Based on geological features and location, the Josemaría and Los Helados deposits are classified as examples of Cu–Au porphyry systems. Porphyries are well documented along the Andes and represent a widespread type of deposit in Chile.

Mineral zones within the Josemaría deposit were defined by the relative abundance of chalcopyrite, pyrite and chalcocite, as well as the mode of occurrence of chalcocite (hypogene or supergene), and the level of oxidation. Six main zones were modeled:
• Cu oxide (CuOx)
• Barren oxide (leached cap) (Ox)
• Pyrite + hypogene chalcocite (PyCc(H))
• Pyrite + supergene chalcocite (PyCc(S))
• Mixed sulphide and oxide (MIX)
• Pyrite + chalcopyrite (PyCpy).

The Cu–Au mineralization at Josemaría is mostly hosted by a Miocene porphyry system which forms an elongated body with minimum dimensions of 800 to 900 m north-south, 600 to 700 m east- west and 600 to 700 m vertically.

The main hypogene Cu and Au mineralization is associated with the upper parts of the potassic alteration zone. Chalcopyrite and pyrite are disseminated through the potassic zone, with minor bornite. Quartz–magnetite ± chalcopyrite veining occurs through much of the main mineralized zone, as discrete veins and locally as a more intense stockwork.

Hypogene chalcocite mineralization is found predominantly along the western part of the system, associated with the roots of the advanced argillic domain. The occurrence of both hypogene and supergene chalcocite in close proximity complicates the differentiation of these two zones, a situation that is compounded by the fact that they are likely in local gradational contact.

A well-developed leached cap overlies the entire Josemaría deposit, related to oxidation at and below the present topographical surface. The leached zone ranges from 10–20 m in thickness over the relatively impermeable felsic volcanic rocks in the west to a maximum of 230 m into the Josemaría structural corridor and the tonalite farther east.

Mineralization at Josemaría would be mined using conventional open pit methods.

Based on the discontinuity data, conceptual level kinematic analyses were undertaken which indicate that, assuming a maximum bench height of 15 m, bench face angles of 70° are achievable for the west wall. However, analysis of discontinuity data suggests that the east wall stability will be affected by the formation of wedges along structures. The geotechnical study recommended that bench faces on the east wall be restricted to bench face angles of 65°, with minimum bench widths of 8 m and 8.5 m for the east and west walls, respectively. Bench face angles of 70° can be utilized for the east wall to simplify PEA designs, as long as additional catch bench capacity (minimum of 10 m) is incorporated.

A geotechnical catch bench of 20 m wide is recommended every 105 m, where no ramp section is present to reduce the overall slope.

For unconsolidated or partially consolidated materials (overburden, gravel, soil, etc.), a bench face angle of 38° should be considered. A 20 m geotechnical berm should be placed between the rock and the overburden contact.

For waste rock facilities, a maximum bench face angle of 38° is recommended and should have a minimum set back of 50 m from the start of the pit crest in the overburden. Maximum lift heights of 150 m are recommended with 50 m berms.

The final pit design incorporates two main ramps with an exit point at the 4400 level on the north side of the pit; the exit point will be in close proximity to the planned primary crusher location.

Mining loss and dilution consists of two types:
- Lateral.
- Vertical.

The quantity of lost and diluting material in the lateral case is determined by the angle of the dig face (70°), which defines a lateral influence of 2.7 m. The assumption for bench elevation control for the vertical case was 0.5 m total or 0.25 m on average in either direction.

Overall loss and dilution was estimated to be less than 1% and has been incorporated into the block model.

The metallurgical testwork programs reviewed two alternative comminution technologies:
- Conventional semi-autogenous grind (SAG) milling.
- High pressure grinding rolls (HPGR).

While both methods appeared to be effective, based on the testwork completed, a decision was made to proceed with HPGR technology due to the estimated significant LOM operating cost savings.

During the first seven years of operation, the plant will only treat material from Josemaría. The Josemaría open pit will be located approximately 4 km from the proposed process plant site. In year 8, mill feed material from the underground operation at Los Helados will be introduced to the plant, and blended with Josemaría open pit material. The blended feed will continue for a six-year period while Los Helados ramps-up to full production. In year 14 of operations, the open pit mining at Josemaría ceases, and for the remaining mine life, only underground feed from Los Helados will be processed.

Mill feed from the Josemaría open pit and Los Helados underground deposit will be conveyed to the process plant facility. Four conveyors will be constructed, not including the tunnel conveyor from Los Helados to the surface:
- JM1: 1.36 km.
- JM2: 1.57 km.
- JM3: 1.99 km.
- LH1: 2.80 km.

For the Josemaría mill feed material, run-of- mine (ROM) material will be trucked from inside the pit to two gyratory jaw crushers that will be located just outside of the Josemaría open pit. The material will be primary crushed and then conveyed to a common primary crushed stockpile located at the plant via the conveying system (JM1, JM2 and JM3).

Primary crushed material from the primary crushed ore stockpile will be control-fed to conveyors and then to a conventional secondary screening circuit where the material will be combined with secondary cone crushed product. The combined new feed and secondary crushed product will then be fed to a storage bin, and control-fed to six conventional vibrating screens. Screen undersize material will be conveyed to a storage bin feeding the HPGR crushers, whilst screen oversize material will be fed to the secondary crushing. The undersize material will be combined with screen oversize product from the HPGR crushers. This combined material will be control-fed individually to four separate HPGR crushers with the product being conveyed to eight vibrating discharge screens located at the discharge of four ball mills (two screens per mill) operating in parallel. Ball mill discharge screen undersize material will flow into the ball mill discharge hopper where it will be pumped to a bank of hydro-cyclones for classification. Cyclone underflow will flow by gravity to a distribution box and be split evenly to the ball mills, whilst cyclone overflow will be directed to the copper sulphide flotation area.

Conventional sulphide flotation will follow the comminution stage. This circuit will commence with a rougher flotation section. Rougher flotation concentrate will be produced and directed to a vertical mill regrind stage in closed circuit with hydrocyclones. Overflow pulp from the hydrocyclones will be directed to three sequential stages of conventional cleaning flotation. Cleaner scavenger flotation will be installed on the discharge from the first cleaning stage.

Tailings from the rougher flotation and cleaner scavenger flotation stages will be combined and will report to the final tails thickener, where 74% of the water will be recovered. The tails will go to the tailings storage facility where approximately 20% of the contained water in the tailings will be recovered and sent back to the process plant.

Copper concentrate from the third cleaner stage will be directed to the concentrate thickener and filtration stages. The water obtained from the concentrate thickener, tailings thickener and concentrate filter will also be recovered and sent back to the process plant.