The Woodlawn VMS deposit lies on the eastern limb of the asymmetric northnorthwest plunging Woodlawn Syncline. The regional structures trend approximately north-northwest with the local foliation dipping moderately to steeply to the west. The Woodlawn deposit is hosted by regionally metamorphosed (green schist facies) fine to coarse grained felsic to intermediate volcanic rocks, volcanogenic sedimentary rocks and minor carbonaceous shale, known as the Woodlawn Volcanics. The axial plane dips at about 60° to the west and is paralleled by a strong slaty cleavage or in places a schistosity through the mine sequence.

Within the Woodlawn mine area, the Woodlawn Volcanics comprise a lower sequence of coarse-grained silica and sericite altered volcaniclastic quartz-phyric rhyolitic rocks with siltstone, mudstones and minor basalt overlain by strongly silica and sericite altered rhyodacitic to intermediate volcaniclastic mudstones and sandstones that host the massive sulphide lenses. Overlying the mineralised sequence is a less altered massive, coherent, quartz–feldspar phyric rhyolite unit. The Woodlawn Volcanics have been intruded by calc-alkaline and tholeiitic dolerites (McKay & Hazeldene 1987) related to the Currawang Basalt. Dolerites postdate hydrothermal alteration and mineralisation at Woodlawn and are boudinaged (along with mineralisation) indicating that they were emplaced prior to deformation.

The Currawang deposit is conformable to and hosted by sub-vertical sheared spilitic pillowed and massive basalts with basaltic tuffs and fine-grained clastic rocks of the Currawang Basalt. The Cowley Hills deposit is also hosted in the Woodlawn Volcanics. The mineralisation consists of a single massive sulphide lens of average 9-15m true width, 80m strike length conformably hosted in altered hydrothermally altered felsic volcanogenic sediments within dolerite and spilitic volcanics of the Currawang Basalt.

The Woodlawn deposit occurs in Silurian felsic volcanic rocks, volcanogenic sediments and carbonaceous shales intruded by doleritic sills. It is considered to be part of suite of Volcanic-hosted Massive Sulphide (VMS) deposits formed along the eastern coast of Australia in the Paleozoic (240 to 540 Ma) (Gemmell et al 1998). Other significant VMS deposits in this suite include Mt Lyell, Rosebery, and Hellyer in Tasmania, Stockmans in eastern Victoria, Captains Flat in NSW and Mt Morgan and Thalanga in Queensland.

Three mineralised horizons hosting twelve known massive sulphide lenses occur within a 400 by 600m wide and 900m+ long northwest plunging envelope at Woodlawn. Each horizon has distinctive differences in mineralogy, geochemistry, metal ratios, alteration, isotope geochemistry and geophysical characteristics which reflect evolving ore-forming processes. Two major mineralisation types are recognised: a “copper-rich” assemblage and a “polymetallic” assemblage (see McKay & Hazeldene 1987; Glen et al. 1995). The “polymetallic” assemblage includes fine- to medium grained, banded to massive pyrite–sphalerite–galena–chalcopyrite with the gangue mineralogy including talc, quartz, chlorite, phlogopite, muscovite and barite. The copper-rich assemblage includes pyrite–chalcopyrite, lesser pyrrhotite as well as chlorite, quartz and calcite as massive sulphide and stock-work veins.

The deposit geometry has been modified by both ductile and brittle deformation. The ductile deformation is preserved as a prominent foliation with minor rodding concentrated along the margins of dolerite, in fine grained units and phyllosilicate altered zones. The dolerites and mineralised lenses display stretching to the northwest in the plane of foliation and are cut by lens parallel and sub-parallel anastomosing brittle faults.

Mining at the Woodlawn Project will proceed from two independent ore sources, fresh ore from underground and surface reclaimed tailings from the existing tailings dams.

The tailings will be hydraulically mined using a bottom up approach with a channel feeding back to a sump.

The underground is scheduled to produce at an annualised maximum rate of up to 864 ktpa using three different mining methods:
-Drift and half uppers;
-Underhand Longitudinal Open Stoping; and
-Underhand Transverse Open Stoping.

The plant will treat either reclaimed tailing at a rate of 190t/h (1.5Mtpa equivalent) or the underground material at a rate of 125t/h (1.0Mtpa equivalent).

Underground ore will be mined and stockpiled until a minimum of two weeks feed stock is available for campaign ore treatment at an annualised rate of 1Mtpa. The aim is to minimise stockpile time to avoid potential effects of ore oxidation.

Fresh ore from underground will be transported via a decline to the surface, where it will be stockpiled on the ROM pad for treatment. ROM ore will be crushed using a two-stage crushing circuit closed by a screen, and consisting of a primary jaw crusher and a secondary cone crusher. Undersize screen material will be conveyed to a fine ore bin and reclaimed to a ball mill grinding circuit closed with hydrocyclones prior to being feed to the common flotation plant.

The flotation circuit employed will utilise a gangue pre-float, copper, lead and zinc differential flotation sequence. Feed will be either cyclone overflow from the primary ball mill (when processing underground ore) or IsaMill product when processing tailings ore.

Pre-float cleaner concentrate will be discarded to final tails to remove some of the gangue and talc ahead of the differential flotation circuit. A differential flotation circuit for separation of copper, lead and zinc will be utilised with concentrate regrind stages in each circuit to produce saleable mineral concentrates. Tailings from the flotation plant will be thickened for recovery of process water with final tailings deposited into a new tailings dam (TSF4).

Flotation concentrates from the copper, lead and zinc flotation circuits will be thickened and filtered for road transport to port (Port Kembla and / or Port Botany). Concentrate will be loaded into half height containers via front end loader (FEL) at site before transport of the concentrate from site by road. Lead concentrate containers will be plastic lined containers. Rail transport options are also being considered.

Raw water will be supplied to the process plant from the Willeroo borefield. Standing acid waters in existing evaporation and tails dams (approximately 1,000 ML; pH. 3) will be neutralised in a water treatment system and used in the process water circuit.

Final copper, lead and zinc concentrates from the flotation circuits will be pumped to 6m, 6m and 10m diameter high rate concentrate thickeners respectively. Flocculant will be added to increase the settling rate and underflow density to 55-60% solids. Underflow will then be pumped to the associated copper, lead and zinc agitated concentrate storage tanks. Overflow from the each of the concentrate thickeners will gravitate to the water treatment rapid mix tank.

The filtration section will be similar for copper, lead and zinc concentrates and will consist of concentrate storage tanks, filter feed pumps and plate and frame pressure filters. The concentrate storage tanks will provide storage capacity ahead of the concentrate filters. The volumes will be 60 m³, 100 m³ and 200 m³ for copper, lead and zinc respectively. The minimal operational surge time will be 21, 21 and 14 hours respectively.

The concentrate storage shed will be a covered building including individual concentrate storage areas, access and working areas for a FEL, and a drive-thru weigh bridge for concentrate load out. The structure will be sized to accommodate stock piles of 520t, 665t and 1055t for copper, lead and zinc concentrates respectively. During underground ore treatment this equates to 6, 5 and 3 days storage respectively.