Pegmont is enveloped by a variable but large volume of late or postorogenic fracture/vein- controlled alteration. Williams et al identified two stages of alteration. Stage 1 veins typically are composed of quartz ± K feldspar ± tourmaline ± biotite ± rutile ± ilmenite associated with K feldspar ± muscovite ± tourmaline, or biotite ± muscovite ± garnet alteration. Stage 2 quartz ± chlorite ± calcite ± ferroan dolomite ± hematite ± sulphide veins are associated with fine-grained muscovite (illite-phengite) + chlorite ± carbonate alteration.

A largest amphibolite (hornblende-plagioclase) body is poorly exposed in sub-crop on the edge of the mining lease, called the “Lease Amphibolite”, it is a flay lying, highly continuous body that transgresses the BIF without any apparent displacement. The Lease Amphibolite displays textural zoning, which may reflect chilled margins and compositional layering. The southeast edge of Zone 3 is defined where the amphibolite intersects the Lens B BIF, to the southeast of this the Lease Amphibolite is below the BIF's and to the northwest it is above the BIF's. Drill data density in the Lease Amphibolite is less than the mineralization.

The Squirrel Hills Granite, part of the Williams Batholith, is over 100 km north-south and up to 25 km wide. It has intruded the project sequence in the northern boundary of the Pegmont lease. Present as several low outcrops of tors exhibiting spheroidal and onion skin weathering. The southern contact is inferred from airborne magnetics. The Squirrel Hills Granite is a non-foliated, porphyritic granite, composed of biotite and hornblende with coarse K feldspar. The Yellow Water Hole Granite, also part of the Williams Batholith, approximately 18 km from Pegmont, is dated at 1493±8 Ma, postdating main tectonism that affected the sequence.
Lead-zinc mineralization at Pegmont is contained within BIF's. The BIF’s consists of banded quartzmagnetite-fayalite-garnet- grunerite-hedenbergite-sulphide. Apatite, gahnite, and graphite are common minor minerals. Bedding is typically on a scale of 1 to 5 mm. In fresh rocks, the main sulphide minerals are galena, sphalerite, with subordinate pyrrhotite, pyrite and chalcopyrite. In contrast to the almandine garnet in the hangingwall and footwall, the garnet in the BIF is manganiferous spessartite, reflecting the higher whole rock MnO content of the BIF.

The overall morphology of the stratiform mineralized banded iron formation at Pegmont is a flat, gently easterly dipping sheet. The known mineralization extends approximately 2 km along strike and approximately 1 km in the down dip direction to the southwest. Mineralization is known to extend to a depth of 350 m below surface but remains open down dip.

The principal lead and zinc mineralized BIF is termed Lens B, it is around 2 to 8 m thick. At surface, the Lens B BIF is present in banded ferruginous, jaspery, manganiferous gossans. Outcrops occur in two areas termed the Mount Lucas Load and Burke Hinge Zone (BHZ). Additional oxide mineralization has been intersected at the Bonanza Lode, which is about 2 km to the northeast of Mount Lucas. Work by Scott and Jones shows that there has been considerable depletion of Mg, Ca, Na, K, S, Ag, Cd, and Zn during weathering. The least mobile elements have undergone residual concentration up to the profile, whereas concentrations of other elements are either unaffected by weathering or vary irregularly, in many cases reflecting local variations within the ore horizon. On oxidation, galena changes either through a pyromorphite ± cerussite assemblage to plumbogummite / corkite or directly to coronadite.

Oxidation reaches depths of 25 m below surface. The surface gossans display delicate boxworks after fayalitic olivine and are composed of goethite, clay, and secondary lead minerals such as pyromorphite, plumbogummite, plumbojarosite, and / or beaverite, together with isomorphous hydrated lead, iron, and copper sulphates. No secondary zinc minerals have been observed. Between 0.5 and 2% graphite has been noted. A yellow-green fibrous vein mineral contained in the gossans is thought to be nontronite.

The lead content of outcropping gossan is the same as, or slightly higher than that of un- weathered ore. Although zinc is strongly depleted at surface (×10), its greater abundance (×2 depletion) only 15 m below the surface reflects the truncated profile and immaturity of the gossan.

In the partially oxidized transition zone, between 25 – 40 m below the surface, galena, sphalerite, magnetite (invariably showing alteration to hematite and goethite), manganese minerals including pyrolysite, and small amounts of pyrite and graphite have been described along with montmorillonite and kaolin, garnet, mica and quartz, and minor amounts of siderite, amphibole, chlorite, and talc.

The un-weathered Lens B BIF consist mainly of galena and sphalerite associated with a finely laminated assemblage consisting of dominant magnetite and spessartite with subordinate iron-magnesium-manganese silicates and apatite. Gangue minerals are apatite, olivine (fayalite), garnet (spessartine – almandine), amphibole (hornblende and grunerite), clinopyrozene, biotite, and greenalite.

