Overview

Deposit type

• Magmatic
• MMS
• Intrusion related

Summary:

Deposit Types
The Wellgreen Deposit is hosted in the Quill Creek Complex, one of a number of mafic-ultramafic sills that are enriched in nickel-copper-PGM mineralization that outcrop within the Kluane Ultramafic Belt of the Wrangellia Terrane in southwestern Yukon.

The Project is located within the Insular Superterrane, which is dominantly composed of two older terranes, the Wrangellia Terrane and Alexander Terrane, that were amalgamated at approximately 320 million years (Ma). These terranes are comprised of island arc and ocean floor volcanic rocks overlain by thick assemblages of oceanic sedimentary rocks that range in age from 220 to 400 Ma. The Project is part of the Kluane Ultramafic Belt, situated in the southwest portion of the Wrangellia Terrane that spans from Vancouver Island, north through British Columbia (BC), into Alaska.

The Wellgreen Deposit occurs within, and along, the lower margin of an Upper Triassic (Kluane) ultramafic-mafic body, within the Quill Creek Complex. This assemblage of mafic-ultramafic rocks is 20 km long and closely intrudes along the contact between the Station Creek and Hasen Creek formations. The main mass of the Quill Creek Complex, the Wellgreen Deposit, and Quill intrusions, is 4.7 km long and up to 1 km wide.

Mineralization
Mineralization on the Project occurs dominantly within the Quill Creek Complex except for a small portion at the contact within the metasedimentary host rocks. This serpentinized, ultramafic-gabbroic body intrudes the Pennsylvanian-Permian sedimentary and volcanic rocks of the Station Creek and Hasen Creek formations. The main zone of mineralization has a strike length of 1.7 km and thickness ranges from 20 m on the western end to almost 300 m at the eastern end. Drilling intercepts have indicated the mineralization ranges in depth from several metres at the west of the deposit to over 500 m at the eastern side. Discontinuous massive and semi-massive sulphide zones are significantly thinner (centimeters to a few metres), are located near the footwall contact and transition into disseminated sulfide zones above. Historic exploration and development programs defined two main zones of gabbro-hosted massive and disseminated sulphide mineralization known as the East Zone and West Zone. These zones have since been determined to be contiguous and have been further broken up and now are known as the Far East, East, West, and Far West Zones with the connecting Central Zone. The North Arm Zone is interpreted to be a splay off of the Far East Zone. All intrusive phases containing primary olivine have been moderately to intensely serpentinized.

Far East Zone
The Far East Zone represents the easternmost part of the Wellgreen Deposit intrusion. The zone lies between 578250E and Aird Creek, at approximately 578750E. In both the current East and Far East Zones, historic exploration efforts focused on defining massive sulfide horizons and lenses near the contact between the Project Intrusion and Hasen Creek metasedimentary rocks and as such this contact is very well defined. This sedimentary contact was historically interpreted to be the steeply dipping southern footwall to mineralization based on the data available at the time, but more recent work in the East Zone showed the sedimentary contact was a wedge of metasedimentary rocks in a much larger ultramafic body. This change in understanding the nature of the sedimentary contact was demonstrated in the Far East Zone by drill holes 154, 160, and 165.

The typical steeply dipping lithological sequence of peridotite, clinopyroxenite, and gabbro with massive sulphide is very well defined in the Far East Zone. The core of the Far East Zone shows a sulphide-rich, clinopyroxenite, and gabbro/skarn horizon with a second clinopyroxenite and gabbro enriched zone at the lower contact with the metasediments.

In the easternmost portion of the Far East Zone, all lithologies exhibit a similar sub-horizontal dip to the symmetrical sequence further west: peridotite transitioning to clinopyroxenite, and gabbro with skarn units and massive sulphide immediately prior to the basal contact with Station Creek volcaniclastics and Hasen Creek metasedimentary rocks. This lower sequence is interpreted to be contiguous with the basal sequence observed 350 m farther to the west. In addition, the foot-wedge pinches out to the east such that in the upper portion of the intrusion, the various contact-proximal lithologies are absent.

