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Ammonium thiosulfate leaching is a promising alternative to the conventional cyanide method for extracting gold from ores. However, strategies for recovering gold from the leachate are less commercially used due to its low affinity to gold. The present study investigated the recovery of gold from the leachate using iron oxides (hematite, Fe2O3 or magnetite, Fe3O4). Cementation experiments were conducted by mixing 0.15 g of aluminum powder as an electron donor and 0.15 g of an electron mediator (activated carbon, hematite, or magnetite) in 10 mL of ammonium thiosulfate leachate containing 100 mg/L gold ions and 10 mM cupric ions for 24 h at 25 °C. The results of the solution analysis showed that when activated carbon (AC) was used, the gold was recovered together with copper (recoveries were 99.99% for gold and copper). However, selective gold recovery was observed when iron oxides were used, where the gold and copper recoveries were 89.7% and 21% for hematite and 85.9% and 15.4% for magnetite, respectively. An electrochemical experiment was also conducted to determine the galvanic interaction between the electron donor and electron mediator in a conventional electrochemical setup (hematite/magnetite–Al as the working electrode, Pt as the counter electrode, Ag/AgCl as the reference electrode) in a gold–thiosulfate medium. Cyclic voltammetry showed a gold reduction “shoulder-like” peak at −1.0 V using hematite/Al and magnetite/Al electrodes. Chronoamperometry was conducted and operated at a constant voltage (−1.0 V) determined during cyclic voltammetry and further analyzed using SEM-EDX. The results of the SEM-EDX analysis for the cementation products and electrochemical experiments confirmed that the gold was selectively deposited on the iron oxide surface as an electron mediator.

The enhanced cementation technique by galvanic interaction of aluminum (Al; electron donor) and activated carbon (AC; electron mediator) to recover gold (Au) ions from the ammonium thiosulfate solution is a promising technique to eliminate the challenges of poor recovery in the system. This study presents the kinetics of Au ion cementation in an ammonium thiosulfate lixiviant as functions of initial Au concentration, size/amount of Al and AC, temperature, and shaking speed. The recovery results basically followed first order kinetics and showed that the cementation rate increased with a higher initial concentration of Au, smaller electron donor size, greater both electron donor and mediator quantity, decrease in temperature, and higher shaking speed in the system, while size of electron mediator did not significantly affect Au recovery.

Grinding conditions affect the flotation of copper sulphide minerals as changes in the properties of the grinding media and their interactions with the sulphide minerals, and between sulphide minerals themselves, affect the chemical environment in the flotation pulp. Galvanic interactions between steel grinding media and sulphide minerals, and between sulphide minerals, can lower the pulp potential, decrease the dissolved oxygen concentration in the mineral slurry, and lead to the dissolution of iron and copper from the media and the minerals. As a result, the formation of hydrophilic iron hydroxides and their adsorption on the copper sulphide minerals can be deleterious to copper flotation while pyrite (when present) can be activated to flotation by dissolved copper lowering the grade of the copper concentrate. Electrochemically less active grinding media (e.g., chrome alloy balls rather than mild steel media) can have beneficial effects on flotation performance due to the lower oxidation of the grinding media and consequently the lower production of oxidised iron species in the pulp. Copper activation of pyrite can be decreased by chemical additions to the pulp. In this paper, relevant experimental data published in the last 15 years are discussed.

