Sulphide mineralization discovered to date in the project area can be characterized as Komatiite-hosted Ni-Cu-Co-(PGE) deposit type, which recognizes two sub-types (Lesher and Keays, 2002):

1. Type I – Kambalda-style: channelized flow theory; komatiite-hosted; dominated by net textured and massive sulphides situated at or near the basal ultramafic/footwall contact with deposits commonly found in footwall embayments up to 200 m in strike length, 10s to 100s of meters in down-dip extent, and meters to tens of meters in thickness; generally on the order of millions of tonnes (generally <5 Mt) with nickel grades that are typically much greater than 1% nickel; tend to occur in clusters (e.g., Alexo-Dundonald, Ontario; Langmuir, Ontario; Redstone, Ontario; Montcalm, Ontario; Thompson, Manitoba; Raglan, Quebec).

2. Type II – Mt. Keith-style: sheet flow theory; thick komatiitic olivine adcumulate-hosted; disseminated and bleb sulphides, hosted primarily in a central core of a thick, differentiated, dunite-peridotite dominated, ultramafic body; more common nickel sulphides such as pyrrhotite and pentlandite but also sulphur poor mineral Heazlewoodite (Ni 3 S 2) and nickel-iron alloys such as Awaruite (Ni 3 -Fe); generally on the order of 10s to 100s of million tonnes with nickel grades of less than 1% (e.g., Mt. Keith, Australia; Dumont Deposit, Quebec).

Sulphide nickel-copper-cobalt-PGE mineralization in the Crawford Ultramafic Complex is interpreted as most similar to Mt. Keith-style.

The project is situated in the Timmins-Cochrane Mining Camp of Northeastern Ontario, in the western portion of the mineral- rich Abitibi Greenstone Belt (2.8 to 2.6 Ga), which is within the Superior Province, Canada. The Abitibi Greenstone Belt of the Abitibi Subprovince, spans across the Ontario-Quebec provincial border and is considered to be the largest and best- preserved greenstone belt in the world (Jackson and Fyon, 1991; Sproule et al., 2003). The Timmins-Cochrane Mining camp has a history of nickel production from komatiite-associated nickel-copper-platinum group element (Ni-Cu-(PGE)) deposits.

Recent work (2003-2012) suggests that the rocks underlying the property are part of the Deloro Assemblage (Monecke et al., 2017). The Deloro Assemblage (2730 to 2724 Ma) hosts the Crawford Ultramafic Complex (CUC) and consists mainly of mafic to felsic calc-alkaline volcanic rocks with local tholeiitic mafic volcanic units and an iron formation cap which is typically iron-poor, chert-magnetite (Ayer et al., 2005; Thurston et al., 2008).

The principal target, the CUC is entirely under cover but based on geophysics and drilling is an approximately 8.0 km long by 2.0 km wide body (original estimated shape) of dunite, peridotite (and their serpentinized equivalents), and lesser pyroxenite and gabbro, as confirmed in recent historical diamond drill holes (Spruce Ridge Resources, 2018) and the current extensive drilling program by CNC. Historical diamond drilling in the 1960s and 1970s also reported intersections of gabbro, peridotite, pyroxenite, dunite and serpentinite (e.g., George, 1970). Descriptions from drill core logs record localized brecciation in the Main Zone at the northern contact between mafic volcanic rocks and dunite.

Sulphide mineralization discovered to date on the Crawford property can be characterized as komatiite-hosted Ni-Cu-Co- platinum-group-elements (PGE) deposit type, which recognizes two sub-types (Lesher and Keays, 2002). Sulphide nickel-copper-cobalt-PGE mineralization in the CUC is interpreted as most similar to Mt. Keith-style. Mt. Keith-style (Type II) is based on sheet flow theory (Lesher and Keays, 2002) and is characterized by thick komatiitic olivine adcumulate hosted, disseminated and bleb sulphides, hosted primarily in a central core of a thick, differentiated, dunite peridotite dominated, ultramafic body. More common nickel sulphides, such as pyrrhotite and pentlandite, are present along with the sulphur- poor mineral heazlewoodite (Ni 3 S) and nickel-iron alloys such as awaruite (Ni 3 -Fe). These deposit types are generally on the order of 10s to 100s of million tonnes with nickel grades of less than 1% (e.g., Mt. Keith, Australia; Dumont Deposit, Quebec).

