB061

DMRB046

DMRB053

DMRB045

DMRB044

DMRB062

DMRB084DM

RB060

DMRB028

DMRB027

DMRB036

DMRB023

DMRB026DM

RB025

DMRB024

DMRB019

DMRB018

DMRB017

DMRB016

DMRB020

DMRB015

DMRB063

DMRB064

DMRB078

DMRB099

DMRB049

DMRB079

DMRB048

DMRB038

DMRB021

DMRB085

DMRB039

DMRB040

DMRB043

DMRB098

DMRB065

DMRB070

DMRB030

DMRB032

DMRB071

DMRB031

DMRB029

DMRB072

DMRB042

DMRB073

DMRB035

DMRB074

DMRB075

DMRB080

DMRB068

DMRB041

DMRB077

DMRB067

DMRB076

DMRB022

DMRB081

DMRB066

DMRB037

DMRB014

DMRB105

DMRB011

DMRB010

DMRB009

DMRB082

DMRB008

DMRB012

DMRB013

DMRB090

DMRB086

DMRB091

DMRB094

DMRB087

DMRB088

DMRB092

DMRB089

DMRB093DMRB095

DMRB096

DMRB097

DMRB001

DMRB033

DMRB034

DMRB002

DMRB006

DMRB003

DMRB007

DMRB004

DMRB005

MERB011MERB012MERB013MERB023MERB008

MERB004

MERB007MERB009

MERB010

MERB003

SERB132

SERB149

SERB129

SERB131

SERB135

SERB134

SERB130

SERB127

SERB128

SERB133

SERB143

SERB136

SERB137

SERB142

SERB043

SERB042

SERB145

SERB041

SERB141

SERB154

SERB018

SERB147

SERB029

SERB033

SERB028

SERB155

SERB030

SERB162

SERB163

SERB027

SERB077

SERB037

SERB078

SERB096

SERB059

SERB038

SERB025

SERB026

SERB055

SERB056 SER

B100

SERB102

SERB101SER

B017SER

B054SERB099

SERB016

SERB053

SERB081

SERB060

SERB061

SERB082

SERB079

SERB080

SERB095

SERB089

SERB094

SERB024

SERB086

SERB088

SERB093

SERB085

SERB092

SERB087

SERB091

SERB057

SERB084

SERB153

SERB045

SERB040

SERB044

SERB140

SERB168

SERB167

SERB039

SERB146

SERB150

SERB151

SERB169

SERB139

SERB138

SERB156

SERB160

SERB161

SERB034

SERB035

SERB031

SERB032

SERB157

SERB158

SERB159

SERB020

SERB014

SERB019

SERB148 SER

B015SER

B049SER

B050

SERB103

SERB013

SERB011

SERB104

SERB144

SERB012

SERB052

SERB008

SERB152

SERB110

SERB009

SERB111

SERB036

SERB010

SERB109

SERB108

SERB112

SERB113

SERB105

SERB106

SERB107

SERB021

SERB022

SERB090

SERB023

SERB083

SERB098

SERB058

SERB097

FSRB022

FSRB016

SERB073

SERB164

SERB165

SERB074

SERB006

SERB007

SERB005

SERB069

SERB066

SERB062

SERB064

SERB065

FSRB018

FSRB019

FSRB135

SERB063

FSRB134

FSRB133

FSRB099

FSRB021

FSRB138FSR

B020

FSRB136

FSRB045

FSRB046

FSRB047

FSRB048

FSRB049

FSRB050

FSRB051

FSRB137

FSRB130

FSRB015

FSRB017

FSRB131

FSRB014

FSRB013

FSRB132

FSRB083

FSRB098

FSRB012

FSRB082

SERB051

FSRB007

FSRB003

FSRB006

FSRB001

FSRB002

FSRB005

FSRB004

FSRB156

FSRB029

FSRB155

FSRB027

FSRB154

FSRB028

FSRB009

FSRB077

FSRB011

FSRB081

FSRB008

FSRB078

FSRB079

FSRB010

FSRB080

FSRB038

FSRB127

FSRB037

FSRB126

FSRB036

BSRB014

BSRB008

BSRB011

FSRB035

BSRB009

BSRB012

MERB020

MERB019

MERB018

MERB034

MERB033

MERB036

MERB017

MERB032

MERB038

MERB068

MERB022

MERB024

MERB025

MERB027

MERB026

MERB031

MERB001

MERB028

MERB030

FSRB054

FSRB100

FSRB139

FSRB086

FSRB087

FSRB088

FSRB053

FSRB085

FSRB140

FSRB141

FSRB052

FSRB084

FSRB142

MERB037

MERB035

MERB040

MERB039

MERB043

MERB060

