Ferguson Lake Project — 2011 Technical Report

This report details the proposed hydrometallurgical processing route for the Ferguson Lake Project, based on testwork and engineering design for a chloride-based recovery system.

Report context

This technical report, dated November 30, 2011, documents the Ferguson Lake Project owned by Starfield Resources Inc. The report describes recovery methods developed through metallurgical testwork conducted since 2001 on mineralized material from the Main West Zone and West Extension. The proposed process was designed to minimize environmental effects, be highly energy efficient, and provide high base metal recoveries. Products include LME-grade nickel, copper, and cobalt metal, hematite, sulphuric acid, gangue material containing platinum and palladium, and power.

Processing route

Overview

The proposed process is a chloride-based hydrometallurgical system. Since 2001, the process has evolved but remains the same in the overall context of minimizing environmental impact and retaining the same product suite. The process design assumes a mining rate of 6,000 tpd of ore, from which the concentrator will produce approximately 3,300 tpd of bulk sulphide concentrate.

Mine site concentrator

Run of mine ore will be crushed and ground to a relatively coarse grind (P90 of 100 mesh, or 150 μm) at the mine site. Ground ore will be subjected to flotation to recover essentially all sulphide minerals and precious metals into a bulk concentrate, while rejecting the bulk of the gangue minerals. The concentrator will produce tailings requiring disposal at the project site, with a portion disposed of as underground backfill. Mill tailings will also be stored in a tailings management facility at the mine site.

Concentrate transport

The bulk sulphide concentrate will be slurried in water and transported by pipeline to the process site. The process design assumes a 285 km pipeline will deliver concentrate to Arviat, where it will be filtered ahead of the process plant.

Primary leach

Slurried concentrate is transported from the mine site by pipeline into the filtration plant. The filter cake is slurried and transported into the primary leach using ferrous chloride solution recycled from the primary leach thickener. Leaching takes place in four cascading reactors with residence time of about 30 minutes per reactor. Primary leaching is carried out with hydrochloric acid at approximately 115°C, producing a ferrous chloride solution, hydrogen sulphide off gas, and a residue containing the base and precious metals. Pyrrhotite reacts with HCl to form FeCl₂ and H₂S. Base metals and precious metals do not react appreciably with HCl under reducing conditions.

Hydrogen sulphide treatment

The H₂S off-gas from the primary leach will be cooled to condense out most of the water and any residual HCl, prior to burning in air to generate high-strength sulphur dioxide, which will be sent to an acid plant to produce 93% sulphuric acid. The H₂S from the primary leach, and sulphur from secondary leach, are burnt to SO₂ using air. A waste heat boiler will cool the gases and generate 80 Bar steam, which will be expanded to 28 Bar in a turbine to recover power for generating electricity. Approximately 1.1 Mt of acid will be produced each year.

Oxidation stage

The solution from primary leach is concentrated and heated in an evaporation stage. FeCl₂ in the liquid from evaporation is oxidized to FeCl₃ and hematite using oxygen in three cascading reactors at 120°C. One third of the iron precipitates as hematite. Each reactor has a residence time of about 37 minutes. A second evaporation stage further concentrates and heats the slurry.

Hydrolysis stage

The temperature of the slurry is raised to 180°C by condensing 28 Bar steam. Hematite is precipitated and concentrated HCl vaporizes in three reactors. A high temperature filter separates the hematite from the liquid. The filter cake is re-pulped with raffinate from cobalt solvent extraction, filtered, and the hematite is stored prior to shipment.

Secondary leach

Primary leach underflow is treated with oxygen in a low pressure autoclave to oxidize ferrous to ferric, dissolve the base metals, and precipitate most of the soluble iron as goethite. A residence time of five hours in the autoclave is assumed. Slurry leaving the secondary leach is thickened and filtered on a belt filter. The filter cake is repulped with recycled concentrated HCl to dissolve the goethite. Elemental sulphur is recovered by flotation from the filter cake. The leach residue contains the PGEs and will be stored until a treatment process is available.

Solvent extraction and metal recovery

The solution from the secondary leach is purified by solvent extraction of residual ferric chloride, then by ion exchange to remove zinc chloride. The purified solution is subjected to further solvent extraction in stages to separate individual base metals.

Copper and cobalt recovery occurs through anionic solvent extraction. A second stage of solvent extraction separates copper from cobalt. Each metal is extracted into a sulphate solution using solvent extraction and electrowon from the sulphate medium to generate LME grade metal.

Nickel is recovered from the solution leaving manganese dioxide electrowinning by a nickel specific cationic organic solvent. The loaded organic phase is stripped with sulphuric acid in spent electrolyte from the subsequent nickel electrowinning step.

Manganese is removed as manganese dioxide from the raffinate after the anionic solvent extraction of cobalt and copper using electrolysis. The manganese dioxide will be discarded and the resulting solution neutralized with lime.

Electrowinning

Copper electrowinning uses 37 cells, each containing 85 stainless steel cathodes of 1 m by 1 m, at a current density of 300 amps/m². Nickel electrowinning uses 42 cells, each containing 50 stainless steel cathodes of 1 m by 1 m, with cloth diaphragms on anodes. Cobalt electrowinning uses eight cells, each containing three stainless steel cathodes of 1 m by 1 m, with cloth diaphragms on anodes.

Acid regeneration and energy recovery

The high-pressure steam generated in the acid plant will be expanded through turbines to generate power. Lower pressure steam will be used in the evaporation, oxidation, and hydrolysis stages. The HCl regeneration section will consume approximately 260,000 tpa of acid, leaving some 900,000 tpa of excess acid for shipment and sale. Approximately 500,000 tonnes of acid will be stored during the non-shipping season in a battery of mild steel tanks.

Key reported parameters

Technical qualifications

The report includes specific limitations and qualifications that should be noted. Capital and operating cost estimates have an accuracy of ±40%, corresponding to a scoping level study. Additional test work should be carried out to confirm operating parameters. The equipment to be used in the test plant should be made of material pre-selected for use in the main plant. Results from tests should be used in continuing engineering studies and to solve materials of construction concerns. Materials of construction for primary leach equipment need more evaluation. Care will be needed in designing the primary leach stage because H₂S is very toxic and mixtures of H₂S and air are explosive between 4% and 46% H₂S. Materials of construction for reactors, heat exchangers and hot filter in the hydrolysis stage need to be further investigated. The leach residue containing PGEs will be stored until a treatment process is available.

Source: Starfield Resources Inc. – Ferguson Lake Project, Technical Report NI 43-101 – November 30, 2011, Sections 17.