The Crawford mine plan currently includes three discrete open pits. The largest and first to be mined is the Main Zone, comprising 77% of the total mined tonnage, 80% of feed to the mill and associated contained nickel, and 85% of the recoverable nickel. Immediately north of the Main Zone is the East Zone, which is divided by a saddle into East Zone-West (EZ-W) and EastZone-East (EZ-E). EZ-W is the larger of the two, comprising 16% of the total tonnage to be mined, 14% of feed to the mill and associated contained nickel, and 11% of the recoverable nickel. Mining of the EZ-W will begin in the last year of Main Zone operation. EZ-E will be mined last, starting 33 months after EZ-W.

All waste generated while the Main Zone pit is active will be impounded on the surface. Tailings from the material processed through the mill will be stored in the tailings storage facility (TSF) located to the east of the pits. This facility will be constructed in two phases. The starter phase will use a downstream construction method to its ultimate 36 m height, utilizing waste rock and gravel from the open pit. The much larger ultimate phase will also use the downstream construction method to a height of 49 m before converting to an upstream method utilizing mainly rehandled tailings for the final 24 m. Gravel that is not used to construct the TSF as well as clay will be impounded in two overburden storage facilities (OBS-S and OBS-E) to the south and east of the Main Zone. Waste rock not used in constructing the TSF will be impounded in a waste rock storage facility to the north of the East Zone (WRS-N). The Main Zone will be mined at a rate faster than is required to satisfy the mill, allowing for accelerated delivery of higher value material to the mill while lower value material will be temporarily stockpiled in one of three low-grade stockpiles. LG1 will be comprised of highest value material and located closest to the mill. Next highest value material will be stored in LG2. Both LG1 and LG2 will be located within the footprint of the future EZ-E pit. At the time that EZ-E starts up, both these stockpiles will have been completely reclaimed, with only LG3 to the north of the East Zone remaining. Waste from the EZ-W will initially be impounded in surface facilities. As soon as operations in the Main Zone cease, all remaining waste from EZ-W and EZ-E (including tailings produced by the mill) will be impounded in the mined-out Main Zone pit. Ultimately, the Main Zone void will be 55% filled by the various waste stored within it.

The pits will be mined with a mixed fleet of equipment. A portion of the deepest clays will be loaded with small backhoe excavators and hauled using 40 t articulated trucks. The remaining clay and a portion of the underlying gravel will be loaded with medium-sized face shovel excavators and hauled using 90 t manually operated haul trucks. The remaining gravel and all the underlying waste rock will be loaded using large-face shovel excavators and 290 t haul trucks that utilize an autonomous haulage system (AHS) and have no onboard operator. Starting in Year 2 of operations, trolley assist will be employed to reduce the diesel consumed while increasing the speed of trucks on uphill hauls.

Pit shells generated using the LG algorithm represent a theoretical design and, while the final walls honour a best estimate of slopes taking account of ramps, the shell cannot be considered practical because ramps have not been located. The engineered design provided for ramps measuring up to 54 m. The engineered design also translates the overall slope angles used by LG into bench face angles and berms as follows:
• Clay (6H:1V / 9.5°): 5 m bench height with a 30° face angle and 25 m berm.
• Gravel (2.5H:1 / 22°): 5 m bench height with a 60° face angle and 10 m berm.
• Rock (1H:1 / 22°): 30 m bench height (double benching) with a 70° face angle and 19 m berm.

It can be seen that the engineered designs honour mill feed contained within the LG shells, with an aggregate planned mining recovery of 99% and dilution of just 2%. However, this was achieved at the expense of a 22% increase in waste stripping, as a result of the inclusion of ramps in the design while adopting a conservative approach of maintaining a 45° inter-ramp slope angle in rock.

Attention is directed to the following:
• The Main Zone operates for 6 years following the completion of clay stripping, while the East Zone operates for 4 years. With more detailed scheduling it may prove possible to defer some clay stripping.
• The Main Zone ramps up to a peak production rate of 190 Mt / a (520 kt / d). The mining fleet peaks at this time, at seven 700 t excavators and forty-nine 290 t trucks. While the duration that peak production is maintained is relatively short, immediately afterwards the initial units of fleet reach their economic lives and are retired and no replacement truck purchases are required (two excavators are replaced over the life of mine). The open pit thus achieves a relatively high level of capital efficiency.
• At the end of Main Zone life, the combined inventory of material in higher grade stockpiles LG1 and LG2 totals 53 Mt with an average NSR value of C $ 30 / t, which is 25% higher than the average value of all East Zone material. Start-up of the East Zone is thus deferred until these stockpiles are largely depleted, leading to the dip in ex-pit production in Year 17. Over the life of mine, 25% of total mill feed is temporarily stockpiled. The average residence time is 14 months in LG1, 33 months in LG2, and 128 months in LG3.