MERB072

MERB041

MERB061

MERB071

MERB042MERB058

MERB059

MERB070MERB069

MERB016

MERB050

MERB005

MERB015

MERB014

MERB048

MERB049

MERB046

MERB047

MERB044MERB045

MERB062

MERB063

MERB006

MERB002

MERB021

MERB051

MERB029

MERB055MERB056

MERB052

MERB054MERB057

MERB073

MERB074MERB053

MERB067

BSRB013

BSRB010

M2SR

B003

BSRB015

M2SR

B002

M2SR

B001

STRB001

STRB003

STRB002

STRB017

STRB004

STRB034

STRB033

STRB018

STRB005

STRB006

STRB010

STRB030

STRB007

STRB02

9

STRB011

STRB032

STRB019

STRB028

M2NRB002M2NRB035

M2NRB003

M2NRB001

M2NRB045

M2NRB014M2NRB004

M2NRB049

M2NRB050

M2NRB036M2NRB037

M2NRB024

M2NRB039

M2NRB006

M2NRB010

M2NRB038

M2NRB005

STRB035

STRB036STRB037

STRB038

STRB014

STRB008

STRB012

STRB013

STRB025

STRB026

STRB021

STRB023

STRB009

STRB027

STRB022

STRB016

STRB020

M2NRB047

STRB031

M2NRB046

M2NRB012

M2NRB044

M2NRB025

M2NRB043

M2NRB026M2NRB008

M2NRB048

M2NRB053

M2NRB054

M2NRB052

M2NRB007

M2NRB015

M2NRB051

M2NRB018

M2NRB016

M2NRB019

M2NRB013

M2NRB017

M2NRB027

M2NRB021

M2NRB028

M2NRB022

STRB024STRB015

M2NRB009

M2NRB011

M2NRB029

M2NRB020

M2NRB030

M2NRB031

242RB017

M2NRB032

M2NRB023

242RB016

242RB007

M2NRB033242RB004

242RB006

M2NRB034

242RB005

242RB003

242RB002

242RB001242RB008

242RB009

242RB010242RB011

242RB012242RB013

242RB014242RB015

NSAR

B011

NSAR

B014

NSAR

B008

NSAR

B009

NSAR

B010

NSAR

B015

NSAR

B012

NSAR

B013

NSAR

B003

NSAR

B004

NSAR

B005

NSAR

B006

NSAR

B007

SERB123

SERB124

SERB125

SERB126

SERB002

SERB001

SERB072

FSRB120

SERB067

SERB068

SERB115

SERB075

SERB117

SERB004

SERB114

SERB003

SERB076

SERB116

SERB071

SERB121

SERB118

SERB122

SERB166

SERB047

SERB048

SERB070

SERB046

SERB120

SERB119

FSRB152

FSRB153

FSRB024

NSAR

B002

FSRB065

NSAR

B001

FSRB064

FSRB066

FSRB067FSR

B101

FSRB068

FSRB095

FSRB118

FSRB093

FSRB091

FSRB116

FSRB026

FSRB117

FSRB090

FSRB115

FSRB097

FSRB108

FSRB025

FSRB089

FSRB094

FSRB151

FSRB149

FSRB150

FSRB148

FSRB110

FSRB074

FSRB075

FSRB043

FSRB073

FSRB041

FSRB042

FSRB105

FSRB106

FSRB092

FSRB030

FSRB023

FSRB119

FSRB031

BSRB006

BSRB007

FSRB129

FSRB032

FSRB125

BSRB004

BSRB005

FSRB058

FSRB121

FSRB033

FSRB124

FSRB034

FSRB122

FSRB123

FSRB128

FSRB102

FSRB147 FSR

B144

FSRB062

FSRB044

FSRB109

FSRB069

FSRB146 FSR

B143

FSRB063

FSRB104

FSRB070

FSRB103

FSRB145

WD

RB027

FSRB071

WD

RB028

WD

RB029

FSRB072

WD

RB030

WD

RB031

WD

RB032

WD

RB033

WD

RB044

WD

RB045

FSRB112

FSRB114

FSRB113

FSRB111

FSRB040

FSRB096

FSRB107

FSRB039

FSRB076

WD

RB020

WD

RB021

WD

RB022

WD

RB016

WD

RB017

WD

RB018

WD

RB019

WD

RB026

WD

RB023

WD

RB024

WD

RB025

WD

RB046

WD

RB047

BSRB002

BSRB003

FSRB057

FSRB059

BSRB001

FSRB060

FSRB055

FSRB061

FSRB056

M2N

RB040

M2N

RB041M

2NR

B042 WD

RB001

WD

RB008

WD

RB009

WD

RB010

WD

RB011

WD

RB002

WD

RB003

WD

RB004

WD

RB005

WD

RB012

WD

RB013

WD

RB014

WD

RB015