The mineralogical characteristics of Copperbelt ores undergo continuous stratigraphic changes due to ongoing mining activities and weathering changes. These variabilities in mineralogy lead to drastically different behaviors in hydrometallurgical processing during the recovery of Cu and Co. Consequently, the technologies currently used in several hydrometallurgical industries have demonstrated their limitation in the effective processing of these ores. This paper focuses on the thermodynamic approach to assess the possibility of extracting Co and Cu from a mixture of ores characterized by Cu–Co oxide and Cu sulfide streams. To achieve the purpose of this investigation, the first phase consisted of thermodynamic predictions, while the second focused on experimental analysis. The thermodynamic feasibility of simultaneous Co and Cu extraction was assessed using potential–pH predominance and speciation diagrams of an actual leaching system. Thermodynamic calculations were used to construct Eh–pH and species distribution diagrams for the Co–Cu–Fe–SO 4 2− –H 2 O system under standard leaching conditions. Software tools such as Hydra-Medusa, HSC Chemistry, Phreeqc, and Geochemist’s Workbench were employed to calculate thermodynamic equilibria. The redox behavior of Fe species at pH 1.5 was also investigated to account for the chemistry and potential mechanisms of minerals during the leaching of the mixed system. Results revealed that experiments conducted under a free and average redox potential of 440 mV, pH of 1.5, solid percentage of 10%, stirring speed of 600 rpm, and temperature of 50 °C for 2 h led to a leaching efficiency of 86 and 70% Co and Cu, respectively. The dissolution of Co and Cu in the mixed system was promoted by the impact of chemical reactions involving the contribution of H + , HSO 4 − , and H 2 SO 4 , as well as redox reactions involving the Fe 2+ /Fe 3+ redox couple and the galvanic interactions between oxidized and sulfidic minerals. Graphical Abstract

Gold grains are observed in a variety of forms, such as coarse-liberated native gold grains, and ultra-fine grains associated with sulfide or non-sulfide mineral particles, in the form of solid solution in sulfide minerals, mainly pyrite. In the flotation of gold ores, bulk sulfide mineral flotation is generally applied to maximize gold recovery. This approach gives high gold recoveries, but it also causes the recovery of barren sulfide minerals (i.e., sulfide mineral particles with no gold content), which increases concentrate tonnage and transportation costs and reduces the grade sometimes to below the saleable limit (approx. 10 g/t Au). This study addresses the differences between gold-bearing and barren pyrite particles taken from various ore deposits and utilizes these differences for the selective flotation of gold-bearing pyrite. The laboratory scale flotation tests conducted on three pyrite samples having different cyanide soluble gold contents show that a selective separation between gold-bearing pyrite and barren pyrite particles could be achieved under specific flotation conditions. Gold recovery is correlated directly with the cyanide-soluble gold in the ore samples. Electrochemical experiments were conducted to elucidate the differences in surface properties of the two types of pyrite. The barren pyrite particles were more cathodic and prone to cathodic reduction of OH− and depressant ions on the surface, and they could be depressed effectively without significantly affecting the gold-bearing particles.

Beneficiation of sulphide ores by flotation is ascribed to the natural electrochemical activity associated with sulphide minerals. Flotation is an electrochemical process comprising many interdependent conditions that are difficult to decouple in terms of controlling flotation performance. The extent of electrochemical activity is mineral dependent and can be measured against a reference cell to differentiate between minerals. This difference in activity is known as the rest potential. The rest potential can be exploited to preferentially float one mineral over another as these properties result in different regions of flotation stability for different minerals. Bornite and chalcocite present an interesting study because when measured against the standard hydrogen electrode (SHE), there is a small difference in rest potential; bornite has a potential of 0.44 V and chalcocite a potential of 0.40 V. The key differentiating factor between the two minerals is the presence of iron in bornite (Cu5FeS4) and the lack thereof in chalcocite (Cu2S). This study considers bornite and chalcocite microflotation, adsorption studies and zeta potential measurements, and three key factors were explored: pH, galvanic interactions and collector adsorption. The overall objective of the study is to understand the response of bornite and chalcocite to changes in pulp chemistry.