East Zone
The East Zone lies between 577900E and 578250E and was historically explored for massive sulphide at the footwall contact. As mentioned above, this zone was the first in which the change in the footwall contacts’ orientation was observed in drill core. The peridotite-clinopyroxenite-gabbro sequence is observed to wrap around the base of the wedge in the East Zone.

Central Zone
The Central Zone lies between 577500E and 577900E. The eastern portion of the zone is similar to the East Zone whereby well-mineralized peridotite gradationally transitions to clinopyroxenite and gabbro, and units are observed near the contact with dominantly Hasen Creek metasedimentary rocks. The western portion of the Central Zone exhibits a sub-horizontal, symmetrical, mineralized unit similar to that intersected at depth in the Far East Zone. Additional drilling will be required to test whether the higher-grade, sub-horizontal, mineralization intersected in the Central Zone connects with that in the East and Far East zones. This represents a high priority exploration target, and currently is the least drilled zone on the Project.

West Zone
The West Zone lies between 577120E and 577500E. Similar to the western portion of the Central Zone, well-mineralized peridotite overlies a comparatively thick package of clinopyroxenite and gabbro with significant semi-massive and massive sulphide zones.

Far West Zone
The Far West Zone lies between 576720E and 577120E, and the northern part of the zone is interpreted to be a branching sill from the main Project intrusion. This sill is generally zoned outwards, well mineralized in the centre, grading from peridotite to clinopyroxenite and gabbro towards the contact with the metasedimentary country rocks. Grades in the Far West Zone are significantly elevated starting at surface with high sulphide content.

North Arm Zone
The North Arm Zone is located in the east-central portion of a narrow 1,200 m long sill, positioned approximately 150 m below the main Project intrusion. It was discovered by Hudson-Yukon Mining in the 1950s and explored in 1987 with three drill holes by All-North. All of these drill holes intersected mineralization. The geology of this zone is similar to both the East and West Zones. Mineralization consists of massive sulphide lenses, disseminated sulphide in gabbro and clinopyroxenite, and fracture fillings in footwall Hasen Creek metasedimentary rocks. The North Arm Zone was tested in 1988 and 2005 by limited drilling and was determined to have a sub-vertical dip. The information collected todate suggests the North Arm Zone is relatively narrow in comparison with the main Project body at surface, but it does represent a prospective area of nickel-copper mineralization that warrants further work and may be contiguous with the main Project intrusion at depth towards the eastern end of the deposit.

Mining Methods

• Truck & Shovel / Loader

Summary:

The Mineral Resources for the Project include the Wellgreen and Arch deposits, however the Arch deposit was only recently explored and will not be included in the Mineral Reserves. AGP’s opinion is that with current metal pricing levels and knowledge of the mineralization and previous mining activities, open pit mining offers the most reasonable approach for development of the deposit. This is based on the size of the resource, tenor of the grade, grade distribution and proximity to topography for the deposits.

The current mine life includes two years of pre-stripping followed by nineteen years of mining. The open pit mining starts in Year -2 and continues uninterrupted until early in Year 19.

Pit Design
The pit design consists of 3 main phases of successive pushbacks. Phase 1 provides the initial low strip ratio mill feed in the schedule. A waste quarry has been designated as phase 1B and is immediately southwest of the phase 1 pit. The quarry material will be mined early in the schedule so that the crusher and stockpile pad can be constructed as soon as possible. These initial phases will be followed by phases 2 and 3 which both extend to the western and higher main portions of the pit. The pit optimization shells used to guide the ultimate pits were also used to outline areas of higher value for targeted early mining and phase development. All pits were developed using 10 metre bench heights.

Phase 1B
Phase 1B is the first mined in the pit. It has been designed as a nearby source of waste for the construction of the crusher and stockpile pad. This phase is mined from 1505 masl down to 1285 masl. Narrow waste benches near the top of the phase will be mined with dozers and pushed down to lower benches. Pioneering access will be shared with phase 1 for some of the upper benches.

Phase 1
Phase 1 is the first significant source of ore mined in the pit. It will be mined concurrently with phase 1B and will release ore as stockpile capacity increases with pad construction. This phase is mined from 1515 masl down to 1245 masl. Pioneering access will be shared with phase 1B for some of the upper benches.

