The Feasibility Study created the process flow sheet consisting of two stages of primary grinding, rougher flotation, rougher concentrate regrinding, first stage cleaning followed by cleaner-scavenger flotation, and two additional stages of cleaner flotation. The cleaner-scavenger concentrate will be returned for regrinding. The rougher tailings and the cleaner-scavenger tailings will be combined and sent to the tailings facility for cycloned sand placement. Test results showed that responses to commonly used flotation reagents will be acceptable.
Gravity test work to optimize recovery of gold to the final concentrate showed considerable potential. The degree of the economic advantage for operating a gravity concentration circuit could not be identified in the test work. However, based on successes at other porphyry copper gold operations, gravity concentration circuits will be included in the primary and regrind circuits. Mineralogical analysis and subsequent test work has also shown a strong potential benefit for flash flotation and as such a simple two cell circuit has been added to this end.
Copper recovery was correlated with the copper head grade for a constant copper concentrate grade. Recoveries were normalized on a 28% Cu concentrate grade for hypogene, on a 26% Cu concentrate grade for mesogene and 58% Cu concentrate for supergene. In all cases results from batch open circuit data were offset upwards by 2% -3% to project plant performance. This was supported by results from locked cycle tests. Overall copper recovery from the supergene ore was variable in the samples tested due to the variable proportions of native copper. Native copper is not readily recovered by flotation but can be recovered by gravity concentration.
Mineralogical studies showed that gold was primarily associated with copper sulphides recoverable by flotation. Only about 4% of the gold was associated with native copper in supergene ore. For hypogene ore, the gold recovery did not correlate well with copper recovery but the gold concentration ratio correlated well with copper concentration ratio. Therefore gold recovery was calculated using the concentration ratio correlation and the calculated concentrate production based on copper.
The mill is designed to process 11,000 tonnes per day of ore at full capacity and will operate 24 hours per day; 365 days per year with two observed holidays and scheduled downtime for maintenance of equipment. Initially planned for two stage ramp up, the current plan allows for full commissioning to design production rate during initial construction. The process utilizes the two stage SAG-Ball mill combination to produce a constant product size to the flotation circuit. A slurry stream will be tapped from the cyclone underflow to feed the gravity concentrator and flash flotation circuit to maximize gold recovery. Gravity and flash concentrates will be directed to the concentrate thickener.
Primary and secondary flotation cells will be installed. The primary cyclone overflow will be directed to the flotation circuit. All flotation separations will be conducted at the natural pH of the ore. Rougher concentrates will be reground in a vertical grinding mill operating in closed circuit with cyclones and upgraded to a final concentrate grade with three stages of cleaner and one stage of cleaner-scavenger flotation. The second cleaner concentrate will feed the third cleaner circuit to produce the final copper concentrate.
The concentrate dewatering circuit will involve thickening and pressure filtration. The water removed from the concentrate will be recycled within the process plant. A 40 square meter pressure filter will produce an 8-10% moisture concentrate that will be sampled, loaded into B-Line trucks and weighed before it is hauled to the rail transfer at Ashcroft and then sent by rail to Vancouver. The majority of the concentrate is expected to be sent from Vancouver via ocean freight to Asian smelters.
The Afton Open pit presently contains about 2.7 million cubic meters of water. This will be pumped to, and stored in, Pothook pit during project development. The capacity of Pothook will be increased to 2.9 million cubic meters by constructing a 12 meter high earth dam on its south west side. In the initial stage of the project, whole tailings will be pumped into Pothook, an equal volume of water will be displaced and pumped back to the mill. When Pothook is full of tailings, approximately coincident with the ramp up to 4 million tonnes per year, tailings will be pumped to the main Tailings Storage Facility (TSF) to be constructed to the south west of Pothook, and east of the existing TSF.
The tailings storage concept is to construct a centre line raised tailings dam using native glacial till and cycloned sand from the flotation tailings to provide an impoundment area. Flotation tailings will be pumped to a non-elaborate cyclone plant where a primary cyclone separation will occur. Diluted underflow from the primary cyclones will be sent to mobile secondary cyclone headers placed around the dam. Underflow from the secondary cyclones will constitute the downstream sand material while both of the fine overflow streams from these two stages will be placed inside the TSF.
Water will be reclaimed via a pump barge and pipeline to be reused in the mill. Centreline raising combined with cycloning was chosen to provide a balance between cost and stability issues. It is also the method used at the nearby Highland Valley and Kemess operations
