Advances in Mineral Processing and Hydrometallurgy: 2nd Edition

1. Introduction and Scope

After a very successful initial debut, this second edition of a Special Issue of Metals has been commissioned that is specifically devoted to aspects of mineral processing and hydrometallurgy. Again, the editors are Prof. Dr. Corby Anderson and Dr. Hao Cui. This is a unique, focused Special Issue, as the specific combination of hydrometallurgy and mineral processing is not common. Both technical regimes will be relied upon to produce the supply of critical minerals and metals needed for renewable energy.

As Free [1] states of water make aqueous processing of metals possible, water is an unusual substance from a chemical perspective. Most substances with similar molecular weights, such as methane and ammonia, are gases at room temperature. However, water is a liquid at room temperature. Water’s unusual properties are due primarily to hydrogen bonding effects. These are related to the tendency of hydrogen to donate electrons and the tendency of oxygen to accept electrons. As such, hydrometallurgy offer superior separations, as well as the capability to treat low-grade materials to produce fungible compounds and metals.

Wills and Napier Munn [2] recently published a seminal book on mineral processing. They stated “As-mined” or “run-of-mine” ore consists of valuable minerals and gangue. Mineral processing, sometimes called ore dressing, mineral dressing, or milling, follows mining and prepares the ore for extraction of valuable metals in the case of metallic ores, and produces a commercial end product of products such as iron ore and coal. Apart from regulating the size of the ore, it is a process of physically separating the grains of valuable minerals from the gangue minerals to produce an enriched portion, or concentrate, containing most of the valuable minerals and a discard, or tailing, containing predominantly the gangue minerals. The importance of mineral processing is taken for granted today, but it is interesting to reflect that less than a century ago, ore concentration was often a crude operation involving simple gravity and hand-sorting techniques performed by the mining engineers. The twentieth century saw the development of mineral processing as a serious and important professional discipline in its own right, and without physical separation, the concentration of many ores, and particularly the metalliferous ores, would be hopelessly uneconomic.

Uniquely, Fuerstenau and Han [3] published a book devoted exclusively to the combination of mineral processing and hydrometallurgy. In this unique publication, they stated that to advance the technology in the production of material resources, nations look to current practicing and future practicing engineers. Current and future mineral processing engineers must obtain sound and rigorous training in the sciences and technologies that are essential for effective resource development. Many industrial and academic leaders have recognized the need for more textbooks and references in this critical area. This was the driving force for writing a comprehensive reference book that covers mineral processing and hydrometallurgical extraction.

Hence, this Special Issue, the second edition, was designed to include submissions for characterization, along with recycling and waste minimization, mineralogy, geometallurgy, thermodynamics, kinetics, comminution, classification, physical separations, liquid–solid separations, leaching, solvent extraction, ion exchange, activated carbon, precipitation, reduction, process economics, and process control. The suggested application areas were gold, silver, PGMs, aluminum, copper, zinc, lead, nickel, and titanium. Critical metal articles on topics such as lithium, antimony, tellurium, gallium, germanium, cobalt, graphite, indium, and the rare earth metals were welcomed. Ten high-quality peer-reviewed articles from around the globe were selected, peer reviewed, and accepted for inclusion.

2. Overview of the Published Articles

Taboada et al. [Contribution 1] provided an article entitled ‘Obtention of Suitable Pregnant Leach Solution (PLS) for Copper Solvent Extraction Plants from Copper Concentrate Using Hydrogen Peroxide and Iodine in a Sulfuric Acid–Chloride Medium’. Copper leaching presents an environmentally friendly alternative to traditional sulfide ore-processing methods. This study investigates an efficient leaching process for copper concentrate, utilizing a solution of sulfuric acid (H₂SO₄) and potassium iodide (KI) in a chloride medium (NaCl), enhanced by hydrogen peroxide (H₂O₂) at room temperature. A significant aspect of this research was optimizing the KI concentration to minimize iodide sublimation into iodine gas (I2). Through the experimental design, the optimal dosages of reagents were determined, leading to a maximized copper extraction of approximately 27% in 45 min of testing at room temperature. The results show that it is possible to obtain a suitable pregnant leach solution (PLS) (i.e., in a range from 3 to 8 g/L of Cu) for treatment in available copper solvent extraction (SX) plants with a cost of less than 4.5 USD/t Cu, according to the economic analysis carried out. The results of this study determine the most effective operational conditions for leaching and ensure a suitable PLS for SX plants in a cost-effective and environmentally friendly manner. This approach could significantly contribute to more sustainable practices in the mining and processing of copper ores.