Phase 2
Phase 2 extends the mine to higher elevations and also to the western portion of the deposit. Once the pioneering road is complete to the upper elevations, the phase retreats back down the mountain and leaves haul road access in its wall for phase 3. This phase is mined from 1925 masl down to 1225 masl.

Phase 3
Phase 3 is the final phase to be mined in the deposit. The haul road access from phase 2 is mined to final wall and the final access to upper elevations is via haulage ramps on the northeast waste rock facilities. This phase is mined from 1925 masl down to 1125 masl. As the ultimate pit has a very high north wall, a wider berm of approximately 26 m width was used at 1475 m elevation. This berm can then be maintained to allow for safer working conditions at lower elevations.

Rock Storage Facilities
The design of the rock waste dumps used a swell factor of 1.30 and were designed with a 37° face slope. The north waste rock facilities (WDN) will be constructed with wrap-around accesses near the bench exits. The south waste rock facility (WDS) will be built from the bottom up with overall slopes of 21.8° (2.5H:1V) and 30 m lift heights.

Mine Schedule
The mining rate or 51 Mtpa was selected based on strategic planning scenarios which demonstrated that the targeted mill capacity 45 ktpd (16.2 Mtpa) of would be achieved. Year 1 of process production was reduced to account for plant ramp-up. Two years of pre-production were utilized to develop pioneering roads to mining areas, construct the crusher and stockpile pads, and ensure adequate stockpile material to allow consistent crusher feed in year 1.

When mining starts, various infrastructure items will be under development. Key significant activities near the pit will include construction of the crusher and stockpile pad. A suitable rock drain will need to be established in the drainage below the stockpile pad. A pioneer road up the mountain will need to be developed early in the schedule to begin mining in phase 2. Phases 1B and 1 will also require a pioneer road so that nearby waste may be sent to the stockpile pad.

During preproduction, phase 1B will act as a quarry for stockpile pad construction material while phase 1 will be an early ore source once stockpile space is available. Once pioneering is completed to the top of the mountain, phases 2 and 3 will commence mining downward with waste being sent to the north wrap-around style waste rock facilities.

By the end of year 1, it is anticipated that the process facilities have completed the ramp-up production and the mining layout has progressed. At this time, the phase 1B will have been completed and reached 1295 m elevation. Phase 1 will be the primary source of ore and will be mined down to 1365 m elevation. Phase 2 will be mined down to 1685 m elevation and access to the phase 3 will be left by a combination of ramps in the pit wall or north waste rock facilities. The upper benches of phase 3 will be mined down to 1915 m elevation.

Years 2 to 5 will have mining in phases 1, 2 and 3 with resulting mining layout for year 5. Phase 1 will be mined down to 1245 m elevation with completion in year 4. Waste from phase 1 will used for widening the stockpile pad or starting the lower levels of the south waste rock facility. Phase 2 will be mined down to 1435 m elevation with waste being sent primarily to the north waste rock facilities, but the lower benches would have the option of using longer hauls to send some waste to the south waste rock facilities. Phase 3 will continue mining down to 1675 m elevation with all waste being sent to the north waste rock facilities.

Years 6 to 10 will have mining in phases 2 and 3 with the resulting mining layout for year 10. Mining in phase 2 will progress down to 1255 m elevation with most waste being sent to the south waste facilities. Mining in phase 3 will progress down to 1465 m elevation with waste being sent to the north waste rock facilities until year 9, and then starting to use the south waste rock facilities as the waste destination. While mining phase 3 in year 10, a geotechnical berm will be established at 1475 m elevation across the full length of the pit wall. This geotechnical berm is planned to be used to facilitate a water diversion around the west side of the pit.

Years 11 to 15 will have mining in phases 2 and 3 with the resulting mining layout for year 15. As the phases are now at lower elevations, all waste will be directed to the south waste rock facilities. The south waste rock facility is expected to reach 1440 m elevation by year 15. Phase 2 will be mined complete to 1225 m elevation in year 11. Phase 3 is expected to reach 1325 m elevation by year 15 with the haul road access transitioning to the east side of the pit.

Years 16 to 19 will have mining only in phase 3 with the final mining layout. The south waste rock facilities will receive all waste during these final years and will reach a final lift elevation of 1470 m. Phase 3 will be mined down to a final elevation of 1125 m by year 19.