Approximately 10 to 40 m below Lens B is a second BIF horizon, Lens C, generally only weakly mineralized and when present is around 1 m thick and is separated from the Main BIF by garnet-bearing quartzite and schist. Lens C becomes more important in Zone 5 where it is present at thickness of around 4 m.

In Zone 5 there is indications of a further four mineralized BIFs below Lens C, to date they have been intersected on the south western most drill section, in an interpreted anticline position.

The Pegmont project includes three resources consisting of the BHZ, the Main Zones (Zones 1- 5), and the Bridge Zone. The BHZ can be developed as an open pit, the Main Zone can be developed as a combination of open pit and underground, and the Bridge Zone is an underground resource.

AMC (AMC Mining Consultants (Canada) Ltd) developed the Mineral Resource block models for the Main Zone and BHZ for evaluation of the open pit mining potential. The block models are proportional models, containing a field for ore percent and waste percent, with block dimensions of 5 m in the X (east) direction by 5 m in the Y (north) direction by 5 m in the Z (vertical) direction for the Main Zone, and 2.5 m in X by 2 m in Y by 2.5 m in Z direction for the BHZ.

AMC proposes to mine the open pits using a conventional truck and excavator mining method. AMC has assumed that a 10 m bench height would be adopted for waste rock, and that mineralized material will be mined in 2.5 m flitches to increase mining selectivity. Mining of mineralized material will occur using a hydraulic excavator in backhoe configuration (example: Komatsu PC1250). Hydraulic excavators in front-shovel configuration (example: Komatsu PC2000) will be used in areas of bulk waste rock to increase the rate of mining in waste rock and reduce operating costs.

The average productivity of the PC1250 excavators in blasted mineralized material is expected to be approximately 1,080 t/op hr. The productivity of the PC2000 in fresh waste is expected to reach 1,513 t/op hr.

Hauling of mineralized material and waste will be undertaken by 90 t trucks (example: CAT 777).

Pits were designed based on the selected optimization shells. Four pits have been designed, one for BHZ and three in the Main area.

Haulage ramps will be 23.1 m wide at a 10% gradient. Single way ramps of 13.3 m width were designed for the bottom twenty vertical meters.

The Main 7 pit is the largest pit measuring approximately 930 m in length, 330 to 550 m in width, and 200 m at its deepest point. The Main 7 pit consists of four stages (Main 4, Main 5, Main 6, and Main 7).

To assess the underground mining potential, a pit shell was generated that incorporates the potential open pit reserves for the Main Zone, all mineralized material outside of the pit shell was considered for underground mining with the focus on the Bridge Zone.

The geometry of the underground resources is primarily flat dipping (23° to 30°) and varying in thickness across each zone (3 m to 12 m). However, the Zone 5 mineralization and part of Zone 3B tends to be narrow in width and fairly vertical in dip.

The geometry lends itself to room and pillar (R&P) mining in the flat dipping zones and longhole stoping in the steeply dipping zones.

It is assumed that the underground zones will each be accessed from the Main pit with individual declines. AMC has developed conceptual mine designs for each zone to determine mine physicals and development costs.

The three main access declines have a total length of 2,166 m, level access development has a total length of 1,786 m and the remaining development includes return air drive access and mineralized development. Vertical development consists of return air raises with a total of 698 m. Total development for the three underground zones including vertical development is 5,547 m.

Stope wireframes were generated using MSO, a check was made to remove any outlying stopes that would not be economic when the cost of access development was included. The cost of access development was determined for each level and each level was evaluated to determine whether it was economic to develop.

TA conventional sequential flotation circuit has been selected for the recovery of the lead and zinc minerals from the Pegmont deposit.

The process plant shall consist of a conventional three stage crushing and a single stage ball mill grinding circuit, followed by differential flotation of the lead and zinc minerals to produce separate saleable lead and zinc concentrates. The concentrates from the lead and zinc flotation circuits will be thickened and subsequently filtered on site for road transport.

Fresh open pit mineralization will be mined and delivered to a Run of Mine (ROM) pad for processing through a three-stage crushing plant designed to provide a crushed product size suitable as feed to the single stage ball mill.

Crushed material will be stored in the fine product bin for reclaim via a conveyor system to the primary ball mill. The single stage ball mill will operate in a closed circuit with hydrocylones to produce a target P80 size of 106 µm, suitable for the downstream flotation circuit. A differential flotation circuit for lead and zinc will be utilized with rougher concentrate regrind stages for both the lead and zinc circuits to produce saleable grade lead and zinc concentrate.

Open pit material will be reclaimed by a FEL from the ROM stockpile into the ROM bin. The open pit material will be reclaimed from the ROM bin by an apron feeder and vibratory grizzly pan feeder to remove fines from the feed to the primary jaw crusher. The open pit material will be reduced in size from an F80 of 317 mm to a P80 of 120 mm by the primary jaw crushing, based on a close side setting (CSS) of 100 mm.