WD

RB037

WD

RB034

WD

RB035

WD

RB036

WD

RB038

WD

RB043

WD

RB039

WD

RB040

WD

RB041

WD

RB042

WD

RB058

WD

RB006

WD

RB007

WD

RB057

WD

RB059

WD

RB048

WD

RB049

WD

RB050

WD

RB051

WD

RB060

WD

RB061

WD

RB067

DMDD008

DMDD006

DMDD009

DMDD010

DMDD007

SEDD

008

SED

D00

9

MEDD001

MEDD002

FSDD

002

MEDD003

FSD

D00

3

FSD

D00

6

STDD009STDD008

STDD010BSD

D012

BSDD

013

STDD007

STDD006

BSD

D01

5

STDD004

STDD011

BSD

D01

6

STDD005

SED

D00

2SED

D001

419D

D00

3

B SDD

0 14

BSD

D01

9

S ED

D00

6

S ED

D0 0

7

SED

D00

4

SED

D0 0

3SE

DD

005

FSDD

005

FSD

D0 0

4

BSDD009

BSDD

0 04

BS DD

006

B SDD

00 7

BSDD

010

BSD

D01

7

BSD

D01

8

B SDD

0 11

M2NDD008

BSDD

00 8

M2S

DD

001

M2NDD005

M2NDD003

M2NDD006

M2NDD004

M2NDD009

242DD001

M2NDD002

M2NDD001

M2NDD007A

242DD002

242DD003

242DD004

DMDD005

DMDD004

DMDD003

DMDD002

DMDD001

FSDD001

BSDD005

BSDD002

STDD003

BSDD003

419DD002

419DD001

BSDD001

STDD00

2STDD001

M2NDD001

DMRC014

DMRC006

DMRC035

DMRC013

DMRC022

DMRC033

DMRC030

DMRC032DM

RC027

DMRC010

DMRC031

DMRC029

DMRC028

DMRC009

DMRC023

DMRC025

DMRC024

DMRC008

DMRC021

DMRC016

DMRC015

DMRC005

DMRC026

DMRC004

DMRC007

DMRC019

DMRC003

DMRC011

DMRC020

DMRC017

DMRC018

DMRC001

DMRC012

DMRC002

MERC006

MERC005

MERC004

MERC003

MERC002

MERC007

MERC008

MERC009

SERC

024SER

C023

SERC

037

SERC

006SER

C022

SERC

058

SERC

035

SER

C04

2

SERC

021

SER

C04

1

SERC

019

SERC

007

SERC

020

SERC

008SER

C025

S ER

C0 4

3

SERC

009

SERC

026SER

C027

SERC

046

SER

C01

8

SERC

004

SERC

013SER

C017

SERC

056

SERC

060

SER

C05

5SERC

057

SERC

036SER

C034

SERC

016

SERC

054

SERC

012SER

C005

SERC

033SER

C048

SERC

053

SERC

002

SERC

038

SERC

011SER

C010

FSR

C00

9

FSRC

009AF SR

C00

9B

SERC

039

SERC

028

SERC

047

SERC

045

FSRC

020

FSRC

006 FSR

C01

0

SERC

040FSR

C011

SER

C04

4

BSRC

003

BSRC

002

BS RC

001

S ERC

03 1

SERC

032

SERC

015SER

C05

0

SERC

003

SERC

030

SER

C01

4

SERC

052

SER

C04

9

SERC

029

SERC

001

S ER

C05

1

419RC

022

419RC

012

419RC

009

419RC

010

4 19 RC

0 11

FSRC

005

419RC

002419R

C013

419RC

014

419RC

003

419RC

021

4 19R

C00

7

419R

C00

841

9RC

0 06

419R

C0 1

74 1

9RC

005

4 19R

C01

8419R

C01

9

419RC

001

419RC

015

419RC

004419RC

016

4 19R

C02

0

FSRC

001

MERC001

FSRC

012

SERC

059

FSRC

004

FSRC

019

FSRC

018

FSRC

021

FSRC

017

FSRC

016

FSRC

008FSR

C003

FSRC

002FSR

C007

FSRC

013FSR

C014

STRC011BSR

C00

4

BSRC

014

STRC012

STRC005

FSRC

015

STRC013

BSRC

012

STRC014

STRC015

STRC010

STRC009

STRC006

STRC008

STRC007

BSRC

011

BSRC

010BSR

C009

M2NRC008

M2NRC007

M2NRC010

STRC004

STRC003

M2NRC009

M2NRC006M2NRC003

M2NRC002

M2NRC012

M2NRC001

M2NRC011

M2NRC005

BSRC

005

242RC015242RC005

242RC011

242R

C014

242RC006

242RC004

M2NRC004242RC019

242RC012