Mine Equipment Selection
The mining equipment selected to meet the required production schedule is conventional mining equipment, with additional support equipment for snow removal and surface ditching maintenance.

Drilling will be completed with down the hole hammer (DTH) drills with a 165 mm bit. This provides the capability to drill 10 metre bench heights in a single pass.

The primary loading units will be 37 m³ hydraulic shovels. Additional loading will be completed by 21 m³ loaders. It is expected that one of the loaders will be at the primary crusher for the majority of its operating time. The haulage trucks will be conventional 240 tonne rigid body trucks.

Comminution

Crushers and Mills

Summary:

Process Description
240-tonne mine haul trucks tip into a single 450kW gyratory crusher station designed with a planned 85% circuit availability and located adjacent to the pit. Should the crusher feed pocket be full on arrival, the mine truck has the option of dumping run-of-mine material (“ROM”) on the ground for later reclamation. Crushed ROM material is transferred from the crusher surge pocket to the mill circuit by an overland conveying system, incorporating a single flight 2,700 kW conveyor that for the most part follows the mine-mill access road down the valley.

Surge capacity between the primary crusher station and the mill circuit is handled by a single 30,000- tonne coarse rock stockpile adjacent to the process plant. Crushed ROM is withdrawn from the stockpile, in a controlled manner, using multiple apron feeders. SAG mill feed control is achieved using variable speed feeding with mill feed size distribution measurement and conveyor weigh scale. The mill circuit includes a 16,500kW twin pinion grate discharge SAG mill with pebble ports and a vibrating pebble screen. Pebbles are crushed using a 600-kW cone crusher to assist the comminution of critical size material that otherwise accumulates within the SAG mill and impacts performance.

Secondary milling to a product size of 80% -110 microns is achieved by two twin pinion 16,500 kW overflow ball mills, operating in parallel. Each ball mill operates in closed circuit with multiple hydro cyclones and discharges into a common mill tank. SAG mill discharge slurry also gravitates from the SAG mill via a single vibrating screen into the mill sump. The mill discharge tank contents are diluted with process water and then pumped via two 1,100 kW discharge pumps to the two parallel cyclone clusters associated with the two ball mills.

Primary Crushing
Run-of-Mine (ROM) material is delivered to the primary tip location by 240t mine trucks at an average frequency of roughly 15 to 16 trucks per hour. Trucks tip ROM directly into the primary crusher throat at a peak delivery rate of approximately 2,400 dmtph where it is crushed between the crusher mantle and the concave shell.

The selected primary crusher, a modern 42-65” gyratory unit, accepts feed material up to 800 mm and runs with a 165 mm open side setting. Crushed rock discharges by gravity into an 800-tonne rail-lined surge pocket, which provides at least 20 minutes of surge capacity between the crusher and the stockpile. An apron feeder is used to withdraw crushed rock from the surge pocket onto a short sacrificial conveyor. This conveyor discharges onto the main overland conveyor, which in turn transports material down the valley to the crushed ROM stockpile area.

Spillage and run-off in the primary crusher building is pumped to surface and co-mingled with mine water as it is discharged down to the lower plant area for treatment. The primary crusher area is served by a contract mobile crane for routine maintenance.

Stockpile
The overland conveyor transports ROM material from the mine to the stockpile in the plant area. The crushed ROM stockpile is uncovered and provides a live capacity equivalent to roughly 12-14 hours of plant production. Mill feed is withdrawn from the stockpile via four lined discharge chutes and up to four apron feeders (nominally two operating, two standby). Each apron feeder is variable speed and capable of providing up to 80% of the nominal mill feed rate. Feeders discharge via lined chutes onto the SAG mill feed conveyor.

SAG & Ball Milling From the stockpile discharge apron feeders, mill feed material is withdrawn at a controlled rate onto the SAG mill feed conveyor, which discharges into the SAG mill feed hopper. Key process variables such as the throughput and the particle size distribution are monitored using an accurate weightometer and high-speed camera system respectively and controlled using standard plant control systems.

