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Solvent extraction scheme to isolate Zn, Cu, and Ni from glutamate media as hypothetical product of electric arc furnace dust alkaline leaching was developed, with concentrations being 17.6, 0.35, and 0.14 g/L, respectively. Three extractants were investigated: Cyanex 272, DEHPA, and Acorga M5640 selectively separated Zn, Ni, and Cu from each other. Aside from pH, parameters investigated included O/A ratio, extractant concentration in kerosene as diluent, and sulfuric acid concentration as stripping agent. pH was the most critical factor in determining the separation factor among three metals since pH controlled metal speciation in pregnant leach solution and regulated the interaction between metal ions and extractant. Based on previous studies, a flowsheet for Zn, Cu, and Ni isolation is proposed, which obeys the following sequence: Zn separated from Cu and Ni by Cyanex 272 (in pregnant leach solution at pH 8, log separation factor Zn–Cu 4.78, Zn–Ni 2.51), followed by Ni separation from Cu by DEHPA (in raffinate at pH 7, log separation factor Ni–Cu 3.92), and finally Cu extraction (in raffinate at pH 4, log distribution coefficient 3.19). Sulfuric acid was proved to be a suitable stripping agent with optimum concentrations of Zn, Ni, and Cu being 0.5, 0.25, and 2 M, respectively.

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) grade or recovery. However, in light of the fact that 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% of 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 up to approximately 59% REO and 2.0% calcium. While not the goal of this study, these results also support previous limited data to 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.

The demand for rare earth elements for everyday technology and applications has initiated much research into the extraction and recovery of these rare earth elements. An otherwise unknown group of minerals, eudialyte is a zircon silicate consisting of rare earth oxides (REO), contained in a unique mineral structure and represented by a complex chemical formula. Research into extraction and recovery of rare earth elements involves various beneficiation and hydrometallurgical techniques. The goal of beneficiation being to efficiently liberate and upgrade the REO content, followed by hydrometallurgy to extract the desired elements. This overview will discuss previous research on treating eudialyte group minerals via different mineral processing and extraction techniques.

The recovery of platinum group elements (PGE (platinum group element coating), Pd, Pt, and Rh) from used catalytic converters, using low energy and fewer chemicals, was developed using potassium bisulfate fusion pretreatment, and subsequently leached using hydrochloric acid. In the fusion pre-treatment, potassium bisulfate alone (without the addition of an oxidant) proved to be an effective and selective fusing agent. It altered PGE into a more soluble species and did not react with the cordierite support, based on X-Ray Diffraction (XRD) and metallographic characterization results. The fusion efficacy was due to the transformation of bisulfate into pyrosulfate, which is capable of oxidizing PGE. However, the introduction of potassium through the fusing agent proved to be detrimental, in general, since potassium formed insoluble potassium PGE chloro-complexes during leaching (decreasing the recovery) and required higher HCl concentration and a higher leaching temperature to restore the solubility. Optimization on the fusion and leaching parameter resulted in 106% &plusmn, 1.7%, 93.3% &plusmn, 0.6%, and 94.3% &plusmn, 3.9% recovery for Pd, Pt, and Rh, respectively. These results were achieved at fusion conditions: temperature 550 &deg, C, potassium bisulfate/raw material mass ratio 2.5, and fusion time within 30 min. The leaching conditions were: HCl concentration 5 M, temperature 80 &deg, C, and time within 20 min.

This paper demonstrates the recovery of valuable metals from shredded Waste Printed Circuit Boards (WPCBs) by bromine leaching. Effects of sodium bromide concentration, bromine concentration, leaching time and inorganic acids were investigated. The most critical factors are sodium concentration and bromine concentration. It was found that more than 95% of copper, silver, lead, gold and nickel could be dissolved simultaneously under the optimal conditions: 50 g/L solid/liquid ratio, 1.17 M NaBr, 0.77 M Br2, 2 M HCl, 400 RPM agitation speed and 23.5 &deg, C for 10 hours. The study shows that the dissolution of gold from waste printed circuit boards in a Br2-NaBr system is controlled by film diffusion and chemical reaction.

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) grade or recovery. However, in light of the fact that 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 purposes of this study is to determine the applicability of gravity concentrators to beneficiate bastnaesite from deleterious calcite bearing flotation feed material. Via the use of a UF Falcon, it was possible to achieve rougher gravity REO recoveries approaching the upper 80% range while rejecting on the order of 30% of the total calcium. In terms of purely REO recovery, this represents a significant improvement over results obtained via a traditional Falcon in previously reported studies.

