(end of period)

Different mining equipment is used in the saprolite and bedrock due to the significantly

different material characteristics. A contractor will mine the saprolite using a fleet of allwheel-

drive trucks and excavators. Crystallex will mine the bedrock using conventional

136 tonne haul trucks and 21m

Several samples of saprolite oxide (SAPO), saprolite sulphide (SAPS), carbonate

leached bedrock (CLB) and carbonate stable bedrock (CSB) ore from the Conductora pit

were examined in bench tests and pilot plant operations by SGS Lakefield Research

(Lakefield) during the months of April through December 2003. Samples of waste from

the Conductora pit and four samples of Mesones ore were also studied. Sub-samples of

Conductora ore were sent to McGill University for gravity recovery testwork. Outokumpu

conducted pilot plant settling tests on several samples. The various test programs were

designed to confirm relevant data generated by Placer, determine the gold recovery and

reagent requirements for the proposed gravity-leach flowsheet, and generate plant

design data.

Grinding data are generally in accordance with data generated by Placer Dome. Pilot

scale gravity concentration tests at Lakefield on Conductora ore show about 30% gold

recovery from both a SAPO-CSB blend and a SAPO-SAPS-CLB-CSB blend at mass

concentration ratios of about 4000:1. Preliminary data for Mesones shows an even

better response. Intensive cyanidation of the concentrates from Conductora gave >99%

leach recovery. Tests at McGill to determine the gravity recoverable gold (GRG) content

of Conductora SAPO and CSB samples showed 39% and 46% GRG, respectively which

would translate into practical recoveries of about 25%.

Thirty-six hour bottle roll leach tests on Conductora gravity tailings confirm that SAPO

leaches very well to give about 99% overall (gravity+leaching) extraction and a 0.02 g/t

tailing. With a 24 h leach time, tailings were 0.03 g/t corresponding to 98% extraction.

CSB gives about 85% overall extraction (0.17 g/t tailing). Cyanide additions for SAPO

and CSB have been less than 1 kg/t ore. Pure SAPS samples with cyanide soluble

copper (CNSCu) levels of 370 ppm or less have been tested and gave 85 to 88%

extraction, albeit with cyanide additions of 1.7 to 1.9 kg/t. Mixtures containing SAPO,

SAPS and CSB gave 85 to 90% overall extraction provided that sufficient NaCN was

present. The NaCN addition varied with the CNSCu level in the ore.

An initial gravity-leach test on each of the four Mesones samples showed an average

85% overall gold extraction and modest reagent consumption. It is believed that higher

extraction could be obtained with optimization of the leach conditions.

Duplicate bench scale tests on a series of samples containing 20%CLB and 80% CSB

and between 1 and 2 g/t gold yielded an average of 88.7% overall gold recovery (gravity

and leaching) with no measurable dependency on head grade.

A 2 kg/h pilot plant was operated for three weeks in which batch-ground/gravity

concentrated Conductora ore was subjected to carbon-in-leach (CIL) processing. During

the first 13 days (PP1), a blend of 20% SAPO and 80% CSB was leached with 0.7 kg/t

of cyanide to give a final overall gold extraction of 89.6% (tailings average of 0.15 g/t). A

SAPO-SAPS-CLB-CSB blend was processed for the last week (PP2). The plant tailing

was 0.15 g/t for an extraction of 89.3% with a cyanide addition of 0.8 kg/t.

Viscosity measurements by Lakefield indicated nothing problematical in the mixtures that

will be handled in the Las Cristinas plant.

Outokumpu conducted high-rate thickening tests on nine sample blends, ranging from

pure SAPO to pure bedrock, using its pilot-scale thickener. At 50% solids in the

underflow, all blends containing 50% SAPO or less could be processed at 0.46 t/m2/h or

greater. Allowing for a 15% scale-up, the data showed that a 50 m diameter thickener

would give at least 47% solids in the underflow when processing up to 20 000 t/d of a

50% SAPO, 50% CSB mixture. Acid-base-accounting (ABA) tests and various

geotechnical studies were performed by Lakefield on several samples to determine the

potential for acid generation. Data are discussed in Section 14 of this study report.

