The Copper Mountain (CM) deposit is an example of an alkalic porphyry deposit, in which copper–gold mineralization is spatially and genetically associated with multiple pulses of volumetrically restricted, and compositionally varied, alkaline porphyry intrusions.

Mineralization
The bulk of the known copper mineralization at Copper Mountain occurs in a northwesterly trending belt of Nicola Group rocks, approximately 5 km long and 2 km wide, that is bounded in the south by the CMS and in the west side by the northernly trending Boundary Fault system. Here copper mineralization occurs as structurally controlled, multidirectional veins and vein stockworks, with peripheral disseminations. Mineralization had been subdivided into four types, as follows: 1) disseminated and stockwork chalcopyrite, bornite, chalcocite, and pyrite in altered Nicola Group volcanic rocks and LHIC rocks; 2) bornite-chalcopyrite associated with pegmatite-like veins (coarse masses of orthoclase, calcite, and biotite; 3) magnetite-(±hematite)-chalcopyrite replacements and/or veins); and 4) chalcopyrite-bearing magnetite breccias (Fahrni et al., 1976; Klue et al., 2020; Preto, 1972; and Stanley et al., 1995). All mineralization types can be found in each pit area, but each pit is unique with respect to the relative quantities and character of mineralization type.

Disseminated and stockwork chalcopyrite–bornite–chalcocite-(±pyrite) mineralization formed much of the core of the deposit in the southern area of the CM Main Pit (formerly known as Pit 3). In the highestgrade areas that were mined underground, bornite and chalcopyrite veins have locally been replaced on their margins by minor amounts of bladed specular hematite, digenite, chalcocite, and epidote. These replacements are interpreted to be a late hydrothermal overprints on the bornite-chalcopyrite bearing veins, possibly during collapse and cooling of the magmatic hydrothermal system.

Bornite-chalcopyrite mineralization associated with “pegmatite-like” veins are either barren of sulphides or contain chalcopyrite and magnetite with either bornite or, less commonly, pyrite. Historically, these have been subdivided into several different groups, based upon mineralogy: 1) barren veins; 2) bornite–chalcopyrite–(magnetite) bearing veins; and 3) chalcopyrite–pyrite–(magnetite)–bearing veins. Sulphide-absent pegmatite-textured veins occur locally across the mine area. In contrast, the distribution of the sulphide-bearing pegmatite-textured veins defines an “inner” bornite–chalcopyritebearing zone and an “outer” chalcopyrite–pyrite-bearing zone. The inner zone is situated along the northern contact of the CMS within the Nicola Group (in the vicinity of the historical Pits 1 and 3); however, these veins do not penetrate more than 1 m into the CMS (Stanley et al., 1995).

Magnetite–hematite–chalcopyrite replacements and veins primarily existed in the historical Virginia Pit (now known as the CM North zone), and within prospects hosted by the LHIC. They are commonly dilatant, planar, and occur in “sheeted” parallel sets, but in some cases form crackle zones a few metres wide. They are mineralogically similar to the chalcopyrite–pyrite-bearing pegmatite-textured veins, but carry significantly higher gold grades (Stanley and Lang, 1993) relative to other mineralized zones to the east. Chalcopyrite precipitated after pyrite and magnetite within these veins and native gold has been observed associated with chalcopyrite over-grown by pyrite. Gangue calcite, is typically interstitial to magnetite and sulphides.

Magnetite filled breccias with potassically altered clasts of Nicola Group volcanic and LHIC dykes occur within the Ingerbelle Pit and the north side of the CM Main Pit. They occur as elongate bodies along faults, or as roughly circular bodies at fault intersections. These breccias are mineralized with chalcopyrite and pyrite and are bounded by a higher-grade zone of copper–gold mineralization. The breccias include stockworks on their margins, crackle breccias, and clast-supported partially-milled breccias in their cores. The alteration mineralogy of these breccias is closely associated with magnetite–sulphide veins, and are interpreted to be a different structural representation of the hydrothermal event that formed the magnetite–sulphide veins (Stanley et al, 1995).

Copper Mountain Mine (BC) Ltd. (CMML) employs conventional open pit mining methods composed of blasthole drilling, blasting, shovel loading, and rigid-frame, rear-dump truck haulage. Blastholes (270 mm or 311 mm diameter) are drilled on a grid pattern, with blasthole spacing between 7 m and 9 m depending on hole diameter, rock hardness, and whether material is anticipated to be ore or non-economic rock (NER). The blasthole cuttings are mapped and sampled, with samples transported to the on-site analytical laboratory. Samples are pulverized and analyzed for copper. Assays are uploaded to the ore control department and combined with the exploration drill database, which is then interpolated onto bench plans together with blasthole grades and geological information. Grade boundaries are selected manually, and depending on the material, the blasting details are determined. Following blasting, the dig plans are uploaded to the shovels and dispatch system to direct mining and haulage.

Mining at the Copper Mountain Mine (CMM) is by conventional open pit methods, using a 15-m bench height. The major components of this mining method are blasthole drilling, blasting, loading, and hauling. Pit walls are designed to geotechnical specifications and vary in slope angle depending on lithology, alteration, and local ground conditions, with bench face angles of 70 degrees to 73 degrees, and berms of 9 m to 14 m, resulting in inter-ramp pit wall angles that range from 37 degrees to 52 degrees. Ramps are generally 33.5 m wide.

The mining fleet consists of two electric shovels, two diesel shovels, 28 (220-t) haul trucks, five drills, and related ancillary equipment. Blasthole pattern drilling varies depending upon drill diameter, rock hardness, and ore or waste determination, and it is typically a 7 m by 8 m pattern.