Ex-pit mining operations at Crawford will be conducted by the following fleets of production mining equipment:
• Areas where the depth of clay exceeds 7.5 m will be mined using 200 t backhoes loading 40 t articulated trucks. The backhoes will load from on top of the clay, requiring the surface to be armoured with crushed rock to prevent sinking. No drilling and blasting will be required for the overburden. The nominal bench height will be 5 m.
• Areas where the depth of clay is = 7.5 m will be mined using a 300 t hydraulic excavator operating in face shovel configuration. The excavator will load from the underlying gravel footwall and deliver all material into 90 t rigid body haul trucks. No drilling and blasting will be required for the overburden. The nominal bench height will be 7.5 m.
• Below the lowest horizon of the clay/gravel interface, the horizon of mixed gravel and rock will be mined using a 700 t, electrically-powered hydraulic excavator operating in face shovel configuration. The excavator will deliver material into 290 t haul trucks using AHS to replace on-board operators. From Year 2 onwards, trucks will also be equipped to use trolley assist on equipped uphill hauls. No drilling and blasting will be required for the overburden, while rock will be drilled using rotary blast hole drills with a nominal hole diameter of 229 mm. Drills will be equipped with ADS to replace on-board operators. The nominal bench height in mixed gravel and rock will remain at 7.5 m.
• Below the lowest horizon of the gravel/rock interface, the bench height will be increased to 15 m with the bulk of material blasted using holes measuring 311 mm in diameter. Where selective mining is required, each bench will be divided into two 7.5 m flitches that will be drilled using 229 mm holes. Selective mining will be conducted using the same 700 t excavators and 290 t AHS and trolley-assist equipped haul trucks as will be used for bulk mining.

Production equipment will be supported by various units of support equipment, including tracked dozers, wheel dozers, front-end loaders, graders, water tankers, and utility excavators.

It has been assumed that a mining contractor would be employed in the initial year of pre- stripping. Thereafter, all mining fleet would be purchased and operated by the Owner.

Primary Crushing
Based on the design throughput and moderate competency ore characteristics, a single 60 x 89 gyratory crusher is considered the most suitable for the primary crushing duty for the initial plant and Phase 2 expansion. For the 120 kt/d expansion, two crushers will be required. The primary crushers will be located at the edge of the ROM pad. A partially buried crusher design has been selected to reduce ROM pad elevation (reduce mine haulage costs) without major excavation being needed. Trucks will dump from both sides of a 323 m³ capacity live hopper above the crushers. Alternatively, ore can be rehandled and fed with a front-end loader (FEL).

Primary crushing will be installed inside an enclosed building. This will help minimize dust emissions and reduce noise. An overhead crane will be installed in the building for maintenance of the crushers. Auxiliary crusher equipment includes a mobile rock breaker and dust suppression system. Water sprays are used to minimize dust at the crusher bin, crusher discharge and crusher belt feeder. The gyratory crushers will crush ore to a product size of 80% passing 68 mm.

Secondary Crushing
For the 120 kt/d expansion only, secondary crushing will be required. The secondary crushing stage will be an open circuit, with a classification screen after primary crushing. The screen oversize feeds the three secondary cone crushers, model MP 1250 or similar, whose products report to the crushed ore stockpile. Screen undersize also reports to the crushed ore stockpile.

Secondary crushing will be installed in an enclosed building. An overhead crane will be installed in the building for maintenance of the crushers. Auxiliary crusher equipment includes a mobile rock breaker and dust suppression system. Water sprays are used to minimize dust at the crusher bin, crusher discharge and crusher belt feeder. The cone crushers will crush ore to a product size of 80% passing 46 mm.