The selected SAG mill is 38-foot diameter x 19-foot long, with grate discharge and with a 2 x 8,250-kW twin pinion drive system. Fresh ROM rock and water are added to the SAG mill and slurry and pebbles exit the mill after passing through the discharge grate or pebble ports onto a vibrating scalping screen. Scalping screen oversize material (consisting mainly of pebbles) is directed by chute onto the pebble recycle conveyor for crushing and recycling to the mill. Scalping screen undersize slurry (-2 mm) gravitates into the common mill discharge sump, where it is mixed with ball mill discharge and further diluted with water.

From the discharge sump, the coarse mill discharge slurry is pumped to the ball mill cyclones.

SAG mill slurry spillage is collected in a drive-in sump, and then returned to process by a submersible slurry pump. The solids from large spillages can be recovered using a bobcat or small loader.

The milling area (SAG and ball mill x 2) is served by two 15t overhead cranes. Mill lining replacement is carried out using a single shared 5-axis relining machine.

SAG mill grinding media (5” diameter) is stored in ball bunkers located partway along the mill feed conveyor belt. The bunkers are served with a small spillage pump and a ball loading crane and magnet. Balls are added to mill feed on the SAG feed conveyor at timed intervals using a ball loading chute.

After SAG milling, the nominal particle size is further reduced by a pair of conventional, closed circuit ball mills operating in parallel. Each mill is 26.5-foot diameter x 40-foot long with overflow discharge. Slurry is pumped from the common mill discharge sump to two independent clusters of cyclones, with each cluster consisting of 8 x 28” units. Cyclone underflow of roughly 70% solids discharges from the cyclone spigots to the ball mill for more grinding. Each ball mill is equipped with a 2 x 8,250-kW twin pinion drive system and discharges ground pulp into the common mill discharge sump.

The overflow from each cyclone cluster gravitates to a sampling launder and automatic sample cutter before passing into conditioning tanks ahead of flotation. Particle size distribution is monitored automatically using an online particle size analysis (PSA) system that provide control system data - viewable and actionable from the control room.

Spillage contained in the ball mill area is pumped to the common mill discharge sump for re-treatment.

Ball mill grinding media is delivered to the Project site in bulk and is stored in the ball mill ball bunkers. The ball bunkers are serviced by a crawl and electric hoist arrangement, allowing balls to be lifted into a kibble using the ball loading magnet, and tipped into the mill feed spout via a rubber lined ball loading chute.

Processing

• Column flotation
• Crush & Screen plant
• Flotation
• Magnetic separation
• Dewatering
• Filter press

Summary:

A nickel-copper-PGM mineral process plant suitable for the Project located near Burwash, SW Yukon Territory. A nominal throughput of 45,000 metric tonnes per day has been selected for the purposes of this study.

The selected flowsheet is in most respects a very conventional mineral processing circuit, consisting of primary gyratory crushing, overland conveying, Run of Mine (“ROM”) stockpiling, SAG and ball mill grinding, froth flotation, low intensity magnetic separation, concentrate dewatering, and tailings thickening. The flotation process includes a cleaning circuit designed to separate copper and nickel minerals (Copper-Nickel Separation).

Studies of the project included a trade-off of different concentrator configurations for optimized costs and benefits. Two different flowsheet configurations were considered:
• an acidic bulk flotation process, complete with magnetite separation and regrinding, to produce a bulk sulphide (Cu+Ni) flotation concentrate
• an alkaline bulk flotation process, complete with magnetite separation and regrinding plus a copper-nickel separation circuit (multiple stages of selective flotation to make separate Cu-rich and Ni-rich concentrates as opposed to a single mixed product)

Process Description
Cyclone overflow slurry gravitates from the cyclone packs through a sampling station and into a pair of surge/conditioning tanks ahead of the rougher flotation circuit. The rougher flotation plant consists of six 500 m3 tank cells in series, with each cell having independent air flow and individual pulp level control.

