The role of amino acids in the thiosulphate leaching of gold

Amino acids L-valine, glycine, DL-α-alanine and L-histidine were applied in the ammoniacal thiosulphate leaching of a pyrite concentrate. Amino acids formed more stable complexes with cupric ions than does ammonia, resulting in lower thiosulphate consumption due to reduced interaction between thiosulphate and the copper complexes. Overall gold extraction was largely improved by the addition of amino acids, despite lower initial kinetics. Thiosulphate consumption decreased with an increase in the amino acid concentration. Among the amino acids, L-histidine improved overall gold extraction and reduced the thiosulphate consumption to the greatest extent.     

Factors influencing the rate of gold cyanide leaching and adsorption on activated carbon, and their impact on the design of CIL and CIP circuits

The carbon in pulp (CIP) and carbon in leach (CIL) processes became firmly established in the gold mining industry in the 1980s, initially in South Africa and Australia, from where they spread rapidly to all the gold producing regions of the world. The percentage of annual global gold production by activated carbon-based processes grew from zero in the 1970s to almost 70% by the turn of the century, which represented a phenomenal rate of acceptance of a new technology by a traditionally conservative industry. The main reason for this rapid acceptance of the new technology was the fact that the first few large industrial plants in South Africa convincingly demonstrated better gold recoveries than the traditional filtration/Merrill Crowe process, with lower capital and operating costs. And as the plants developed an operating track record over their first few years of life, they proved to be remarkably robust mechanically, and highly tolerant of plant upsets, changes in feed composition and solution phase contaminants that had caused great problems in Merrill Crowe plants.These stellar attributes of the carbon-based gold plants have led to complacency and laziness in the industry, both at the new plant design stage, and with on-going optimization of existing plants. In many cases, basic “rules of thumb” that were developed as design criteria for the early CIP plants are still used today, with no appreciation of the factors that may cause one plant to perform quite differently from another. There seems to be little incentive to improve performance when it is well known that most CIP and CIL plants operate quite well with minimal optimization and, in many cases, minimal understanding of the factors that influence performance. Consequently, almost all CIP and CIL plants are overdesigned at the construction stage and are then operated sub-optimally. This can lead to higher gold losses and/or higher capital and operating costs than necessary.This paper examines the factors that influence CIP and CIL plant design and performance, and demonstrates a very simple methodology that can be used to arrive at something close to an optimum plant design. It can also be used as an on-going tool by plant metallurgists to transform a fairly well run plant into an exceptionally well run plant.     

Enhanced leachability of gold and silver in cyanide media: Effect of alkaline pre-treatment of jarosite minerals

Composite samples of tailings containing gold (1.35 g/t) and significant amounts of silver (155 g/t) were subjected to batchwise cyanide leaching to assess the feasibility of extracting gold and silver. The tailings are waste solids arising from flotation and leaching operations whereby the flotation product (sphalerite concentrate) is calcined and then solubilised into dilute sulphuric acid solution and eventually sequestered from the electrolyte by electrowinning. Silver and gold are part of the zinc refinery residue, flotation tailings and to a limited extent the calcine leach tailings. Mineralogical results showed that composite tailings are refractory in nature (44% quartz, 17% silico aluminates and 12% jarosites).The concept of enhancing gold and silver recovery from the tailings focused on firstly decomposing the jarosite minerals by alkaline pre-treatment and then secondly leaching with cyanide solution. These two steps ensured that free gold and silver found in the zinc refinery residue and in the jarosite minerals could be leached simultaneously. The composite tailings were treated with Ca(OH)₂ solutions and then heated to 90 °C for 2 h to decompose the silver-bearing mineral (Ag,PbFe₃(SO₄)₂(OH)₆). The alkaline pre-treated tailings were then subjected to cyanide leach tests at different NaCN dosages (2.5–10 kg/t) and particle size (96–200 μm). Without an alkaline pre-treatment stage, leach efficiencies achieved were 41% and 25% for gold and silver, respectively at 40 °C and 8 h mixing time. But, better leach efficiencies (55% for Au, 81% for Ag) were achieved after the feed was pre-treated with Ca(OH)₂. The leaching mechanism of gold was explained by the shrinking sphere model denoted by surface chemical reaction.     

Review of rate constants for thiosulphate leaching of gold from ores,concentrates and flat surfaces: Effect of host minerals and pH

A review of literature data for different types of sulphide concentrates and gold ores has been carried out to examine the impact of host minerals and pH upon gold leaching. Analysis of initial rate data over the first 30–60 min of gold leaching from sulphide concentrates or silicate ores over a range of ammonia, thiosulphate, and copper(II) concentrations, pH (9–10.5) and temperatures up to 70 °C shows the applicability of a shrinking sphere kinetic model with an apparent rate constant of the order k_ss = 10⁻⁶–10⁻³ s⁻¹. The dependence of apparent rate constant on pH and initial concentrations of copper(II) and thiosulphate is used to determine a rate constant k_Au(ρr)⁻¹ of the order 1.0 × 10⁻⁴–7.4 × 10⁻⁴ s⁻¹ for the leaching of gold over the temperature range 25–50 °C (ρ = molar density of gold, r = particle radius). These values are in reasonable agreement with rate constants based on electrochemical and chemical dissolution of flat gold surfaces: k_Au = 1.7 × 10⁻⁴–4.2 × 10⁻⁴ mol m⁻² s⁻¹ over the temperature range 25–30 °C. The discrepancies reflect differences in surface roughness, particle size and the effect of host minerals.     

Paste backfill of high-sulphide mill tailings using alkali-activated blast furnace slag: Effect of activator nature,concentration and slag properties

The effect of activator type, concentration and slag composition on the strength and stability properties of paste backfill (CPB) of high-sulphide tailings using alkali-activated slag (AAS) as binder (7 wt.%) were investigated in this study. Acidic and neutral (AS–NS) slags were activated with liquid sodium silicate (LSS) and sodium hydroxide (SH) at 6–10 wt.% concentrations. Ordinary Portland cement (OPC) results were used for comparison. The strength development was found to remarkably improve with increasing the concentration from 6 to 8 wt.%. Further increase in concentration did not enhance the strength. SH was determined to produce higher early-age strength whilst LSS produced higher long-term strengths as an indication of slag selectivity for activators. More extensive gypsum formation was observed at lower concentrations in SEM/EDS studies. An increase in Na₂O concentration raised the activator consumption. High concentrations also led to poorly crystallized C–S–H gel, loose structure and drying shrinkage cracks especially in NS–SH samples. A reduction in total porosity up to 20% was obtained in AAS samples compared to OPC. Amorphous structure, chemical modulus ratio and/or basicity index (BI) values were seen to control the pozzolanic reactivity, and therefore, the alkali-activation and hardening process.