Crushed material will be milled in a single stage ball mill operating in closed circuit with hydrocyclones. Crushed material will be reclaimed from the fine product bin by a variable speed conveyor for mill feed rate control. The single stage ball mill, with dimensions of 5.23 m Ø x 6.67 m EGL, will be equipped with a fixed speed 3300 kW mill motor. The cluster of 400 mm diameter hydrocyclones will have an overflow size P80 of 106 µm and will be designed to operate at a recirculating load of up to 300%.

Lead flotation circuit.
Cyclone overflow from the grinding circuit will gravitate via a trash screen to the lead rougher conditioning tank. The lead flotation circuit will consist of a conditioning stage, roughing and scavenger flotation, regrinding of the lead rougher and scavenger concentrate, followed by two stages of cleaning. The final lead concentrate will be pumped to the lead concentrate dewatering circuit and the lead rougher flotation tailings and lead cleaner tailings will be pumped to the zinc flotation circuit.

Slurry will gravitate from the lead rougher conditioning tank to the feed box of the lead rougher flotation cells. Flotation will take place in a conventional lead rougher, cleaner and recleaner flotation circuit using mechanically agitated, forced air cells. The lead flotation circuit will consist of the following equipment:

• One 25 m3 rougher conditioning tank.
• Two 30 m³ lead rougher mechanically agitated, forced air tank cells in series.
• Three 30 m³ lead scavenger mechanically agitated, forced air tank cells in series.
• One 355 kW stirred media detritor (SMD) lead regrind mill.
• One 2 m³ lead cleaner conditioning tank.
• Five 16 m³ mechanically agitated, forced air conventional lead cleaner cells.
• Four 4.3 m³ mechanically agitated, forced air conventional lead recleaner cells.

Zinc flotation circuit.
The lead scavenger tailings and lead cleaner tailings will be pumped to the zinc flotation circuit. The zinc flotation circuit will consist of the zinc roughing and scavenger circuit followed by a zinc regrind of the rougher and scavenger concentrate and two stages of cleaning. The zinc concentrate will be pumped to the zinc concentrate thickener and the final tailings will be pumped to the tailings thickener.

Zinc rougher – scavenger circuit.
• Two 30 m³ zinc rougher mechanically agitated, forced air tank cells in series.
• Four 30 m³ zinc scavenger mechanically agitated, forced air tank cells in series.
• One 185 kW SMD zinc regrind mill.

Zinc cleaner and zinc recleaner circuit.
• One 5.0 m³ agitated zinc cleaner conditioner tank.
• Five 8 m³ mechanically agitated, forced air conventional zinc cleaner cells.
• Three 4.25 m³ mechanically agitated, forced air conventional zinc recleaner cells.

Concentrate production.
Final lead and zinc concentrates from the flotation recleaner circuits will be pumped to 10 m and 8 m diameter high rate concentrate thickeners respectively. No concentrate thickening test work has been completed and a conventional specific settling rate of 0.25 t/m2.h has been used to size the concentrate thickeners based on benchmark data from similar applications – a 1.25 safety factor on required area has been applied in addition given the fineness of the reground concentrates.

Flocculant will be added to increase the settling rate and underflow density to a target of 60% and 57% solids (w/w) for lead and zinc respectively. Thickened underflow will then be pumped to the associated lead and zinc agitated concentrate storage tanks by peristaltic pumps. To protect the lead and zinc filters, in-line strainers will be installed on the discharge lines from the thickener underflow pumps to remove trash. Overflow from lead and zinc concentrate thickeners will gravitate to the process water pond.

The filtration circuits for the lead and zinc concentrates will be similar. Each concentrate filtration circuit will consist of an agitated concentrate storage tank, filter feed pumps and plate and frame type pressure filters with ancillaries.

The lead concentrate storage tank will have a live volume of 200 m³, and the live volume of the zinc concentrate storage tank will be 100 m3.

Thickened lead and zinc concentrate slurries will be pumped from the respective concentrate storage tanks for dewatering in the pressure filters. The filtration area requirements are:

• 80 m² expandable to 100 m2 for the lead concentrate filter.
• 44 m² expandable to 60 m2 for the zinc concentrate filter.

The pressure filter will dewater the slurry to produce a filter cake with low residual moisture suitable for transport. Based on vendor advice on similar application, moisture levels in the range of 9% to 12% (w/w) can be expected for the concentrates. The filtrate will be pumped to its respective concentrate thickener to recover contained solids and water reclamation. The dewatered lead and zinc filter cakes will be discharged automatically from each filter press into the separate sections of the concentrate storage shed to minimize cross contamination.

Tailings thickening.
The final tailings, consisting the zinc scavenger and zinc cleaner tails will be transferred to a 16 m diameter high rate final tailings thickener. Flocculant will be added to increase the settling rate and produce an underflow density of 50 – 55 % solids (w/w). A flux rate of 1.0 t/m2.h has been used to size the tailings thickener – a 1.2 safety factor on required area has also been applied in the absence of specific test work. Underflow will then be pumped to tailings storage facility which will consist of the first of the smaller open pits used in the initial years of operation. Thickener overflow will gravitate to the process water dam.