All blastholes are analyzed for copper for grade (ore) control purposes and blasthole samples are composited and analyzed for sulphur and carbonate to determine potential for acid rock drainage. Segregation of NER based on acid rock potential is not required at CMM.

Blasthole copper assays are imported into general mine planning, software and interpolated using an inverse distance to a power of two (IDP-2) algorithm method, and search criteria that closely match those used in the Mineral Resource estimation, into 7.5 m by 7.5 m by 15 m blocks. The interpolated block grades are used to design the subsequent dig plans with economic, and non-economic rock boundaries.

Life-of-Mine Production Schedule
Mining is carried out on a 24 h/d, 7 d/week basis, and the mining rate varies from 150 kt/d to 200 kt/d, depending on haulage distances, ore:waste ratios, equipment availabilities, and other factors.

Phase designs and ultimate pits were used as a basis for the schedule. There were 18 phases incorporated from the CM Main and North pits, and 6 phases from the NI pit. All phases used were designed pits that included berms (catch benches), batters, and roads.

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Since 2020 several projects have been completed to debottleneck downstream unit operations. These include Ball Mill 3, rougher expansion, Cleaner Column 3, and Filter Press 2. The completion of these project will allow the concentrator to increase production from 40 kt/d to 45 kt/d while maintaining a primary grind P80 150 µm without sacrificing recovery.

The CMM concentrator flowsheet is a relatively simple two-stage crushing, SAG, pebble crusher, ball milling, and sulphide flotation circuit design. The current capacity supports 45 kt/d of ore processing.

Flotation
The ball mill product is sent to the rougher flotation circuit, where it is treated by two banks of five TK-160 m3 mechanically agitated tank cells. Rougher concentrate is pumped to a 14 ft by 28.5 ft regrind mill for further liberation. In September 2022, two TK-300 m3 mechanically agitated tank cells will be commissioned as further expansion to the rougher flotation circuit. The rougher circuit expansion will provide 50 kt/d capability to the rougher circuit but also allow for increased retention time for difficultto-float ores at lower throughputs.

In August 2018, a new flash flotation circuit was installed to process a portion of regrind cyclone underflow. This circuit comprises an SK240 flash flotation cell followed by a TC5 cleaner cell. The intent of the circuit is to remove coarse liberated material to final concentrate, creating additional rougher mass pull capacity, while preventing unnecessary overgrinding, which historically resulted in gold losses.

Regrind cyclone overflow is pumped to two, 3.7 m diameter by 12 m high column cells for upgrade to final concentrate. In February 2018, these column cells were upgraded with a dynamic sparger system intended to improve fine particle and overall unit metal recovery.

In June 2022, a 6.0 m diameter column 3 began operation to upgrade the first cleaner flotation circuit, processing reground rougher flotation concentrate. This increased the cleaner recovery from 97% to 98% and was sized to handle up to 65 kt/d throughput.

In July 2019, a new second cleaner flotation circuit was installed to process flotation column concentrate. This comprises three 1.9 m direct flotation reactor (DFR) cells. This increased the final concentrate grade from 25% to 28% Cu while maintaining current levels of cleaner circuit performance.

Tailings from these columns are sent to a bank of five TK-70 m3 cleaner-scavenger tank cells. Concentrate from these cells reports back to the regrind cyclone feed pump box for further liberation. Tailings from this bank is joined by rougher tailings and reports to the TMF.

By means of smaller pump upgrades, the installation of the flash circuit, and column cell sparger upgrades, the cleaner circuit capacity has improved, allowing for higher mass pull despite the increased throughput compared to the original flowsheet design.

Gold and silver are recovered as by-products in the final copper concentrate. The historical relationships indicate that recoveries of both metals have a relationship with copper recovery, supporting an association with chalcopyrite.

Concentrate Handling
The dewatering circuit comprises a 16 m diameter, high-rate concentrate thickener, followed by a concentrate storage tank with 24 hours of capacity and a filter press containing 62 plates, each plate measuring 1.5 m by 1.5 m. In July 2022, Filter Press 2, containing 60 plates each measuring 1.5 m by 1.5 m, began operation. With the installation of Ball Mill 3, the primary grind P80 was dropped from 225 µm to 150 µm. The installation of Filter Press 2 was to maintain filtration rates when dealing with finer ore and was also sized to handle up to 65 kt/d throughput.

Filter cake concentrate drops into a storage shed with seven days of storage capacity. Concentrate is loaded into concentrate highway trucks and trucked to the port in Vancouver to be loaded onto vessels bound for Asian smelters.

Tailings Management Facility
The CM TMF is a centre-line, cyclone sand dam construction. The TMF is naturally confined by topography from the north and south, with dams constructed on the east and west extents. This facility started during previous historical operations on the property and was reactivated in 2011. Each dam crest uses a header pipe and four to five cyclones to generate sand for approximate 5 m annual raises.

The CM TMF design, as it is currently permitted, provides sufficient storage capacity to operate until end of 2027. Providing additional tailings storage to operate until 2044 at 45 kt/d.

The 65ktpd Expansion includes a new primary crusher feed hopper, modifications to the primary gyratory crusher, the installation of a High-Pressure Grinding Roll (HPGR) circuit, the addition of a fourth ball mill, a regrind verti-mill, additional rougher and cleaner flotation circuit capacity, and electrical system upgrades. The existing SAG mill will be retired. The fourth ball mill, a 22 ft by 38 ft mill, will be installed adjacent to the third ball mill within the existing building. With the addition of the fourth ball mill, the ball milling line will comprise four mills operating in parallel: two identical 24 ft x 30 ft mills, and two identical 22 ft x 38 ft mills (see Appendix 2 for the proposed 65ktpd process flowsheet). This work will allow for increased throughput with a slightly coarser grind size P80 of 165 µm as compared to the current grind size of 150 µm.

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