Crushed Ore Stockpile
Crusher product will be conveyed from the crusher discharge vault by the variable-speed primary crusher discharge belt feeder and discharged onto the sacrificial conveyor. Ore is transferred via the stockpile feed conveyor to the crushed ore stockpile. A weightometer will be installed on the sacrificial conveyor to provide production rate data for the crushing circuit. A cross belt self-cleaning electromagnet, followed by a metal detector, is fitted over the sacrificial conveyor to detect and remove tramp steel prior to discharge onto the stockpile feed conveyor. The stockpile will provide a minimum of 17 hours’ live capacity at the nominal 42.5 kt/d SAG mill fresh feed rate. The live capacity reduces to 12 hours once the plant is expanded to the 60 kt/d throughput. In the event of the crushing circuit being out of operation for extended periods, a bulldozer can be used to reclaim the dead material in the stockpile to provide emergency feed to the milling circuit. Three apron feeders have been selected to reclaim ore from the stockpile, each able to deliver 60% of the nominal mill feed rate. The stockpile will be enclosed to minimize fugitive dust emissions. The cover will be a dome of galvanized structural steel construction and cladding.

Grinding Design Criteria
The grinding circuit was designed to be capable of processing the required tonnage of 42.5 kt/d and will double throughput to 85 kt/d by mirroring the first line and thereafter will raise production to the ultimate rate of 120,000 t/d through the addition of secondary crushing and a third ball mill.

Flotation testwork and mineralogy have indicated that Crawford ores are relatively insensitive to grind sizes (P80 ) up to about 180 µm in the laboratory. In order to achieve the specified design recovery, Ausenco has nominated a primary grind size target of P80 of 200 µm.

The installed ball mill power of 17,500 kW incorporates allowances for drive train losses as well as a design contingency to account for the accuracy of the models, calculations and testwork used to determine the expected average pinion power.

The installed motor power for the SAG mills incorporates similar allowances, as well as an additional contingency to allow adjustment in the mill operating conditions to handle ore variability. These allowances and contingencies require an installed power of 17,500 kW per mill.

Reclaim, SAG & Ball Mill Circuit
The crushed ore will be reclaimed from the ore stockpile by three apron feeders onto the SAG mill feed conveyor. The feeders will be equipped with variable speed drives.

A SAG mill feed weightometer will be installed on each SAG mill feed conveyor to provide feed rate data for control of the reclaim feeders. The reclaimed crushed ore will be fed at a controlled rate to the SAG mill.

Discharge from the SAG mill will gravitate through a trommel screen. Oversized pebbles from the trommel screen (scats) will be recycled back onto the mill feed conveyor. A cross belt self- cleaning electromagnet removes broken and worn mill balls and other tramp steel from the scats stream. Pebbles will be reintroduced onto the mill feed conveyor via the recycle pebble conveyors. Undersize from the SAG trommel screen will gravity flow into the cyclone feed hopper.

A pebble circulating load of 18% to 22 % of the fresh feed rate has been assumed in the design, based on typical industry experience with ores of similar competency. The conveyors are designed to handle peak loads of up to 30% of fresh feed.

The SAG mill discharge slurry will be pumped via dedicated cyclone feed pumps to the ball mill cyclone cluster in Phase 1 and three clusters in Phase 3, each operating in a closed-circuit configuration with a single ball mill. Water is added to the cyclone feed hopper as needed to achieve the required cyclone feed pulp density.

Hydrocyclone underflow from each cluster will gravity flow to a dedicated 17.5 MW twin-pinion ball mill (two 8.75 MW motors operating in parallel). Discharge from each ball mill will gravity flow through a trommel screen, into the cyclone feed hopper for reclassification. Cyclone overflow will gravity flow to the first stage deslime cyclone feed hopper.

The process plant and associated service facilities will process ROM ore delivered to primary crushers to produce nickel concentrate and tailings. The proposed process encompasses:
• crushing and grinding of the ROM ore;
• desliming via hydrocycloning;
• slimes rougher and cleaning flotation;
• coarse rougher, scavenger and cleaning flotation;
• regrind of the coarse flotation circuit final concentrate;
• high-grade material flotation;
• magnetic recovery of coarse scavenger tailings;
• regrinding of magnetic concentrate, scavenger concentrate and coarse first cleaner tailings;
• fines rougher and cleaning flotation of the reground material;
• magnetic recovery of fines rougher and first cleaner tailings.

Concentrate will be thickened, filtered and stockpiled on site prior to being loaded onto railcars or trucks for transport to third-party processing facilities. The coarse scavenger magnetic separation tailings wil ........