The Casino deposit is best classified as a calc-alkalic porphyry type deposit associated with a tonalite intrusive stock (the Patton Porphyry). Primary copper, gold and molybdenum mineralization was deposited from hydrothermal fluids that exploited the contact breccias and fractured wall rocks. Higher grades occur in the contact breccias, and grades gradually decrease outwards away from the contact zone, both towards the centre of the stock and outward into the host granitoids and schists. A general zoning of the primary sulphides occurs, with chalcopyrite and molybdenite occurring in the core tonalite and breccias, grading outward into pyrite-dominated mineralization in the surrounding granitoids and schists. Alteration accompanying the sulphide mineralization consists of an earlier phase of potassic alteration and a later overprinting of phyllic alteration. The potassic alteration typically comprises secondary biotite andK-feldspar as pervasive replacement and includes veins and stockworks of quartz and anhydrite veinlets. Phyllic alteration consists of replacements and vein-style sericite and silicification.

Hydrothermal Porphyry Alteration
Crystallization and exsolution of hydrothermal fluids from Patton Porphyry (PP) magmas produced porphyry style CuMo-Au mineralization. Therefore, the Patton Porphyry, and associated Intrusive Breccia (IX), is genetically related to the Cu-Mo-Au mineralization of the deposit.

Hydrothermal alteration at the Casino property consists of a potassic core centered on and around the main Patton Porphyry body, in turn bordered by a contemporaneous, strongly developed and fracture controlled phyllic zone, a weakly developed propylitic zone, and a secondary discontinuous argillic overprint. Mineralized stockwork veins and breccias within the Casino Property are closely associated with the hydrothermal alteration.

Supergene Porphyry Mineralization
The Supergene Zone is comprised of the Supergene Oxide (SOX) zone and the more extensive Supergene Sulphide (SUS) zone. The Leached Cap (oxide gold zone) is copper-depleted due to supergene alteration, mainly leaching, processes, and has a lower specific gravity relative to the other supergene zones. It averages 70 m thick and is characterized by boxwork textures filled with jarosite, limonite, goethite, and hematite. This weathering has completely destroyed rock textures and has replaced most primary minerals with clay. The resulting rock is pale gray to cream in colour and is friable to the touch, and the clay is commonly stained yellow, orange, and/or brown by iron oxides. The weathering is most intense at the surface and decreases with depth.

The poorly defined Supergene Oxide zone (SOX) is copper enriched with trace molybdenite. It occurs as a few perched bodies within the leached cap, likely due to more recent fluctuations in the water table. This zone is thought to be related to present-day topography and is best developed where oxidation of earlier secondary copper sulphides occurs above the water table, typically on well drained slopes. Where present, the supergene oxide zone averages 10 m thick, and may locally contain chalcanthite, malachite and brocanthite, with minor azurite, tenorite, cuprite and neotocite. Where present, the supergene copper oxide zone grades into the better-defined supergene copper sulphide zone.

Supergene copper mineralization occurs in a weathered zone up to 200 m deep, below the leached cap and above the hypogene zone. It has an average thickness of 60 m and is positively correlated with high grade hypogene mineralization, high permeability and phyllic and/or outer potassic alteration. Grades of the Supergene sulphide zone vary widely, but are highest in fractured and strongly pyritic zones, due to their ability to promote leaching and chalcocite precipitation. Thus, secondary enrichment zones are thickest along contacts of the potassic and phyllic alteration halos; accordingly, the copper grades in the Supergene Sulphide zone are almost double the copper grades in the Hypogene zone (0.43% Cu versus 0.23% Cu). Grain borders and fractures in chalcopyrite, bornite and tetrahedrite may be altered to chalcocite, diginite and/or covellite. Chalcocite also locally coats pyrite grains and clusters, and locally extends along fractures deep into the hypogene zone. Molybdenite is largely unaffected by supergene processes, other than local alteration to ferrimolybdite. In drill-core, the SUS zone is generally broken with decreasing clay alteration and weathering with depth and is ‘stained’ dark blue to gray.

Hypogene Mineralization
Mineralization of the Casino Cu-Au-Mo deposit occurs mainly in the steeply plunging, in-situ contact breccia surrounding the Patton Porphyry intrusive plug. It was formed by crystallization and exsolution of hydrothermal fluids from late Cretaceous magmas of the Casino Plutonic Suite. The breccia forms an ovoid band around the main porphyry body with dimensions up to 250 m and has an interior zone of potassic alteration surrounded by discontinuous phyllic alteration, typical of some porphyry deposits.