Rougher flotation concentrate is reground in a 2,300 kW inert media vertically stirred mill to a P80 of 25 µm and then cleaned in a three-stage cleaner circuit with a cleaner scavenger circuit on 3rd cleaner tails. Bulk concentrate from the 3rd cleaner is pumped to a dual-purpose concentrate thickener from where the underflow slurry is either:
• pumped directly to concentrate pressure filtration equipment for dewatering and sale as a bulk (Cu + Ni) concentrate
• pumped to the head of the copper-nickel flotation circuit for separation of copper and nickel minerals to give separate copper and nickel concentrates

Rougher flotation tailing slurry is pumped to a magnetite removal circuit, consisting of two stages of 1.2 m diameter x 4 m long wet low intensity magnetic separation (LIMS). The LIMS equipment is configured to remove a magnetite concentrate from the rougher tailing slurry and direct this stream to a pair of 5,000 kW inert media regrind mills for regrinding, prior to rougher flotation for further recovery of valuable minerals. Magnetite rougher flotation concentrate is pumped to the first sulphide cleaner scavenger cell whilst magnetite rougher tailing slurry is sampled and pumped to the tailing thickener for dewatering and disposal at the TMF.

The copper nickel separation process included in the flowsheet is a selective flotation process that generally occurs at high pH. Lime and other reagents are added to the bulk concentrate prior to Cu/Ni separation flotation. The pulp must be aerated for approximately 10 minutes in the presence of activated carbon at this point to facilitate adsorption of any excess sulphide collector. In copper nickel separation, the copper minerals are collected very selectively, and the nickel minerals are depressed. Column flotation cells are used for cleaning the Cu/Ni separation rougher concentrate, and column flotation concentrate is pumped to the copper concentrate dewatering circuit. Cu/Ni separation scavenger tailing pulp is pumped to the nickel concentrate dewatering circuit.

Bulk Flotation
Cyclone overflow pulp is conditioned in a pair of conditioning tanks – each sized to give six minutes conditioning time ahead of rougher flotation. The bulk rougher/scavenger bank consists of six 500 m3 cells operating in series. Flotation air is supplied using flotation blowers via a low-pressure manifold, and air flow to each cell is controlled by modulating valves and flow meters. Pulp level within cells is maintained by modulating dart valves and ultrasonic pulp level instruments.

Bulk Cleaner Flotation
After regrinding, rougher concentrate slurry is pumped to the first cleaner circuit, which consists of 4 x 130 m3 cleaner tank cells and 4 x 200 m3 cleaner scavenger tank cells. First cleaner concentrate is collected and pumped to the head of cleaner 2, while first cleaner tails gravitate together with magnetite rougher concentrate into the cleaner scavenger cells. Cleaner scavenger concentrate is pumped back to the head of the first cleaner and cleaner scavenger tailing slurry is pumped to the tailing thickener for dewatering and disposal.

Magnetite Scavenger Circuit
The bulk rougher flotation tailing slurry is sampled as it discharges the final bulk rougher flotation cell into the rougher tailing tank. From here it is pumped to a distribution box above the magnetic separation plant. This plant consists of two stages of low intensity wet magnetic separation in series, with tailing slurry from the second stage being pumped to the rougher tailing thickener. Three rougher and twelve scavenger drum separators have been allowed in the design, with scavenger machines utilizing slightly higher gauss configurations than the rougher machines.

Concentrate Regrinding Circuit
The concentrate regrinding area consists of two independent regrind streams, namely the bulk rougher concentrate regrind circuit and the magnetite concentrate regrind circuit.

Magnetite Concentrate Regrind
Magnetic separator circuit concentrate gravitates to the magnetite regrind circuit, which grinds the concentrate to approximately 80% -17µm. The regrind circuit consists of a circuit feed tank, a cluster of scalping cyclones, two 5MW vertically stirred mills in parallel, and an agitated product tank.

The 2 inert regrind mills are served by a single ceramic media addition system.

Rougher Concentrate Regrind
Bulk rougher concentrate gravitates to the bulk rougher regrind circuit, which grinds the concentrate to approximately 80% passing 25µm. The regrind circuit consists of a circuit feed tank, a cluster of scalping cyclones, a single 3.5MW vertically stirred mill, and an agitated product tank and is very similar in configuration to the Magnetite Concentrate Regrind circuit. The inert media mill is served by a single ceramic media addition system.

After regrinding, concentrate slurry is pumped to the bulk cleaner flotation circuit for further upgrading.