Bruce et al. [Contribution 2] researched and published the article, ‘Effect of Incorporation of Sulfation in Columnar Modeling of Oxidized Copper Minerals on Predictions of Leaching Kinetics’. Mathematical modeling of columnar leaching is a useful tool for predicting and evaluating the kinetics of copper extraction. One commonly used model for this process is the shrinking core model (SCM). In this study, the aim was to develop a model for column leaching of oxidized copper ore based on the SCM, which incorporates the ore sulfation stage before leaching. In sulfation and leaching laboratory-scale tests, we studied the effect of acid dosage (at 22.8, 34.2, and 45.6 kg/t), humidity (at 90%, 100%, and 110% of the saturation humidity of the mineral), ore granulometry (between −3/4″ and −3/8″), and rest time (at 24, 48, 72, and 96 h) on sulfation. We found that the highest sulfation reached 49.7% for both granulometries in the studies. In the column tests, the effects of acid dosage (at 34.2 and 45.6 kg/t), ore granulometry (−3/4″, −3/8″), and rest time (at 24 and 48 h) were studied. When the SCM was applied to these tests, we obtained fit qualities between 63.4% and 74.9%. By incorporating the sulfation factor into the SCM predictions, we observed an average increase in adjustment between 24% and 28%. This method is effective for minerals and operating conditions different from the ones studied.

Karimov et al. [Contribution 3] provided an article entitled ‘Purification of Copper Concentrate from Arsenic under Autoclave Conditions’. This study presents the results of a two-stage autoclave processing of a copper–arsenic concentrate. Copper concentrate is an important raw material to produce copper and other metals. However, in some cases, the concentrate may contain increased amounts of arsenic, which makes further processing difficult. Therefore, the development of modern hydrometallurgical methods for processing copper concentrate with a high arsenic content is an urgent task which could lead to the optimization of the raw material processing process and an improvement in the quality of the concentrate. It has been established that the optimal conditions for the sequential two-stage autoclave processing of the copper–arsenic concentrate are as follows: t = 220–225 °C, τ oxidation = 20 min, τ tot = 90 min, Po₂ = 0.4 MPa, and L:S = 10:1, [H₂SO₄] initial = 40 g/dm³. In this case, 85% of zinc, 44% of iron, and 78% of arsenic, respectively, are extracted into the solution during both stages, and the loss of copper was about 0.01%. This is explained by the fact that, at the first stage (oxidation) of the autoclave processing of the copper–arsenic concentrate, copper, together with iron, leaches into the solution, and at the second stage (reduction), copper precipitates out of the solution in the form of chalcocite. Copper in the residue after autoclave leaching takes the form of Cu2S, iron takes the form of pyrite (FeS2), and lead takes the form of anglesite (PbSO4), respectively. The obtained micrographs and EDX mappings clearly show no iron arsenates. This confirms that, at the oxidative stage of the developed process, arsenic, removed by 78%, remains in the solution. The remaining arsenic is associated with tennantite, indicating the effectiveness of the treatment process in removing arsenic from the copper–arsenic concentrate. A second important observation is the presence of pronounced areas of copper sulfides in the micro photos without iron and arsenic impurities. This confirms that copper is deposited as chalcocite during the reduction phase of the process, which is the desired result.

Askarian et al. [Contribution 4] contributed a paper entitled ‘Towards Using MMO Anodes in Zinc Electrorefining: Mn Removal by Simulated Plant Off-Gas’. Implementing mixed-metal oxide (MMO) anodes in zinc electrowinning is highly desired due to the considerable reduction in electrical energy consumption. However, the presence of manganese in the electrolyte is a major obstacle for implementing MMO anodes in the zinc cell houses. In this work, we explore the possibility of using plant off-gas containing SO₂ to remove manganese. An SO₂/air–gas mixture with different SO₂ and O₂ concentrations was therefore used for the oxidative precipitation of manganese. It was shown that the manganese oxidation reaction is highly pH dependent. Calcium hydroxide was used to control the pH during the process. Different operating parameters, i.e., pH, SO₂/air ratio, reaction time, and effect of cobalt as a reaction catalyst, were investigated. Optimal conditions for manganese removal were reported. Under the optimal conditions, the manganese concentration decreased from 1 g L⁻¹ to less than 1 mg L⁻¹ within 30 min. Precipitates were characterized using EDS, XRF, and XPS techniques and showed coprecipitation of manganese, zinc, gypsum, and cobalt.

