Thiosulfate leaching is a promising alternative to cyanidation, and the main hindrances for its wide commercial application are the high thiosulfate consumption and the difficult recovery of dissolved gold. In this review, the four solutions to reduce the consumption of thiosulfate, including the control of reaction conditions, the use of additives, the generation of thiosulfate in situ, and the replacement of traditional cupric-ammonia catalysis, are introduced and evaluated after the presentation of background knowledge about thiosulfate consumption. The replacement of cupric-ammonia catalysis with other metals, such as nickel- and cobalt-based catalysts, is proposed. The reason is that it not only reduces thiosulfate consumption observably via decreasing the redox potential of leach solution significantly but also is beneficial to gold recovery mainly owing to eliminating the interference of cuprous thiosulfate [Cu(S₂O₃)₃]⁵⁻. Based on the comparative analysis for five common recovery techniques of rare-noble metals from pregnant leach solution, ion-exchange resin adsorption is considered to be the most appropriate to recover aurothiosulfate [Au(S₂O₃)₂]³⁻ because the resin can be employed in the form of resin-in-leach/pulp and, furthermore, is able to be eluted and regenerated simultaneously at ambient temperature. At last, how to reduce the process cost of the resin adsorption technique is discussed. In order to simplify the complex two-stage elution process for loaded resins, the traditional catalysis is suggested to be replaced.

This review is devoted to an integrated evaluation of the current use and future development of the resin adsorption technique in gold recovery from pregnant thiosulfate solutions. Comparisons are firstly made with other recovery techniques, including precipitation, activated carbon adsorption, solvent extraction, electrowinning and mesoporous silica adsorption. A detailed discussion about the recent advances of the technique in gold recovery from pregnant thiosulfate solutions is then presented from the aspects of gold adsorption on the resins and gold-loaded resin elution, respectively. On the basis of summarizing the present research, the major limitations of the resin adsorption technique are eventually pointed out and future development will also be prospected.

The effect of pyrite and the role of ammonium alcohol polyvinyl phosphate (AAPP) during gold leaching in ammoniacal thiosulfate solutions were investigated using pure gold foils. The results showed that pyrite catalyzed the decomposition and also significantly increased the consumption of thiosulfate. This detrimental effect became more severe with increasing pyrite content. Further, the presence of pyrite also substantially slowed the gold leaching kinetics and reduced the overall gold dissolution. The reduction in gold dissolution was found to be caused primarily by the surface passivation of the gold. The negative effects of pyrite, however, can be alleviated by the addition of AAPP. Comparison of zeta potentials of pyrite with and without AAPP suggests that AAPP had adsorbed on the surface of the pyrite and weakened the catalytic effect of pyrite on the thiosulfate decomposition by blocking the contact between the pyrite and thiosulfate anions. AAPP also competed with thiosulfate anions to complex with the cupric ion at the axial coordinate sites, and thus abated the oxidation of thiosulfate by cupric ions. Moreover, the indiscriminate adsorption of AAPP on the surfaces of gold and passivation species prevented the passivation of the gold surface by surface charge and electrostatic repulsion. Therefore, AAPP effectively stabilized the thiosulfate in the solution and facilitated the gold leaching in the presence of pyrite.

High-aluminium-content iron ore is one of typical intractable iron ores, and magnetic separation and floatation processes are found impracticable to remove alumina from the ore effectively. In this article, a new process, roasting with addition of soda followed by leaching, is developed to remove aluminium from the ore. Results show that Al₂O₃ content decreases from 8.16% in raw ore to 2.13% in iron concentrate, and total iron grade increases from 48.92 to 63.21% when the ore is roasted at 1000°C for 15 min with the addition of 14.0% (wt.) sodium carbonate. Mechanisms of aluminium–iron separation were studied by using XRD, SEM, and thermodynamic methods, and it is shown that aluminium is transformed into sodium aluminosilicate, sodium aluminate, and corundum during roasting; sodium aluminate is able to be leached by water, so is sodium aluminosilicate by dilute acid solution, while corundum remains in the iron concentrate.