Copper Nickel Separation Flotation – Area 450
This uses a change to the pulp electrochemistry to depress nickel and iron sulphide minerals, while allowing copper sulphide minerals to float. In this way, the Cu/Ni separation concentrate becomes enriched in copper, and the Cu/Ni separation tailing stream is depleted in copper and becomes richer in nickel.

Copper Concentrate Dewatering – Area 500
When in use, the Cu/Ni separation circuit pumps final copper concentrate slurry to the copper concentrate thickener sampling launder and sampler before entering the thickener tank for settling and dewatering. These 10 m-diameter thickeners are equipped with a rake lift, bed level detection, and bed mass monitoring. Thickener overflow gravitates to a common spray water tank for recycling as flotation sprays, while the thickener underflow is withdrawn from the cone by a centrifugal underflow pump and pumped to mechanically agitated storage tanks ahead of the multi-use pressure filter.

Bulk/ Nickel Concentrate Dewatering
The bulk concentrate thickener serves as the nickel concentrate thickener during periods where highCu ROM is fed to the plant. Nickel concentrate slurry is pumped from the Cu/Ni separation circuit to the nickel concentrate thickener-sampling launder and sampler before entering the thickener tank for settling and dewatering.

Water Supply

Summary:

Water supply adequate for drilling operations can be pumped from local creeks. Potable and nonpotable water has been sourced from a shallow well at Lower Camp. In 2015, a new well was drilled at Lower Camp to provide water to the lodging facilities during exploration. It is assumed sufficient water supplies from pit dewatering and surface run-off will be available for the mill processing needs of the Project.

Process Plant, and Tailings Management Facility (TMF)
The process plant requires approximately 5,000 m3 /h of process water from the holding pond along with roughly 70 m3 /h of raw water which is assumed to be substantially free of solids and neutral pH. The water balance assumes a 92% availability for the process plant. Thickener overflows from the process plant report directly to the holding pond and will do so year-round. Other water discharges from the process plant include approximately 1,300 m3 /h of water in tailing slurry to the TMF, and <20 m3 /h of water shipped in trucks as copper/nickel concentrate.

Decanted water from the TMF is pumped back to the holding pond for re-use in the process plant. This amounts to almost 800 m3 /h of water on average. It is assumed that approximately 60% of the tailings water reports back to the holding pond at all times of the year (the remainder being held interstitially within the TMF) and also that up to 40% of the water within the TMF becomes frozen during the winter months.

Holding Pond and Effluent Treatment Plant (ETP)
Water that has been collected from the crushing area can report either to the holding pond or to the TMF by way of ditches, channels, and pipelines. Thickener overflows from the process plant along with decanted water from the TMF shall also report to the holding pond.

Water flows from the upper site area do not provide sufficient make-up volume for the Process Plant at all times of the year. Between April and September, a surplus of water will be present, whilst in the winter months, a shortage will tend to occur. The Holding Pond and the storage facilities at the mine and the TMF have been sized to buffer these seasonal imbalances, thereby limiting the Holding Pond size to 250,000 m3 .

While there may be fluctuations during the year, on average the site has a water surplus and therefore treatment and discharge at certain times has been allowed for. The overall capacity of various ponds also allows for year-round operation of an Effluent Treatment plant, should that be required.

Based upon the various pond capacities and the expected seasonal variation of water flow, a 250 m3 /h effluent treatment plant (ETP) has been specified, assuming an availability of 92%. At the time of this study the Kluane River has year-round flow, with the lowest flow being 5,300 m3 /h in April and the mean yearly flow being almost 34,000 m3 /h.

Based upon the above, it was estimated that the holding pond should have a capacity of roughly 200,000 m3 . Due to the process plant water requirement, it has been assumed that the holding pond will be designed such that there will be liquid water available for the process plant at all times of the year.

Production

Operational metrics

Personnel

Job Title	Name	Phone	Profile	Ref. Date
Consultant - Mining, Infrastructure & Costs	Gordon Zurowski			Sep 20, 2023
Consultant - Recovery Methods & Costs	Andy Holloway			Sep 20, 2023
President and CEO	Stuart Harshaw	1-416-304-9318		Jan 30, 2024