The Thackaringa mineralisation comprises stratabound units of moderate to steeply dipping, pyritic quartz-albite gneiss that form three deposits referred to herein as Pyrite Hill, Big Hill and Railway. Pyrite Hill is geographically separate from the other deposits. Conversely, Big Hill and Railway are considered to reflect the same mineralised body, separated by a zone of low grade mineralisation and minor structural dislocation.

Controls on mineralisation are considered to include:

- Primary foliation of the host lithology as a fluid flow pathway and depositional site for the cobaltiferous pyrite; and
- Bedding parallel shear zones, generally occurring along the quartz-albite gneiss contact, responsible for evident fold thickening.

Pyrite Hill
The Pyrite Hill deposit extends over 1.2 km along strike, approximately 300 m down dip and varies in thickness from approximately 10 to 100 m. Mineralisation is hosted by quartz-albite gneiss with both the hanging wall and footwall comprised of quartz-albite-biotite gneiss with lesser quartz-albite gneiss and amphibolite sills.

The northern-western extent of the deposit is generally undeformed and dips at approximately 50° to the northeast. In the central part of the deposit, rapid thickening of mineralisation, resultant of near isoclinal folding, occurs in correlation with a general change in strike to the south and coincident steepening of dip to approximately 60° to the east.

The Mineral Resource estimate extends from the base of partial oxidation (approximately 20–25 m below surface) to 35mRL (approximately 270 m below surface).

Big Hill
The Big Hill deposit has an overall strike length of 1.2 km and comprises two mineralised zones separated by a late stage dextral fault with approximately 150 m of apparent displacement. The southwestern zone occurs over 800 m of strike varying in thickness from 30–100 m due to steep isoclinal folding. The northern-eastern zone is a relatively linear, steeply dipping zone extending for some 400 m with an average thickness of 35–40 m.

The base of partial oxidation occurs approximately 10–25 m below surface with narrow zones of deeper, structurally controlled oxidation evident at the southern extent of the deposit. The Mineral Resource estimate extends from the base of partial oxidation to 150 mRL (approximately 150 m below surface).

Railway
The Railway deposit is considered a north-eastern extension of the Big Hill deposit with continuous mineralisation observed over some 2.5 km. The southern extent of the deposit is generally linear with an average thickness of 30 m increasing to approximately 60 m in correlation with evidently upright isoclinal folding. The central part of the deposit is characterised by extensive ductile deformation and complex folding resulting in a rapid thickening of mineralisation up to 300 m. At the northern-eastern extent, the mineralisation is increasingly discontinuous, comprising a series or narrow lenses within a weakening low grade mineralised envelope.

The base of partial oxidation generally occurs approximately 15–20 m below surface. The Mineral Resource estimate extends from the base of partial oxidation to 50 mRL (approximately 230 m below surface) with a section between 6540950mN and 6451400mN at 0 mRL (approximately 300 m below surface).

The PFS considers a multi-open pit mining scenario that will extract ore using conventional drill and blast, load and haul and dump processes. The operation is planned to use excavator and rigid body trucks along with a fleet of auxiliary equipment. This proposed mining method is considered appropriate for the deposit style.

Approximately 5.25Mt of ore will be hauled annually to a stockpile area (ROM) proximal to the processing plant located centrally to the pits and waste material hauled to the waste emplacements located in close proximity of each pit. During periods where the quantity of ore mined exceeds the quantity processed, additional temporary long- term stockpile areas may be utilised.

As the project consists of a simple bulk massive style deposit with no internal waste, a mining recovery of 95% and mining dilution of 5% has been assumed.

The key target metal in the ore is cobalt, and this is present within a pyrite mineral lattice as a solid-solution substitution for iron atoms.

The two major components of the ore are pyrite and albite, along with minor amounts of quartz, mica, silicate. The ore is differentiated from waste rock on an economic basis of cobalt content ~ 500 ppm (mining ore reserve cut-off). On account of the substitution of cobalt into pyrite, and the lack of any other cobalt-bearing minerals, the ore grades can be broadly indicated by monitoring the pyrite content, i.e. higher pyrite content in the rocks indicate higher cobalt grades.

Using the processing options studied in the Scoping Study (2017) and the Pre-Feasibility Study (2017–2018), Cobalt Blue has selected a preferred flowsheet for development of the Thackaringa cobalt project. The flowsheet is focused on concentrating the pyrite from the ore, and then processing the pyrite to recover cobalt. After a review of the current and forecast market for cobalt over the next fifteen years, the final form of cobalt selected for production was cobalt sulphate heptahydrate crystals. These salts are used in the production of lithium ion batteries.

CONCENTRATION OF PYRITE FROM ORE
The mined ore is crushed to p80 ~ 800–900 um (p100 1.2mm), and passed over gravity spirals to produce a pyrite concentrate. The gravity tails are screened and the fines fraction (<125 um) is sent to a scavenger flotation circuit to recover any sulphides. The use of gravity spirals, takes advantage of the coarse pyrite grains (p80 200-800 um), and limits costs associated with crushing and milling the ore, as would be the case for a typical flotation circuit requiring feed at p80 100–200 um.

In the PFS testwork program, 820 kg of ore at 600 ppm cobalt was trialled using a full-sized gravity spiral and a 14 L flotation cell. The recovery of cobalt to concentrate was 92%, at a grade of 3326 ppm. The ore was tested on a continuous pilot basis.

THERMAL DECOMPOSITION (PYROLYSIS) OF PYRITE CONCENTRATE
The pyrite mineral is thermally decomposed into pyrrhotite and elemental sulphur by heating to 650–700°C. A nitrogen atmosphere is used to prevent any oxidation. The off-gas is collected, and cooled to recover the sulphur. In the PFS testwork program, 100 kg of concentrate grading 3326 ppm cobalt was processed in a custom built rotary furnace. Variations in operating conditions were tested, with the best results showing that >95% of the pyrite could be converted into pyrrhotite along with the simultaneous recovery of 40% of the head sulphur. The calcine was then passed through a magnetic separator to prepare a magnetic fraction containing pyrrhotite for leaching, and a non-magnetic fraction containing unreacted pyrite for recycle to the concentrator circuit.

LEACHING AND PRODUCTION OF MIXED HYDROXIDE PRECIPITATE
The artificial pyrrhotite is leached in a low- temperature (130°C) and pressure (10–15 bar) autoclave. The resulting leach residue is screened, and the coarse fraction is sent for sulphur recovery by distillation or remelting. The fines fraction is discarded as tails from the process plant. The resulting leach solutions are treated to remove iron, copper and zinc before precipitating the cobalt as a mixed hydroxide (along with nickel and manganese).

In the PFS testwork program, ~ 30 kg of calcine product from the furnace was leached in batches of 250g to 1kg. Variations in the operating conditions were tested, with the best results showing that 97-98% of the cobalt could be leached consistently from the pyrolysis calcine.

REFINING OF THE MIXED HYDROXIDE PRECIPITATE TO PRODUCE COBALT SULPHATE CRYSTALS
In the PFS testwork program, variations on the ion-exchange and solvent extraction circuits were tested. The best conditions resulted in the production of cobalt sulphate heptahydrate grading ~20.5% with total impurities at ~800 ppm copper and 800 ppm manganese. Further optimisation of the parameters for the ion-exchange circuits, is expected to reduce the copper and manganese content reporting to the cobalt sulphate in future testwork.

An engineering design study was completed using the metallurgical testwork results as the design criteria basis. The throughput rate was fixed to 5.25 Mtpa of ore.