Hypogene mineralization occurs throughout the various alteration zones of the Casino Porphyry deposit as mineralized stockwork veins and breccias. Field relationships show that the potassic alteration occurred first as mineralized quartz veins of the phyllically altered zones, which cut those of the potassically altered zones. Significant Cu-Mo mineralization is related to the potassically-altered breccias surrounding the core Patton Porphyry, as well as in the adjacent phyllically-altered host rocks of the Dawson Range Batholith.

Mineralization in the potassic zone comprises mainly finely disseminated pyrite, chalcopyrite and molybdenite, as well as trace sphalerite and bornite. The phyllic alteration zones have increased gold, copper, molybdenite and tungsten values concentrated within disseminations and veins of pyrite, chalcopyrite and molybdenite along the inner part of the pyrite halo. The pyrite halo occurs within the phyllic alteration zone along the potassic-phyllic contact and discontinuously surrounds the main breccia body. It is host to the highest copper values on the property.

Pyrite to chalcopyrite ratios range from less than 2:1 in the core of the deposit, to greater than 20:1 in the outer phyllic zones. Locally, coarse grained bornite and tetrahedrite are intergrown with chalcopyrite. Molybdenite is not generally intergrown with other sulphides and occurs as selvages in early, high temperature, potassic quartz veins and as discrete flakes and disseminations. Native gold can occur as free grains (50 to 70 microns) in quartz and as inclusions in pyrite and/or chalcopyrite grains (1 to 15 microns). High grade smoky quartz veins with numerous specks of visible gold have been reported to exist.

Structurally Hosted Gold Mineralization
Structurally controlled gold mineralization within the Canadian Creek portion of the Casino property occurs mostly in the northwestern part of the property. Drilling in 2009 and 2017 discovered widespread anomalous gold mineralization associated with clay altered-shears, sheeted pyrite veins and quartz-carbonate veins hosted in both intrusive and metamorphic rocks. To date, the identified structures are generally less than 3 m thick and of short strike length. Gold is accompanied by silver, arsenic, antimony, molybdenum, barium and bismuth.

Mine operations will consist of drilling large diameter blast holes (31 cm), blasting with a bulk emulsion, and loading into large off-road trucks with cable shovels and a hydraulic shovel. Resource amenable to processing will be delivered to the primary crusher or various resource stockpiles. Waste rock will be placed inside the limits of the tailings management facility (TMF). There will be a fleet of track dozers, rubber-tired dozers, motor graders and water trucks to maintain the working areas of the pit, stockpiles, and haul roads.

The following general parameters guided the development of the mining plan:
- Mill material is limited to about 1.2 billion tonnes, CMC elected to limit the capacity of the TMF to be comparable to the concept and overall physical characteristics of the TMF design favored in the Best Available Tailings Technology Study (BATT study);
- Total mine waste to be co-disposed with tailings is limited to about 600 million tonnes;
- Mill capacity is a nominal 120,000 tonnes per day (t/d), but actual plant throughput for the schedule is based on hardness of the various material types, and usually exceeds 120,000 t/d.

Forty-five-degree inter-ramp angles were recommended for most of the slope sectors. The north sectors of the main pit and west pit were recommended to be designed at 42-degree inter-ramp angles. For the small amount of overburden on the north wall, the recommended angle was 27 degrees. The slope angle recommendations also specified that there be no more than 200 m of vertical wall at the inter-ramp angle without an extra wide catch bench (16 m instead of 8 m).

The overburden is placed in the overburden stockpile in Canadian Creek, north of the pit. The remaining waste is disposed in the tailing management facility in three facilities for mine waste: 1) the North Waste area which contains 248.4 million tonnes, 2) the Divider Dam which contains 134.4 million tonnes, and 3) the West Waste storage area which contains 164.6 million tonnes. About 5 million tonnes of mine waste will be used in the Starter Dam for the TMF embankment. The material will be placed by trucks and dozers; the rising water and tailings level in the TMF facility will cover the material before the end of the mine life.

The stockpiles are all constructed in lifts from the bottom up. The low-grade stockpile, leach stockpile, and SOX stockpile are designed with 30 m lifts at angle of repose with a 20 m setback between lifts to make the overall slope angle about 2H:1V. This is assumed to be adequate since these are not permanent facilities.

The mine production schedule is based on five mining phases. The designs utilized 40 m wide roads at a maximum grade of 10%. The road width will accommodate trucks up to the 370-tonne class such as the Komatsu 980E.

Total mill ore is 1.22 billion tonnes at 0.189% copper, 0.217 g/t gold, 0.0213% moly, and 1.66 g/t silver. For Years 2 through 26, full production years, ore throughput varies from a low of 45.0 million tonnes in Year 19 to a high of 47.1 million tonnes in Year 2. The average recovered copper grade is 0.164%, indicating an average copper recovery of 86.4%.