Natural degradation tests and continuous INCO Air/SO2 cyanide destruction tests have

been performed on pilot plant tailings. Natural degradation under Lakefield climatic

conditions reduced CNWAD to below 20 ppm in about 40 d for pilot plant tailings from

PP1 and 100 d for PP2 tailings. The INCO process then reduced CNWAD to <0.3 ppm

and Cu to about 1 ppm under industry-typical operating conditions. INCO tests on

naturally-degraded PP2 tailings solution gave <0.1 ppm CNWAD and <0.5 ppm Cu.

All equipment, with the exception of secure areas such as gravity circuits and refinery,

electrical and control rooms, is located in open sided structures or outdoors. The

processing plant is fenced for gold security reasons. Installed spare pumps are provided

for all critical process streams.

The process plant consists of single line crushing, semi-autogenous primary grinding

(SAG) followed by secondary grinding using a ball mill. A pebble crusher is incorporated

in closed circuit with the SAG mill.

A gravity circuit is included in closed circuit with the cyclones in order to recover any

coarse, free gold prior to regrinding in the ball mill.

Gold extraction is achieved in a conventional carbon-in-leach (CIL) circuit. Gold is

removed from the loaded carbon by pressure stripping, electrowinning and smelting a

gold dore product.

CLB and CSB ore is delivered by mine truck to the double dump primary crushing station

which is permanently located to the east of the process plant. The primary crusher

product discharges via an apron feeder on the stockpile feed conveyor.

CLB and CSB ore is reclaimed from the coarse ore stockpile using apron feeders

located in the reclaim tunnel situated below the stockpile. The ore is loaded onto the

SAG mill feed conveyor.

SAPO and SAPS ore is delivered by mine truck to the double dump saprolite crushing

station. The mine trucks direct dump into a feed hopper which is positioned over top of

an apron feeder. The apron feeder passes the saprolite ore into a mineral sizer in order

to reduce over sized clumps before being fed on to the SAG mill feed conveyor.

The SAG mill feed conveyor delivers a combination of saprolite, CLB and CSB ore

directly to the SAG mill. The SAG mill is driven by a wrap around, variable speed motor

with a cycloconverter. The SAG mill discharge is screened by two double deck vibrating

screens to remove over sized 12 mm pebbles. The 12 mm pebbles from the vibrating

screens are crushed in a cone crusher prior to being recycled back to the SAG mill feed

chute. Provision has also been made so the pebbles can be recycled directly back to

the SAG mill without further size reduction or can be stockpiled outside the process plant

building.

The under sized product from the vibrating screens drops into the cyclone feed pump

boxes where it is combined with the discharge from the parallel ball mills. The ball mills

are driven by a wrap around, variable speed motors through a cycloconverter drive units.

The combined SAG and ball mill discharges are diluted in the parallel grinding circuit

pump boxes with process water and pumped to dedicated cyclone clusters which sorts

the ore particles by size and returns the over size to the ball mills for further size

reduction.

Also included in each parallel grinding circuit is a gravity recovery circuit. A portion of

each ball mill discharge is diverted over a vibrating screen with the under size fed to one

of two centrifugal concentrators. Gravity concentrate from each centrifugal concentrator

is stored in the same secured holding cone until it is leached in a semi-batch, high

intensity cyanide leach reactor. Gravity and leach reactor tailings are pumped backed to

each parallel grinding circuit. The gold loaded solution from the leach reactor is pumped

to a dedicated electrowinning circuit located in the secured gold room.

Slurry from each parallel grinding circuit cylcone overflow, after trash removal, is gravity

fed to one of two thickener feed collection boxes where slurry flows into a 50 m diameter

thickener. The two thickener overflows flow by gravity into the process water tank. Each

thickener underflow is pumped at 50% solids by weight into dual parallel 12-stage CIL

circuits. Cyanide and lime are staged added to each tank train. On an intermittent

basis, loaded carbon is pumped counter current to the slurry flow in order to increase

the gold loading. Loaded carbon is removed from the head end of each tank train and is

transferred to the acid wash vessels via a dedicated loaded carbon vibrating screen.

Loaded carbon from the two parallel CIL circuits captured by the vibrating screens drops

by gravity into the acid wash vessels. Each acid wash vessel is part of a dedicated,

parallel elution and carbon regeneration circuit consisting of duplicate elution columns,

electric immersion heater, heat exchangers, barren eluate tanks and kilns. The

description to follow refers to either parallel elution and carbon regeneration circuit.

A 3% acid solution is pumped into the acid wash vessel and overflows the top and

returns to the acid mix tank. Acid washing takes approximately 1 hour. After acid

washing is complete, the spent acid is neutralized with sodium hydroxide before

discarding it to the tails pump box.

