The Cannington mineralisation is hosted within a 70 m to 90 m thick of interbedded garnetiferous psammite and basic volcanic units that enclose a fine to medium grained central amphibolite body up to 100 m thick. The psammite is itself intercalated with muscovite-sillimanite schist and enclosed by migmatitic quartzon feldspathic gneiss. Pegmatite horizons that are semi-conformable within the deposit sequence are thought to be predominantly the result of partial melts of clastic rocks within this package (Mark, 1993).

The psammite sequence is folded into a tight, northerly striking isoclinal recumbent fold that plunges to the south with limbs dipping between 40O and 70O to the east as the result of the second of four structural events (D1 to D4) described by Bailey (1998). Two major northwest trending D4 faults offset the sequence. The Trepell Fault divides the ‘Northern Zone from the ‘Southern Zone’ and the Hamilton Fault forms a southern limit of the Southern Zone’. This Proterozoic host sequence is truncated up-plunge in then Northern Zone at the sub-horizontal unconformity with the younger Cretaceous sediments that are between 10 m and 60 m thick in the Cannington Operations area.

The Southern Zone has a strike length of 600 m to 700 m and extends to approximately 650 m below surface. The metamorphosed sediment package has been folded into a synform structure and is displaced on a local scale by north northeast and northeast trending faults. The northeast trending faults are interpreted as conjugate sets to the Trepell and Hamilton Faults as a result of sub- horizontal sinistral movement.

The Northern Zone strike is approximately 500 m and extends to around 400 m below surface. It is less deformed than the Southern Zone and typically dips east, interpreted as being the eastern limb of an antiform.

Silver, lead and zinc mineralisation is stratiform within the garnetiferous psammite and mafic host rocks and varies in textural, mineralogical and geochemical characteristics that have been classified as ‘mineralisation types’. These are divided into footwall (western limb), hangingwall (eastern limb) and fold hinge lead-rich and zinc-rich domains, with the mafic rocks more complex in mineralogy than the siliceous rocks as mineralisation grades from outer Pb-Ag to inner zinc zones. The sulphide mineralogy is dominated by coarse grained galena (PbS), sphalerite (ZnS), with silver mainly contained in freibergite (Cu6(Ag,Fe)6Sb4S13).

Both open-stope and bench mining methods are used at Cannington, with approximately 3.3Mt of ore processed each year. The ore travels to the surface from depths of up to 620 metres (m) via two methods; either by a nine tonne capacity vertical shaft hoisting system or by 50t haul trucks. The ore is then stockpiled for processing.

The Cannington deposit is characterised by competent ore with much weaker host rock outside of the orebody. Much of the mine access, ore handling and service infrastructure is located in the amphibolite core of the deposit or within the ore to avoid poor ground conditions. The main orebody (Southern Zone) is separated from the satellite North Zone by a major regional structure, the Trepell Fault. A number of other significant structures impact the mining operations. These are managed primarily by leaving a skin of ore between stopes adjacent to the structure. Mining exclusion zones have been delineated around the crusher, shaft and access infrastructure. Provision has been made in the mine plan for a replacement crusher to allow extraction of high-grade ore within the crusher exclusion zone in 2018.

The processes for extracting the silver, lead and zinc from the ore involve grinding, sequential flotation and leaching techniques that produce high-grade, marketable lead and zinc concentrates with a high silver content.

Ore is crushed in an underground double-toggle jaw crushing facility with the crushed ore hoisted to surface and trucked to an open, conical stockpile. In addition, some underground ore is trammed to surface using LHD units, where it is crushed and then placed on the same stockpile. The crushed ore is moved to the grinding feed stockpile using a front-end loader (“FEL”). The grinding circuit comprises a single 5.8 MW autogenous grinding (“AG”) mill at 8.53 m diameter and 3.96m effective grinding length. The AG mill operates in closed circuit with a pebble crushing circuit and primary hydrocyclones.

The hydrocyclone overflow product at a particle size distribution of 80 per cent passing around 110 microns (P80 = 110µm) is treated in two vertical ball mills (“Vertimills”) operating in closed circuit with a second bank of hydrocyclones. The overflow from this circuit is treated in a talc pre-flotation step to remove talc and a significant amount of fluorine. The concentrate from this pre-float step is disposed as a final tailing product. The floatation tailings is fed to another stage of hydrocycloning that allows the fines overflow to be treated in a bank of rougher flotation cells, while the underflow is treated in another bank of flotation cells in what is termed a “split” flotation circuit.

The fines portion is floated firstly in a rougher bank with rougher tailings joining the coarse stream feed. The fines rougher concentrate is cleaned with lead (“Pb”) concentrate sent directly to the Pb concentrate thickener. Cleaner tailings are recycled to the rougher circuit. The hydrocyclone coarse underflow along with the fines rougher tailings is treated in a coarse rougher flotation stage.

This rougher concentrate is sent to regrind mill before being treated in a Pb cleaning flotation bank. The cleaner Pb concentrate is delivered to the Pb concentrate thickener while the cleaner tails are recycled to the Pb rougher circuit. The Pb rougher tailings are sent to the zinc (“Zn”) flotation circuit.

The Zn rougher concentrate is reground in a Vertimill prior to being floated in a cleaner bank of cells with cleaner concentrate sent to the Zn thickener. The tailings from cleaners and from roughers are treated in a scavenger bank of flotation cells with scavenger concentrate recycled to the rougher flotation bank with scavenger tailings sent to a tailings thickener as the final tailings product.

The thickened Pb concentrates are pumped to an agitated storage tank where it is leached using aluminium sulphate to reduce the fluorine (“F”) content to a level acceptable to smelters. Once leached the Pb concentrate is filtered using a horizontal belt filter. The filtrate from the filter is disposed in a separate fluorine storage pond adjacent to the TSF. The resultant filter cake at about 7.5 per cent moisture is shipped to Townsville port and then on-shipped to the market smelters. The same process is applied to the Zn concentrates and the final filter cake ready for market runs about 9.5 per cent moisture.

The tailings thickener underflow is pumped to a past-fill station where much of the water is extracted and the dense solids are mixed with cement powder and then pumped to underground sites as backfill for mined out voids. The tailings not used as backfill is pumped to the TSF.

The tailings thickener overflow is recycled to the plant, while the Pb concentrate thickener overflow is recycled to the Pb flotation circuit and the Zn thickener overflow is recycled to the Zn flotation circuit.

The Ore Treatment Plant commenced operation at a rated throughput of 1.5 Mtpa. Over the years, the throughput capacity has increased to 3.2 Mtpa and is expected to increase to 3.4 Mtpa.