Direct feed ore is ore that is scheduled to be processed the same year it is mined. This amounts to 913.9 million tonnes at 0.206% total copper, 0.230 g/t gold, 0.0240% moly and 1.77 g/t silver. This is about 75% of total ore. The average NSR value of this ore is $27.77 per tonne. Note that Year 1 ore production is 34.6 million tonnes, about 73% of capacity in terms of plant hours and is made up of ore mined during preproduction and Year 1.

The Supergene Oxide (SOX) ore in the mining phase 1 is stockpiled and processed during Years 4 through 13 at the rate of 3.6 million tonnes per year. This is done to maintain the ratio of weak soluble copper to total copper at relatively low levels by year. This material amounts to 35.3 million tonnes at 0.275% total copper, 0.098% weak soluble copper, 0.500 g/t gold, 0.0254% moly, and 2.31 g/t silver.

The operating schedule also results in a significant amount of low grade material that is stockpiled and processed at the end of the mine life during Years 21 through 27. This amounts to 267.8 million tonnes at 0.120% total copper, 0.136 g/t gold, 0.0116 moly, and 1.19 g/t silver.

The reclaim schedules for both the SOX and low grade are on a last-in-first-out (LIFO) basis, consistent with stockpiles build up in lifts and reclaimed in reverse order.

Based on the schedule, the commercial life of the project is 27 years after an approximate 3-year preproduction period.

The schedule of mineral resource mined from the leach cap zone by year. This will be processed by crushing, and heap leaching. Leach resource is defined as leach cap material. This amounts to 209.6 million tonnes at 0.265 g/t gold, 1.95 g/t silver, and 0.036% total copper.

Life of mine total material from the pit is 2.04 billion tonnes. Preproduction is 75.0 million tonnes staged over three years. Year 1 total material is scheduled at about 95 million tonnes after which the peak material movement of 100 million tonnes per year is maintained through Year 14. Total waste is 611.3 million tonnes, so the waste ratio is about 0.44 if mill resource (including SOX), low grade, and leach resource are all counted as resource.

Stacking schedule for the leach resource.
This is based on the ability to crush and stack 9,125 kt/y (30,417 tpd for 300 days/year).

Mine production of leach resource.
Up to 9,125 ktonnes of mined material as direct crusher feed and the excess going to a stockpile. The stockpile gets to a maximum size of 79.2 million tonnes with this scenario.

Total waste in the mine plan amounts to 611.3 million tonnes. The waste material by material type is as follows:
- 58.5 million tonnes of overburden.
- 144.6 million tonnes of leach cap material.
- 33.2 million tonnes of supergene oxide material.
- 125.1 million tonnes of supergene sulphide material.
- 249.8 million tonnes of hypogene material.

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The Casino process plant will consist of two processing facilities, one for sulphide ore and one for oxide ore. The sulphide ore processing facility will produce mineral concentrates of copper and molybdenum using conventional flotation technology. The copper concentrate will be dewatered and transported as a filtered cake by highway trucks. The molybdenum concentrate will be dewatered and packaged in super sacks for transport. Gold and silver contained in the sulphide ore will be recovered as a fraction of the copper concentrate. The oxide ore processing facility will produce gold and silver Doré bars via heap leach and carbon adsorption technology. Copper contained in the oxide ore will be recovered as a copper sulphide precipitate using SART technology.

SULPHIDE MINERALIZED MATERIAL PROCESS PLANT DESCRIPTION
Bulk (Copper/Molybdenum) Flotation
Hydrocyclone overflow will flow by gravity to the bulk (copper-moly) flotation circuit. The copper moly flotation circuit will consist of two rows of mechanical rougher flotation cells, two rows of mechanical first cleaner flotation cells, two concentrate regrind mills operated in closed circuit with hydrocyclones, one row of mechanical second cleaner flotation cells, and two copper-moly third cleaner flotation column cells. Rougher flotation concentrate will flow by gravity to a sump and will be pumped by variable speed, horizontal centrifugal pumps to the first cleaner flotation circuit. Tailing from the rougher flotation cells will flow by gravity to the pyrite scavenger flotation circuit. Pyrite concentrate will join the first cleaner tailing at the pyrite thickener. Tailing from the pyrite flotation circuit (final tailing) will flow by gravity to the tailing thickeners. First cleaner flotation concentrate will flow by gravity to the regrind cyclone feed sump. Tailing from the first cleaner flotation cells will be combined with the concentrate from the pyrite scavenger flotation section in the pyrite thickener. Copper-moly concentrate regrinding will be performed in two vertical grinding mills operated in parallel. The vertical mills will operate in closed circuit with hydrocyclones. Vertical mill discharge will be combined with copper-moly first cleaner flotation concentrate in the regrind cyclone feed sump and will be pumped by variable speed, horizontal centrifugal slurry pump to a dedicated hydrocyclone cluster for each regrind mill. Hydrocyclon underflow will report back to the respective regrind mill. Hydrocyclone overflow (final regrind circuit product), with a target particle size distribution of 80 percent finer than 25 microns, will flow by gravity to the copper second cleaner flotation circuit. Second cleaner concentrate will flow by gravity to the third cleaner feed sump and will be pumped by horizontal centrifugal pumps to the third cleaner columns for upgrading. Tailing from the second cleaner flotation cells will return to the first cleaner flotation circuit. Third cleaner concentrate will flow by gravity to the molybdenum separation circuit. Tailing from the third cleaner flotation columns will return to the second cleaner flotation circuit.