The desorption elution cycle starts with the preparation of a 3% sodium cyanide and 2%

sodium hydroxide solution in the barren eluate tank. The solution is initially pumped

through the strip solution heater and returns to the barren eluate tank until its

temperature reaches 80

carbon fines tank. Water collected in the carbon fines tank is pumped through a plate

and frame filter press to capture any carbon fines.

The activity of the stripped carbon is restored in a kiln. After passing through the kiln,

the carbon drops out into a quench tank and is transported to the reactivated / fresh

carbon sizing screen.

Screened carbon from either parallel circuit drops by gravity to same the reactivated

carbon transfer hopper where it is mixed with washed fresh carbon. Screen fines flow by

gravity to the carbon fines tank. Water collected in the carbon fines tank is pumped

through a plate and frame filter press to capture any carbon fines.

There is some carbon loss through attrition and is made up with fresh carbon. Mixed

regenerated/fresh carbon in the transfer hopper is moved to the last leach tank in each

CIL train via a horizontal recessed impeller pump.

Pregnant eluate solution from the desorption circuits reports to the pregnant eluate tank.

Pregnant eluate solution is pumped to six electrowinning cells (three rows of two in

parallel). Gold metal is electrowon loosely on the stainless steel wool cathodes in the

electrowinning cells. Depleted solution flows from the outlet of each cell to the barren

eluate return tank and is then transported either back to the barren eluate tank or

recirculated back through the electrowinning cells via the pregnant eluate tank.

Pregnant eluate from the concentrate leach circuit is pumped to the leach reactor

pregnant eluate tank in the refinery area. Pregnant eluate solution is transported from

the tank to two electrowinning cells in series. Gold metal is electrowon loosely on the

stainless steel wool cathodes in the electrowinning cells. Depleted solution flows from

the outlet of the last cell to the leach reactor barren eluate return tank and is then

transported either to the CIL circuit or recirculated back through the electrowinning cells

via the leach reactor pregnant eluate tank.

At the end of the run, the cathodes are removed from the cells; the gold bearing sludge

is washed off and then pumped to a plate and frame filter press. The filter cake is mixed

with fluxes, usually borax, soda ash and occasionally sodium nitrate and fed to an

electric induction furnace. The doré metal and slag separate in the furnace, and the slag

is poured off to slag pots then the doré metal is poured into bars for shipment.

The cyanide destruction process is air/SO

The Las Cristinas site is situated in south eastern Venezuela and is some 6 km west of

the village of Las Claritas on Troncal 10 the main highway running from the Brazilian

border to the Venezuelan port of Puerto Ordaz on the Orinoco River. The site is some

360 km by road from Puerto Ordaz and the road presents no significant obstacles to the

transportation of goods and materials to the site.

Access to the site will be from Troncal 10 along an existing unpaved road that will be

upgraded to take construction and operational traffic. This route is 19 km long and being

north of Las Claritas circumnavigates all the local villages and will thus avoid any

disruptions to the local population.

An existing 400 kV power line parallels Troncal 10 and a new substation was

constructed in 2001 just south of Las Claritas at Km 86 to service the area. The

substation has two 150 MVA power transformers and provision has been built in to

supply Las Cristinas with a 230 kV power line.

The site power demand is estimated at 60 MW which can be adequately supplied by the

substation.

Power to the site will be carried via a new overhead power line, a distance of

approximately 6 km, and will terminate at a new substation to be built on site from where

power will be distributed at 6.6 kV.

Potable water will be drawn from on-site wells and will be chlorinated prior to distribution

for consumption. Make-up water for process requirements will be drawn from the Potaso

Pit, an old mining pit that is permanently flooded. During operations the Potaso pit will

be charged with water from the diversion ditch.

Domestic sewage will be collected by a system of gravity sewers and treated biologically

with the resulting effluent being pumped to the tailings pond.

In 1998 a 3,058 person construction camp was constructed at the Las Cristinas property.

The camp included dormitories for workers and supervisors, kitchen and canteen

facilities, administration building, water and firewater plant and a sewage treatment

plant. The camp was subsequently abandoned and has been subject to neglect and

minor vandalism.

For the current project the construction camp will not be utilized except that the

administration building will be refurbished and will serve as the main administration

centre, the kitchen and canteen will be converted to a construction and operations

warehouse and the existing water plant will be brought on line. Sewage from the camp

site will be redirected to the new sewage treatment plant.