STRC002

STRC001

242RC013

242RC017

242RC016

242RC018

242RC020

242RC010

242RC021

242RC002

242RC023

242RC003

242RC001

242RC008

BSR

C01

6

BSRC

015

BSRC

019

BSRC

007

BSRC

007ABSRC

008

BSRC

006

BSRC

018

BSRC

017

BS RC

01 3

242RC007

242RC009

242RC022

381

3

380

435 473

378

300

396

377453

376

398

399

400

371371B

311

313

293

338

337

283

DMH1

286

611

630B

620B

618B

544

553

543B

695

467

471

477

576

536

545

256

100

503

506

279C

660

697

665

103B

474

640

541

355

369

354B354

368

346B

353

352

343B

359

360B

360C

185

205

237

233

232

234

902

465

315

154

3 20

044B44

900

265

319

221

260157157B

61

261

328314

602

899

055B

224

203

255

151

50050B

152B

201201B

327

254326

25

253226B

251

252

161

505

222

270

202

330

269B

176C

240

227

714

243

446

726

445B

241

792

740

731

739B

739

742

732

DMDMH

DMH27

DMH32

725

383

DMH34

395

DMH30

310

DMH31

DMH29

394

302

308

384

393

DMH75

DMH76

DMH28

DMH33

379379B

301

DMH26

DMH37

DMH21

DMH22

DMH24

331DMH25DMH19

297

DMH20

299

DMH23

DMH35752

DM

H14

DM

H15

309

DM

H13

751

DM

H12

397

433433B

434

432

295

DMH17

DMH18

296

298

340

294

DMH16347

292

331B

331C

307

385

DM

H11

306

304

DM

H10

386

303

391

305

333

387

DM

H9

280

388

760

756

621

810

348

291

332

289

349

281

290

312

3503 36

284

339373

375

285

DMH2

372372B

DMH55DM

H3

287

282

334

DM

H8

335

757

DM

H4

DM

H5

288

DM

H5 7

DM

H6

822

816

815

813

814

896

812

620

654

617

SE7564

SE9

554

SE12

765

SE8

619SE5SE6

575

761

737

565

555

566

724

719556

556B

738

557

558

546891

547

548

831

731

DMH5

6

1014

ME11

ME12

ME1

820

821

897

767

895

785

786

766

819

770

771

773817

818

769

774

600

642

758

811

598641

74175975

9B

599

603917

976

977978

573

572SE10

571

694

SE4

SE2

SE3

692

918

919

94

535

ME2

537

ME4

589

ME3

930

890

588

ME6

585

590

ME5

ME8

ME7

829

ME10

677

678

ME9

755

753

634

635

638

629

630

935639942

943

631

628

898

936

451

450

452

SE26

754

439 851

605

561

562

563SE14

SE18

604

SE59552

551

SE19

SE17542

SE13

SE61

939SE60

937

938940

941

658

651

494

543

SE20

SE21

356

SE15

10011012

357

1002 365SE22

1013

358

1003

364

SE16

SE231004

1005SE56

SE54SE55

SE57

698

995

664

996

669

846

993

663

743

994

847

696

848

848B

101

709

710

1020

1021

744

1019

SE63

SE64

SE62

649

874

958

SE39

SE40

957SE37

SE43

SE42

SE41956

1031

495

SE38

1006

1007

346

SE27

SE24

SE25

1011

SE28

343

1010

SE53

SE29

SE50

875

345

SE33

SE36

SE32SE34

367

SE31366

SE35

SE30

SE48

SE49

SE51

SE47

SE52

SE46

998

999997

SE44SE45SE65

SE66

850

965

966

967

984

862

861

863

982

983985

864865

661

655

9641034

1032

1033

448

472

SE11

475

476

466

1009

363

1008

461

462

478

362

470463

459

469

464

361

93

085B

699

992

102

711

713