Molybdenite Flotation
Concentrate from the final cleaner of the bulk flotation circuit will report to a copper-moly separation thickener. Thickened copper-moly concentrate will be pumped by variable speed, horizontal centrifugal slurry pumps to the molybdenite (moly) flotation circuit. The moly flotation circuit will consist of two agitated rougher conditioning tanks, one row of separation (rougher) flotation cells, one row of first cleaner flotation cells, a concentrate regrind circuit, one second cleaner flotation column, one third cleaner flotation column, and one fourth cleaner flotation column. Concentrate from the moly rougher cells will be pumped to the moly first cleaner flotation cells. Tailing from the moly rougher cells, which will be the final copper concentrate flotation product, will flow by gravity to the copper concentrate thickener. Concentrate from the moly first cleaner cells will flow by gravity to the moly concentrate regrind circuit. Tailing from the moly first cleaner flotation cells will flow by gravity to the copper concentrate thickener. In order to reduce consumption of NaHS and nitrogen, the rougher and first cleaner cells will be covered, and the flotation gas will be recycled. Moly concentrate regrinding will be performed in a vertical mill, operated in open circuit. First cleaner moly concentrate will flow by gravity to a regrind sump and be pumped by a variable speed, horizontal centrifugal slurry pump to the mill. Reground moly first cleaner concentrate will be pumped to the moly second cleaner flotation column. Concentrate from the moly second cleaner column will flow by gravity to the moly third cleaner flotation column. Tailing from the moly second cleaner flotation column will be pumped to the moly first cleaner flotation circuit. Concentrate from the moly third cleaner column will flow by gravity to the moly fourth cleaner flotation column. Tailing from the moly third cleaner flotation column will be recycled to the moly second cleaner flotation column. Concentrate from the moly fourth cleaner column, which will be the final moly concentrate flotation product, will flow by gravity to the moly concentrate dewatering circuit. Tailing from the moly fourth cleaner column will be recycled to the moly third cleaner flotation column.

OXIDE ORE PROCESS PLANT DESCRIPTION
Heap Leaching
Barren process solution will be applied to mineralized material lots. Solution will be applied with drip emitters to minimize evaporation losses. When a mineralized material lot has completed the primary leach cycle, solution application will be stopped and another mineralized material lift (or layer) will be placed on top of the previous lift. Leach solution application will resume. The process of layering and leaching the mineralized material will repeat for a maximum of eight mineralized material lifts or layers on the leach pad. When the last process leach cycle is completed on the last lift, the mineralized material heaps will be rinsed with fresh water to recover the remaining gold and rinse the residue. Pregnant solution discharging from the mineralized material heaps will be collected in a network of pipe placed throughout the overliner material that will direct the solution to the in-heap collection area. Pregnant solution will be pumped from the in-heap collection area using horizontal, centrifugal pregnant solution pumps. The pump discharge pipes will be combined in a single pipeline to the carbon-in-column (CIC)/SART circuit for recovery of gold and copper.

The process steps required to recover gold and silver by the carbon adsorption method include:
- loading gold and silver on activated carbon in a CIC circuit;
- acid washing of the carbon to remove water scale and acid soluble copper;
- cold stripping of carbon (elution) to remove copper;
- stripping gold and silver from the carbon using a hot caustic solution;
- electrowinning gold and silver from the stripping solution in a precious metal sludge using an electrolytic cell;
- reactivating stripped carbon by thermal regeneration, and
- melting the precious metal sludge in a crucible furnace to produce Doré bars.
The process steps required to recover copper by the SART method include:
- bleeding a portion of the pregnant solution to the SART process;
- adding sodium hydrosulphide to the solution;
- decreasing the pH of the solution with acid, thereby precipitating copper;
- removing the copper precipitate from the solution by thickening, filtration, and drying;
- increasing the pH of the solution with lime, thereby precipitating gypsum;
- removing the gypsum from the solution by thickening, and
- shipping the filtered copper sulphide product to a smelter for refining.

Doré buttons will be the final product of the operation.