Aside from the main administration facilities located in the existing construction camp

additional buildings will include a Guard House at the entrance to the process plant area,

a Mill Administration and Dry, a Truck Maintenance and Mine Dry and a Truck Wash.

Tailings area water management forms a large component of overall site water

management. Therefore the tailings area water balance was developed in combination

with the overall site water balance to support the development of the site water

management scheme. The water balance provides an indication of average process

water flow rates, range of tailings pond operating volumes, average treatment rates for

water treatment plants, average pumping rates from water management ponds and

average discharge rates of excess water to the environment.

The site water management scheme has been developed so that pumping and

treatment costs are minimized by isolating clean runoff from potentially contaminated

runoff and process water streams. Environmental impact is reduced by providing

appropriate containment and treatment to all potentially contaminated site water before

discharge, and by maximizing the use of water recycling.

Six site water management ponds are proposed in addition to the tailings pond. All

runoff from waste rock dumps and saprolite waste dumps will be collected in ponds to

provide settling of suspended solids. All runoff from waste rock dumps will be monitored

for acid drainage.

Process reclaim water from the Tailings Management Facility (TMF) water reclaim

system will pass through a cyanide destruction facility before use in the process plant.

Freshwater makeup will be supplied by pumping from the Potaso Pit. This water will

require treatment in a sedimentation/filtration plant before entering the process stream.

Any seepage from the TMF dike will be collected by a perimeter ditch and pumped back.

Excess TMF pond water will be considered suitable for discharge to the environment

following cyanide removal if suspended solids are within an acceptable range.

Clean surface water from upstream drainage areas will be collected into a diversion

channel and conveyed around the perimeter of the site. Clean surface runoff from

undisturbed drainage areas within the mine site will be collected and diverted to the

Potaso Pit which overflows into the river diversion system. Site drainage was designed

for a 1:25 year flood event and the river diversions for 1:100 year events.

A field program for the Tailings Pond area was undertaken by Bruce Geotechnical

Consultants Inc. (BGC), in 1994 and 1995 and reported in the Las Cristinas Feasibility

Study (BGC, 1996). BGC drilled 9 boreholes, excavated 27 test pits and carried out

geologic mapping of outcrops. The geologic horizons were described as follows. The

upper horizon consists of a thin laterite soil horizon from 0.5 to 1.0 m thick. The next two

units are saprolite which will form the foundation immediately beneath the tailings dikes.

The upper layer of saprolite oxide (SAPO) is from 0 to 40 m in thickness, while the

thickness of the underlying layer of sulphide stable saprolite (SAPS) varies from 0 m to

65 m. Below the saprolite is a layer of saprock, generally less than 1 to 2 m thick.

Beneath the saprock, bedrock is subdivided into CLB (carbonate leached bedrock) and

CSB (carbonate stable bedrock).

The results of BGC’s geotechnical investigation were used by SNC-Lavalin for the

feasibility level design of the Tailings Management Facility (TMF). In addition, SNCLavalin

carried out analysis of samples collected from the sand and gravel deposit in the

tailings area. Results show that the sand and gravel is suitable material for filter,

drainage and other granular usages.

Design criteria for the TMF were selected to optimize groundwater protection, physical

stability and mine closure conditions, and to make maximum use of mine waste

materials on a cost effective basis. Due to the presence of cyanide in the tailings slurry,

the TMF was designed to withstand a maximum credible earthquake and to contain the

runoff from a 24 hour probable maximum flood event, based on internationally and

nationally accepted practice and risk ratings.

A cyanide destruction plant will be built to treat the Weak Acid Dissociable cyanide

concentration in the water discharged with the tailings slurry. It is not expected that

treatment of an acidic runoff will be required during mine operations.

Previous studies by BGC (BGC, 1996) identified a tailings facility comprised of two cells.

A two celled facility is no longer required due to changes in mining plan and therefore, a

single celled facility is proposed. In addition, the proposed dike alignment differs from

that proposed by BGC. The tailings dike does not extend as far south as the previous

layout since there is an area of sand and gravel deposit along the old south dike

perimeter. This material is not suitable as a foundation material and therefore, the

alignment was adjusted.

The alignment of the tailings dike was selected to provide a natural low permeability

foundation, to provide sufficient storage for tailings and water management, and to utilize

available natural topographic conditions.

