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.

Being a large, near-surface orebody, the Josemaría project will be developed as an open pit mining operation. Waste and ore will be drilled and blasted, loaded by hydraulic shovels and loaders, and transported by haul trucks to external waste storage facilities (WSF), long-term stockpiles, or a primary crusher for mineral processing.

During a strategic mine planning study prior to the PFS, SRK established, along with NGEx, that an optimal processing throughput rate was 150 ktpd.

In establishing a mill throughput rate of 150 ktpd, SRK has assumed that large scale mining would be appropriate for the Josemaría project. Consequently, SRK envisages 290-t (320-ton) haul trucks being loaded with 36-m3 (47 yd3) hydraulic front shovels with rock drilled by rotary blasthole drills capable of drilling 381-mm (15-inch) holes. Both the large hydraulic shovels and the rotary blasthole drills are to be electric-powered. The haul trucks have been specified to be autonomous, saving on labour costs while at the same time increasing equipment utilization.

This primary production fleet is to be complemented by 20-m3 (26-yd3) front-end-loaders for operation at the run-of-mine (ROM) pad and occasional truck loading in the mine and by small rotary drills for wall control blasting and supplemental blasting, 5.3-m (17.3-ft) blade track dozers, 4.6-m (15.2-ft) blade wheel dozers, 4.9-m (16-ft) blade graders, 135-t (150-ton) water trucks, and an assortment of smaller backhoes, dozers, utility dump trucks, and maintenance vehicles to support the operations.

SRK envisages electric rotary drills capable of drilling 381-mm (15-inch) holes as the primary production drill. These would be supported by smaller rotary drills capable of drilling 171-mm (6.75-inch) holes for wall control holes and some make-up drilling in-pit.

All rock at Josemaría is to be blasted. SRK has assumed that the supply and loading of explosives will be a contracted service. To this end, cost estimates were received from Orica (Argentina & Bolivia division) for the supply of all facilities, equipment, and personnel. All explosives and accessories were quoted on a unit basis. The cost of the personnel and construction and maintenance of facilities is recovered by a monthly service charge.

SRK has assumed a 50%/50% explosives blend of ammonium nitrate-fuel oil (ANFO) and emulsion. A powder factor of 260 kcal/t was targeted for blasting.

SRK has assumed all waste and most ore loading will be done with three 36-m3 (47-cu.yd.) electric hydraulic front shovels and one 20-m3 (26-cu.yd.) wheel loader. An additional 20-m3 wheel loader will be stationed at the ROM pad to facilitate stockpile reclaim, whether this is tramming medium and high-grade ore or loading trucks at the low-grade stockpile to transfer to the crusher.

SRK has evaluated the adoption of autonomous haulage for the Josemaría project. In consideration of the increased availability, utilization and efficiency of autonomous haulage, there is a significant reduction in the number of trucks required for the haulage fleet. SRK further increased the operating hours per day of loading equipment, with overlapping operators, to allow for near- continuous loading of the trucks. Additional operating costs for the system were offset by tire and labour cost savings.

Roads and dumps will be maintained by a fleet of support equipment including 5.3-m blade track dozers, 4.6-m wheel dozers, and 4.9-m blade motor graders. Additionally, 135-t water trucks will be used to control dust.

The process facilities are designed to treat a nominal rate of 150,000 tpd of sulphide ore and produce copper concentrate. The process design is based on existing technologies and the largest available and proven equipment sizes.

The crushing circuit will consist of two parallel trains of identical primary gyratory crushers (1,000 kW installed power), treating a total of 8,333 tph (dry). ROM material will be delivered to the primary crushers by mine trucks into a single dump pocket with a 500-t live capacity to the top of the primary crusher for each train. The primary crusher will receive ROM (F100 of 1,200 mm) and reduce it to a P80 of 137 mm. The primary crusher will discharge the crushed ore into a 500-t live capacity surge pocket where a variable speed apron feeder will deliver ore from the surge pocket to the coarse ore stockpile feed conveyor. The stockpile feed conveyor will deliver the ore to a coarse ore stockpile for reclaim.

Dual reclaim tunnel systems will be used to house a total of six reclaim apron feeders (three pertrain) that draw ore from the coarse ore stockpile (454,000 wet t) to two coarse ore reclaim conveyors. The coarse ore stockpile will have a total live capacity of 12 hours of mill feed at the nominal feed rate (91,000 wet t). The stockpile material will be reclaimed from the stockpile at an average rate of 3,676 tph (dry) per train to the coarse ore surge bins. A magnet will be installed on each conveyor to remove tramp iron from the ore, as well as a metal detector to indicate any metal not picked up by the magnet.

There will be four secondary cone crushers (933 kW installed power), each of which will operate in an independent line. Each line will consist of a surge bin section, a belt feeder and a standard cone crusher.

Tertiary crushing will be performed by four independent HPGR lines (6,800 kW installed power). Each line will include an HPGR surge bin, a feeder, and a HPGR unit. Each line will share common feed distribution and product conveyors.

There will be four independent grinding lines, each consisting of a fine ore surge bin, two reclaim feeders, two screens, a pump box and cyclone feed pump, a cyclone cluster and a ball mill (7.9 m diameter x 12.5 m effective grinding length with 19,000 kW installed power).

The flotation circuit is designed to receive feed material grading 0.29% Cu and 0.21 g/t Au over the LOM. Nominal copper recovery to the flotation concentrate is 86.4%, with a gold recovery of 70.9%. Nominal copper concentrate grade is expected to be 25.1% Cu with 14.5 g/t Au.

Flotation comprises rougher scavenger flotation, regrinding of the rougher concentrate and three stages of cleaner flotation to produce a final copper concentrate.

Copper flotation will be promoted using Potassium Amyl Xanthate (PAX), Sascol 95 and Matcol TC-123 collectors with Methyl Isobutyl Carbinol (MIBC) added as a frother. Lime will be used for pH modification, as required.

The final copper concentrate will be thickened with flocculant to 65% solids in a high-rate thickener (35-m diameter). The thickened concentrate will feed the copper filtration stock tank. Thickener overflow water will be recycled to the process.

An agitated tank will have capacity to store six hours production of thickened concentrate ahead of the filtration stage. One pressure filter (61 plates, 2.5 m by 2.5 m plates) will reduce moisture content of the concentrate to approximately 9% for shipment. Filtered concentrate will fall into a covered storage area and will be transferred to a covered copper concentrate stockpile area by front-end loader. There will also be an open emergency storage area in the event of shipment disruptions.

A front-end loader will be used to load concentrate from the storage area to copper concentrate trucks. Each concentrate truck will be weighed and sampled to ensure maximum filling without overloading.