Fu et al. [Contribution 5] published the article, ‘Adsorption of Gold from Copper–Tartrate–Thiosulfate Solutions with Ion-Exchange Resins’. The adsorption behavior of gold from copper–tartrate–thiosulfate solutions with ion-exchange resins was studied in this paper. Experimental parameters include resin dosage, pH, temperature, copper, tartrate, and thiosulfate concentration. A moderate increase in resin dosage, pH, temperature, and tartrate concentration is beneficial for gold adsorption, but an excessive tartrate concentration or higher temperature depresses the adsorption process. Increasing copper and thiosulfate concentrations may competitively occupy the active sites on the resin surface, leading to a reduction in the gold adsorption capacity. The XPS and FT-IR analyses indicate that copper and gold on the resin after adsorption exist in the form of Cu+ and Au+, and sulfur exists in the form of SO₄²⁻ and S₂O₃^2−..This implies that the use of resin for gold recovery from thiosulfate leachate may face critical challenges because there is inevitably a higher content of copper and thiosulfate.

Zhang et al. [Contribution 6] authored a contribution entitled ‘Effects of Ammonium Salts on Rare Earth Leaching Process of Weathered Crust Elution-Deposited Rare Earth Ores’. To reveal the influence of ammonium salts on the rare earth leaching process of weathered crust elution-deposited rare earth ores, ammonium acetate, ammonium chloride, and ammonium sulfate were used as leaching agents. The effects of the leaching agent on the rare earth leaching efficiency and the expansion, dissolution, and transformation behavior of clay minerals in the rare earth leaching process were studied. The results show that rare earth leaching efficiency followed the order of ammonium acetate > ammonium chloride > ammonium sulfate, with values of 90.60%, 85.96%, and 84.12%, respectively. The swelling ratio of clay mineral followed the order of ammonium acetate < ammonium chloride < ammonium sulfate; the clay mineral swelling ratio was 2.09% when ammonium acetate was the leaching agent. Thermogravimetric analysis showed that the interlayer water content was the lowest when ammonium acetate was used as the leaching agent. Under the conditions of different leaching agents, the clay mineral contents changed from illite and halloysite to smectite and kaolinite. When ammonium acetate was used as the leaching agent, the relative conversion of illite was 1.49%, and that of smectite was only 0.17%. SEM analysis showed that the clay minerals expanded and dissolved obviously when ammonium chloride and ammonium sulfate were used as leaching agents. Meanwhile, the clay mineral layered structure was complete when ammonium acetate was used as the leaching agent. Therefore, when ammonium acetate was used as the leaching agent, it had the least effect on the swelling, dissolution, and transformation of clay minerals. This can provide a theoretical basis for the safe production of weathered crust elution-deposited rare earth ore and for the screening of green and efficient leaching agents.

Stallmeister and Friedrich [Contribution 7] researched and submitted an article on the ‘Influence of Flow-Gas Composition on Reaction Products of Thermally Treated NMC Battery Black Mass’. The recycling of lithium-ion batteries (LIBs) is becoming increasingly important regarding the expansion of electromobility and aspects of raw material supply. Pre-treatment and liberation are crucial for a sufficient recovery of all relevant materials from LIBs. Organic removal and phase transformations by thermal pre-treatment are beneficial in many respects. This study deals with the influence of flow-gas composition on reaction products and water-based lithium recovery after thermal treatment. Therefore, a spent NMC black mass was thermally treated at 610 °C in a moved bed batch reactor under an N₂ atmosphere and mixtures of N₂ with 2.5% and 5% O₂. Since the phase transformation of the lithium content to Li₂CO₃ is targeted for water leaching, treatment under a CO₂ atmosphere was studied as well. The resulting off-gas was analyzed using FTIR, and the black mass was characterized using XRD. Afterward, water washing was carried out on the black mass for selective lithium recovery. The gained lithium product was analyzed for the purity and phases present. The addition of O₂ resulted in reduced reduction reactions of lithium metal oxides and lower Li-yields in the water leaching compared to the other two atmospheres. In the case of CO₂, the formation of Li₂CO₃ is favored compared to LiF, but the Li-yield of 56% is comparable to N₂ treatment.