Iron oxides (hematite, Fe2O3, and magnetite, Fe3O4), previously used as electron mediators in the galvanic system with zero-valent aluminum (ZVAl), have been shown to recover Au upon cementation in Au–Cu ammoniacal thiosulfate media selectively, and this warrants further investigation. This research is focused on investigating the role of the semiconductive properties of metal oxides by performing a cementation experiment by mixing 0.15 g of electron mediators (Fe3O4, Fe2O3, TiO2 (anatase and rutile)) and 0.15 g of zero-valent aluminum powder as an electron donor in various electrochemical experiments. The results revealed that upon the cementation experiment, synthetic Fe2O3 and Fe3O4 were consistently able to selectively recover Au at around 90% and Cu at around 20%. Compared to activated carbon (AC), TiO2, in anatase and rutile forms, obtained selective recovery of gold, but the recovery was utterly insignificant compared to that of iron oxides, obtaining an average of 93% Au and 63% Cu recovery. The electrochemical and surface analysis supports the results obtained upon the cementation process, where TiO2, upon cyclic voltammetry (CV), obtained two reduction peaks centered at −1.0 V and −0.5 V assigned to reducing Au and Cu ions, respectively. Furthermore, various electrochemical impedance spectroscopic analyses revealed that the flat band potential obtained in the Mott–Schottky plot is around −1.0 V and −0.2 V for iron oxides and titanium oxides, respectively, suggesting that the electrons travel from semiconductor interface to electrolyte interface, and electrons are accessible only to Au ions in the electrolyte interface (reduction band edge around −1.0 V). The determination of this selective cementation mechanism is one of a kind. It has been proposed that the semiconductive properties of Fe2O3, Fe3O4, and, by configuring their relative energy band diagram, the travel of electrons from the iron oxide–electrolyte interface facilitate the selective cementation towards Au(S2O3)23+ ions in gold–copper ammoniacal thiosulfate solutions.

In this paper, experimental data are compared for an aluminum/copper galvanic system with model predictions from a commercially available software package. The experimental data consisted of optical profilometry measurements of damage accumulation in a bimetallic Al-Cu strip as a function of immersion time in a chloride environment. Model predictions of corrosion rate were accomplished with software that used a solution of the Laplace equation to determine the potential distribution for a mesh of the system. Current density and, ultimately, corrosion rates were determined with this model by correlating the potential at a mesh point with a current density from the experimental potentiodynamic polarization curves for the materials. Although this is a proven method for systems where the anode material is under activation control, this methodology presents a problem for the case of pitting corrosion, such as Al and its alloys in chloride solution. In the case of pitting corrosion, the surface average current density from the polarization curve (i.e., the couple current density) is not equal to the current density at the base of the pit resulting in a model prediction that underestimates the corrosion rate. To overcome this shortcoming, an alternate method for determining the polarization curve for Al is offered from artificial pit electrodes (pencil electrodes) and model predictions using that method are presented.

The galvanic interaction between chalcopyrite and monoclinic pyrrhotite and its effect on flotation separation were studied using monomineral flotation tests, adsorption capacity tests, X-ray photoelectron spectroscopy (XPS) characterization, and scanning electron microscopy (SEM) test. These results showed that the interaction promoted the reduction of O2 on the cathodic chalcopyrite surface and accelerated the generation of Fe(OH)3, which was not conducive to collector adsorption; hence, the flotation recovery decreased by 10–16%. On the other hand, galvanic interaction accelerated the oxidation of S on the anodic monoclinic pyrrhotite surface to S0 or SO42− and produced a large amount of H+, thus preventing the formation of Fe(OH)3. Meanwhile, the Cu2+ eluted from chalcopyrite surface activated monoclinic pyrrhotite; hence, the flotation recovery increased by 3–10%. Galvanic interaction reduced the floatability difference between the two minerals, and the separation difficulty was significantly increased. Even with an increase in the amount of lime, the separation could not be improved.

The use of galvanic interactions between zero-valent aluminum (ZVAl) and activated carbon (AC) to recover gold (Au) ions is a promising technique to overcome the challenges due to the poor recovery in ammonium thiosulfate systems, but the applicability to practical Au ore processing remains elusive so far. The present study describes (1) the recovery of Au ions from low Au concentrations, which are typical concentrations used in Au ore processing; and (2) an investigation into the effects of various coexisting base metal ions that can be present in pregnant ore-leached solutions. The results showed that high Au recovery (i.e., over 85%) was obtained even at low Au concentrations under the following conditions: 1:1 of 0.15 g of ZVAl and AC with 10 mL of ammonium thiosulfate solution containing 5 mg/L of Au ions at 25 °C for 1 h in an anoxic atmosphere. Selected coexisting metal ions (i.e., copper, iron, cobalt, nickel, and zinc) were studied to establish their effects on Au recovery, and the results showed that the Au recovery was enhanced (about 90%) when copper ions coexist in the solution with minimal effects from other competing base metal ions.