Kemess East is a copper-gold-silver-molybdenum porphyry deposit and is typical of calc-alkaline porphyry copper-gold deposits in the western cordillera. The deposit is deeply buried and mineralization starts at an average depth of 900 m below surface and extends to 1500 m below surface. Unlike Kemess Underground, there is no significant low grade mineralization associated with Kemess East.

Copper-Gold mineralization forms a relatively flat-lying zone within the Kemess East deposit.

The phyllic zone consists of pyrite associated with sericite-chlorite alteration. The phyllic zone is barren of significant copper and gold but spikes do occur due to remobilization of mineralization from fluids at depth. Pyrite within the phyllic zone ranges from 3-5% and is dissemined and veined. Minor quartz veining is present within the phyllic zone with pyrite +/- chalcopyrite.

The potassic zone contains a high percentage of the copper-gold mineralization with an upper zone of molybdenum mineralization. Molybdenum is present in the transition zone from phyllic to potassic alteration and is present with chalcopyrite and pyrite within quartz veins, late stage zeolite and carbonate veins, and within joints. The main copper-gold mineralization is present within the Kemess East pluton and is associated with chalcopyrite and pyrite finely disseminated within the host pluton. To a lesser degree, copper-gold mineralization is associated with quartz veining with chalcopyrite and pyrite.

The most appropriate mining methods for Kemess East (KE) deposit are block caving, sub-level caving, and sub-level stoping. Based on mining costs, deposit grade, geometry, and depth, block caving was selected as the preferred mining method.

A footprint at elevation at 370 m has the most value, and the cave design was based on this footprint. Several cave clipping boundaries were trialled before determining the final shape that produced the most value. A cylinder with a 150 m diameter was used to rationalize a realistic caving boundary while extracting some non-economic material in order to include the narrow arms at the northern and southern extremes of the footprint.

The area of the footprint is 107,000 m2 and the perimeter is 1.4 km long. The widest span of the footprint is 470 m.

The underground mine at KE would begin production as the KUG mine production is decreasing, continuing to feed the process plant. The crushed material would be conveyed up from the underground crusher and then connect to the conveyor in the KUG decline, where it would use existing KUG infrastructure to be conveyed to the surface process plant.

Access for personnel, equipment, and supplies to the KE underground would be through a decline collared off the Kemess UG (KUG) ventilation decline. A parallel conveyor decline would be driven for material handling from above the KUG conveyor decline to the KE underground. The approximate location of the connections between the KE twin declines and the KUG triple declines is 2,700 m from the KUG portals (1,150 m elevation, slightly higher than the connection to the KUG workings).

The underground water management system at the KE underground would be designed to handle 1.5 m3/s (24,000 gallons per minute). This accommodates both the groundwater and peak surface water inflows during a 1-in-200-year storm event.

The development stage of the KE mine is approximately five years, which includes mine access, initial footprint development, and construction of major mine infrastructure, such as underground crusher, material handling system, workshop, dewatering system, and primary ventilation system. Following the development stage, the initial ramp-up period is 3 years to reach steady state production of 10.9 million tonnes per year (Mt/yr; 30 kt/d) for 6 years, followed by ramp-down production for another 3 years, for a total production operating life of 12 years.

Primary crushed material from the KE underground mine would be conveyed to the coarse ore stockpile prior to processing in the KE process plant. The material would be reclaimed by feeders below the coarse ore stockpile and conveyed to the grinding circuit. The grinding circuit would be composed of a SAG mill (34 ft diameter by 15 ft 3 inches long; 12,000 hp) with a discharge size of about 2,800 microns. The grinding work index for the SAG mill varies from 9.3 to 10.3 kWh/t. The SAG mill discharge is directed to a vibrating screen with the screen oversize returned to the SAG mill and the screen undersize sent to cyclones. The cyclone underflow is directed to a ball mill (22 ft diameter by 36 ft 6 inches long; 12,000 hp) for grinding operating in closed circuit with the cyclones. The ball mill work index for the KE deposit is similar to that used for the design of KUG. The cyclone overflow represents the final product size at 150 microns K80 to rougher flotation.

The process plant for processing the Kemess East (KE) deposit would be composed of conventional processing circuits for the production of a copper concentrate containing gold and silver as by-products. The existing KUG process plant would be expanded from 25,000 tpd to 30,000 tpd for processing the KE deposit. The results of the metallurgical test programs on the KE sample composites indicate that the same KUG processing circuits would be suitable for KE.

The processing circuits would be composed of the following:
- coarse stockpile with reclaim feeders
- SAG ball mill combination
- rougher flotation
- rougher concentrate regrind
- cleaner flotation
- copper concentrate thickening
- copper concentrate filtering.

Modifications and equipment additions to the KUG process plant for KE would be primarily for the grinding and flotation circuits. The currently planned work programs for the KUG process plant are focused on re-conditioning, repairs and replacement of existing equipment, and the installation new regrind IsaMills™.

Rougher flotation at 36% solids would be done in mechanical cells. Chemicals and reagents added during flotation would include lime for pH control, frothers, and collectors. The tailings from rougher flotation would be pumped to a cyclone for final tailings disposal. The rougher concentrate at about a 13% mass pull would be pumped to cyclones in the regrind circuit ahead of cleaner flotation. The cyclone underflow would operate in closed circuit with the regrind mills. The cyclone overflow would be sent to the first flotation cleaners. The tailings from the first cleaners would be processed in cleaner scavengers, with the cleaner scavenger tailings pumped to the Kemess South Pit (KUG TSF). The concentrate from the cleaner scavengers would be returned to the regrind circuit. The concentrate from the first cleaners would be pumped to the second cleaners (column cells). The concentrate from the second cleaners would be the final copper concentrate and would be pumped to the concentrate thickener. The concentrate thickener underflow at 65% solids would be filtered to approximately 8% moisture in horizontal pressure filters before being shipped to smelters.

There is a potential for installing a Knelson gravity concentrator in the cleaner circuit to recover gold from the cleaner tailings.