Shoppert et al. [Contribution 8] studied and published research on ‘Enhanced Precipitation of Gibbsite from Sodium Aluminate Solution by Adding Agglomerated Active Al(OH)₃ Seed’. The addition of an active seed for increasing the precipitation rate leads to the formation of fine Al(OH)₃ particles that complicate the separation of the solid from the mother liquor. In this study, the enhanced precipitation of coarse Al(OH)₃ from sodium aluminate solution using active agglomerated seed was investigated. Aluminum salt (Al₂(SO₄)₃) was used for active agglomerated seed precipitation at the initial stage of the process. About 50% of the precipitation rate was obtained when these agglomerates were used as a seed in the amount of 20 g L⁻¹ at 25 °C within 10 h. The agglomerated active seed and precipitate samples were characterized using X-ray diffraction (XRD), scanning electron microscopy (SEM), and Fourier-transform infrared spectroscopy (FTIR). SEM images showed that agglomerates consist of flake-like particles that can be stuck together with bayerite (β-Al(OH)₃) acting as a binder. The precipitation temperature above 35 °C and the high concentration of free alkali (αk = 1.645Na2_Ok/Al₂O₃ > 3) led to the refinement of the agglomerates, which can be associated with the bayerite dissolution.

Norgren and Anderson [Contribution 9] provided a rare earth beneficiation-focused article entitled [Contribution 9] ‘Ultra-Fine Centrifugal Concentration of Bastnaesite Ore’. Historically, the ability to effectively separate carbonate gangue from bastnaesite via flotation has frequently proven to be challenging without sacrificing significant rare earth oxide (REO) grades or recovery. However, because the rare-earth-bearing minerals often exhibit higher specific gravities than the carbonate gangue, the possibility exists that the use of gravity separation could be used to achieve such a selective separation. This, however, is complicated by the fact that, in cases such as this study, when the liberation size is finer than 50 µm, most traditional gravity separation methods become increasingly challenging. The purpose of this study is to determine the applicability of gravity concentrators to beneficiate bastnaesite from deleterious calcite-bearing flotation feed material. With the use of a UF Falcon, it was possible to achieve rougher gravity REO recoveries approaching the upper 80% range while rejecting about 30% of the total calcium. In terms of pure REO recovery, this represents a significant improvement over the results obtained using a traditional Falcon in previously reported studies.

Norgren and Anderson [Contribution 10] also provided a second companion article, also focused on rare earth beneficiation, for the ‘Recovery of Rare Earth Oxides from Flotation Concentrates of Bastnaesite Ore by Ultra-Fine Centrifugal Concentration’. Historically, the ability to effectively separate carbonate gangue from bastnaesite via flotation has frequently proven to be challenging without sacrificing significant rare earth oxide (REO) grades or recovery. However, because the rare-earth-bearing minerals often exhibit higher specific gravities than the carbonate gangue, the possibility exists that the use of gravity separation could be used to achieve such a selective separation. This, however, is complicated by the fact that, in cases such as this study, when the liberation size is finer than 50 microns, most traditional gravity separation methods become increasingly challenging. The aim of this study is to determine the applicability of centrifugal concentrators to beneficiate ultra-fine (UF) bastnaesite and calcite-bearing flotation concentrates. By using a UF Falcon, it was possible to achieve initial gravity REO recoveries exceeding 90% while rejecting, on the order of 25% to 35%, the total calcium from an assortment of rougher and cleaner flotation concentrates. Additionally, when additional stages of cleaner UF Falcon gravity separation were operated in an open circuit configuration, it was possible, from an original fine feed of 35 microns containing 50.5% REO and 5.5% Ca, to upgrade to up to approximately 59% REO and 2.0% calcium. While not the goal of this study, these results also support previous limited data that suggest that UF Falcons are potentially capable of treating a wider range of materials than they were originally designed for, including feeds rich in heavy mineral content.

3. Summary and Outlook

This second Special Issue of Metals was again well supported by diverse submissions and a final book publication of ten high-quality peer-reviewed articles. Due to this continued success, another Special Issue (“Advances in Mineral Processing and Hydrometallurgy III”, website: https://www.mdpi.com/journal/metals/special_issues/9259176SUL (accessed on 20 November 2024)) has been commissioned as a follow-up to accept continued global contributions from the hydrometallurgy and mineral processing community.