The starter dike will form the first stage before operations begin and subsequent stages

will be constructed during operations. The starter dike will be sized to provide tailings

storage and water management for the first three years of operation. It will be of low

permeability design with foundation preparation and seepage control measures for

adequate structural and hydraulic stability. The TMF basin floor is saprolite 20 m to 40

m thick with permeabilities ranging from 8x10

Seepage analyses were carried out to estimate the seepage through the tailings dike for

the purpose of sizing the chimney and finger drains as well as the perimeter collection

ditch. Stability analyses were carried out using parameter values based on analyses

carried out by BGC as reported in the Feasibility Report (BGC, 1996). The dike

structure is stable under various loading conditions and suitable for post-closure

environment.

The minimum and maximum normal operating volumes used for design of the tailings

pond are 500,000 m

For the 20,000 t/d feasibility study Buckland Harapiak (B-H) was engaged by Crystallex

to carry out a study and make recommendations for the appropriate organization

structure for the Las Cristinas operation, with a particular focus on the Finance &

Administrative functions. SNC-Lavalin has revised the B-H work to account for the higher

number of employees that will work on the 40,000 t/d operation. The organization

proposed would support the 40,000 t/d open pit mine and CIL processing facility with a

total peak work force of approximately 450 employees. Research for this report included

interviews with senior management from Crystallex, including in-country management;

feasibility work previously completed by SNC-Lavalin and MDA; the 1996 Socio

Economical Study conducted on behalf of PDI; the Las Cristinas Development Plan

(presented earlier this year by Dr. Sadek El-Alfy and Julio Rojo) and various other

sources of information on Venezuela and comparable mining operations around the

world.

A number of conclusions and recommendations can be drawn from environmental

analyses and the preliminary assessment of the potential environmental impacts

conducted during the feasibility stage of the Las Cristinas Project, as well as

development of the preliminary concept for site closure and rehabilitation.

A detailed Environmental Impact Study (EIS) of the Las Cristinas project is required by

both Venezuelan legislation and World Bank standards/guidelines. A significant amount

of environmental baseline data and impact analysis necessary for EIS preparation was

undertaken by PDI throughout the early to mid 1990’s. PDI submitted an EIS document

to the Venezuelan Government for review and approval in 1996. A Land Occupation

Permit was issued for the project in July 1997, and a Permit to Impact Renewable

Natural Resources was issued in August 1997. PDI withdrew from the project in July

2001.

Building on the work conducted by PDI, and in consultation with the Venezuelan Ministry

of Environment and Natural Resources (MARN), Crystallex recently developed an

update of the approved EIS to reflect changes in project design and environmental

characteristics, which will be submitted to MARN in the near future. The main

environmental activities conducted during this feasibility stage were:

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The following are key conclusions of the feasibility stage environmental impact

assessment and preliminary site closure and rehabilitation concept for the Las Cristinas

project:

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low-permeability saprolite soils (to retard downward migration into subsurface

soils and ground water) and covered/buffered by non-acid generating or net acid

consuming waste.

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examining alternate employment opportunities for other small-scale miners in the

next stage of project design.

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A number of recommendations for specific work tasks to be conducted in subsequent

design stages are considered routine (such as preparation of a detailed environmental

impact assessment), and are not provided in the following summary. Crystallex will

develop a work plan that contains these activities prior to initiating the next stage of

design. The recommendations provided below are those that are considered most

significant in the consideration of project feasibility.

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The Las Cristinas estimated capital costs are summarized in the Table 1-7.

Mine 11,777

Process Plant 142,074

Tailings Management Facility 62,904

Infrastructure 29,289

SUB-TOTAL DIRECT COSTS 246,044

Owner’s Costs 15,000

Indirect Costs 104,356

TOTAL COSTS 365,400

In addition, sustaining capital totalling $169.5 million over the 20 years of the mine will

be required.

These estimates do not include VAT of which is recoverable once gold sales commence.

The estimated Operating Costs for the project, based on life of project averages are in

Table 1-8 as follows (before royalties).

Mining $2.531 $75.93

Processing $3.206 $96.17

G & A $0.227 $6.80

Note: *Does not include royalties

The findings in Table 1-9 were estimated by the financial analysis using a gold price of

$325/oz.

Capital Cost (excluding financing and

inflation) 2003 US$ before start of

operation

US$ 368,670,000

IRR Project (without debt financing) 17.3% 16.4% 11.8%

Payback Period-Years 4 6