103

1022

344

1023

1025

342

1026

1024

223

143

193

129

963955

496

962954

500

784

852

783

946953

187

498

1030

341

499

1029

497

610

191

192

189

190

959

960

961

842

419-6

10171018

849

1016

424

781

1037

779

419-5780

422

10351035B

1036

454

458

457

455456

460

468

084B

092B

091B

095B

090B

083B

080B

089B

188

857

135

104

969

137

182

970

856

972

971

128

105

856B

438991989

990419-1

841843

845

840844

839

427

429B

142

136

84 945

92

316 91

88

156

968974

134

317

87

155

264

95

31889

90

266

263

80

83

144

077B

419-4

449

411

428987

988

419-3

425

419

986

419-2

429

420426

437

436

418

947952

419-7782

8531027948

1028951

949950

186

184

854

138

238

858

183

196

239

235

236

145

916973

181

139

855

126

106

127

180

195

79

825826

198199

82

197

158

159

97

278

76

110 606

109

271

112

115

116

269

860 122

859

124

96

108

194

160

609

121

107

123

119

120

117117B

118

272

272B

204

410

417

409

416

408

415

407

406

866

867

230

728

277

231231B

870

868

869

871

872

873

229

228

241B

876

114B114

175

146275

174

113

832

276242

441

443

442

444

834

717717B

55

29

34

52

736

54

56

51

550

540

747

892

746

894915

827788 909

772

893

913

914

768

775

878

879

880

38

601

480

745

30979

1015

35

980

981

28

64

71

883

884

70

26

889

679

828 830908

053A

735

929

680

888

718

257

053B

927

928885

886

887

787

482

479

31

36

071B

32

481

59153

58

33

152

57

37

86

920

921

922

85

502

923

081B

81

944

4075

069B 69

43

42

975

4174

78

171

27

49

48

6

910

911

54

45

7

3

208

173

050A

903

131

926

904

905

925

906

907

912

1321038

130

1039

149

279279B

877

2

150

1

63

258

225

24

259

2268

23

39

22

921

172

1817

16

19

164

20

163

165

162

68

078B

262

67

321

73

901

77

66

72

924

608

98

65

99

024B

018B

62

023B

12

140

15

11

14

10

60

13

141

1000

329

169

111

267

170

147

168

176176B

177268

273

324

933274

881

166

931

148

167

504

932

46

178

250

133

125

179

47

351

789

734

325

245

244

716

882

934

323

715

322

440837

247

838

445

833

447

419 ZONE

Main 2 North

Starter Pit

21000

Kubekro River

Deadman's Hill

Mid East

Power

Plant

Main 2 South

Main 1

Camp1

PW Yard

Administration and

Warehouse

ROM Pad

Kubekro Road

B-Shoot

Agglomerator

F-Shoot

345/80

080/69

071/64

Scale: 1:125000

Drawing:

Projection: Ghana National Metric (Clarke 1880/War Office 1924)

Date:10/2/2003

Author: J. Amekudi

Office: Wassa

LithologyPHYGRAFPOFELDIOBMUBAS

Site Excavation Areas

Pit Excavation Areas

RoadsAccess

WASSA GOLDMINES

WEXFORD GOLDFIELDS LTD.

FIG 3.1

ReWASSA

LOCAL GEOLOGYDRILL HOLES LOCATION




Documents

Technical Report for the First Disclosure of a …s1.q4cdn.com/789791377/files/doc_downloads/Wassa/wassa...Oxide Ore 3.38 1.16 Fresh Ore 6.91 1.69 Total Ore 10.29 1.51 Total Waste