The Bear Lodge Project is an important rare earth deposit in the United States. While extensive previous studies have settled on a crushing, screening, gravity and magnetic separation process, this study sets out to investigate an alternative flowsheet based on flotation and wet high intensity magnetic separation (WHIMS) to effectively beneficiate the rare earth oxide (REO) content of the Bear Lodge ore. Mineral characterization found ancylite was the dominant rare earth mineral, associated mainly with calcite and strontianite. Electrokinetic studies on the effects of pH, concentrations of various ions (Sr2+, HCO3− and CO32−) and hydroxamate concentrations were performed to establish the electric nature of the Bear Lodge ore. The isoelectric points (I.E.P) of the material in distilled water was around 5.27. The Sr2+ and CO32− ions in solution significantly affected the surface charge of the material. Adsorption studies suggested that the mechanism of hydroxamate adsorption is chemisorption, as hydroxamate adsorption increased on the Bear Lodge ore with an increase in temperature. WHIMS was employed to remove the iron content to reduce the interference of iron in following flotation process and consumption of hydroxamic acid. After cleaner flotation a REO grade of 11.2% at 72.7% recovery from a feed material of 4.5% REO was obtained. In light of the loss of REO in WHIMS process, it is possible to produce a concentrate containing 11.2% REO grade at 61.2% recovery.

Flash furnace electrostatic precipitator dust (FF-ESP dust) is a recycle stream in some primary copper production facilities. This dust contains high amounts of copper. In some cases, the FF-ESP dust contains elevated levels of bismuth and arsenic, both of which cause problems during the electrorefining stages of copper production. Because of this, methods for separation of copper from bismuth and arsenic in FF-ESP dust are necessary. Hydrometallurgical leaching using a number of lixiviants, including sulfuric acid, sulfurous acid, sodium hydroxide, and water, were explored. Pourbaix diagrams of copper, bismuth, and arsenic were used to determine sets of conditions which would thermodynamically separate copper from bismuth and arsenic. The data indicate that water provides the best overall separation between copper and both bismuth and arsenic. Sodium hydroxide provided a separation between copper and arsenic. Sulfurous acid provided a separation between copper and bismuth. Sulfuric acid did not provide any separations between copper and bismuth or copper and arsenic.

Because of their selective depressing power, sodium silicate and lignosulfonate have been widely used as depressants in rare-earth mineral flotation to separate minerals from specific types of gangue minerals. In this project, the microflotation of a xenotime pre-concentrate and pure samples of the selected gangue minerals ilmenite, zircon, schorl and staurolite was carried out. This utilized octano-hydroxamic acid and sodium oleate as the collectors and the research was conducted in a Partridge-Smith microflotation cell. The flotation of the mixed samples (weight ratio = 1:1) of xenotime and each one of its gangue minerals was also investigated at both room temperature and 80 °C, using sodium silicate or ammonium lignosulfonate in the presence of octano-hydroxamic acid and sodium oleate respectively. The flotation results are described and compared with those observed by previous researchers. The effects of sodium silicate and ammonium lignosulfonate on weight recoveries and grade of xenotime are also discussed and compared.

The fundamentals of the surface chemistry of ancylite, strontianite, and calcite in the presence of hydroxamic acid (HXY) were investigated based on their zeta potential, adsorption, infrared measurement, and microflotation. Zeta potential studies indicate that the isoelectric points of ancylite, strontianite, and calcite are around 5.46, 4.50, and 5.50, respectively. HXY is chemically adsorbed onto the surface of ancylite, which was confirmed by both zeta potential and infrared measurements. At room temperature, the monolayer coverage of HXY on ancylite is shown as 20 µmol/m2, which is much higher than the monolayer coverages for strontianite and calcite. In the comparison of adsorption densities of strontianite, calcite, and ancylite at both room temperature and 50 °C, the results show that strontianite and calcite appear more sensitive to temperature than ancylite. Microflotation studies of pure minerals show that theoretically, calcite could be separated from strontianite and ancylite at pH 7.5 in the presence of 5 × 10−4 M HXY, and ancylite can be separated from strontianite in the presence of 2 × 10−4 M HXY when pH is around 9. However, as indicated from the zeta potential results, the dissolved species from minerals significantly change the flotation behavior of minerals’ mixture. Thus, a successful flotation separation could not be achieved without any modifiers.

Recently, leaching of gold with cyanide has come under scrutiny for its possible detrimental effects to the environment. Also, in some gold bearing ores and concentrates, conventional cyanide leaching is not the best method for gold extraction. This is due to gold particle encapsulation, cyanicides or pregrobbing carbon such as found in refractory ores and concentrates. Hence, there is much interest in developing an alternative way to extract gold from ore. Possible benefits of leaching gold with something other than cyanide include easier processing of refractory ore bodies including: carbonaceous, cyanicides, and sulfides, cheaper reagent and operating costs and opening of mine properties in places where cyanide use has been limited by the law. Hence, this paper details the fundamentals and development of the Alkaline Sulfide Gold Leaching system.

Summary Separating rare-earth-element-rich minerals from unwanted gangue in mined ores relies on selective binding of collector molecules at the interface to facilitate froth flotation. Salicylhydroxamic acid (SHA) exhibits enhanced selectivity for bastnäsite over calcite in microflotation experiments. Through a multifaceted approach, leveraging density functional theory calculations, and advanced spectroscopic methods, we provide molecular-level mechanistic insight to this selectivity. The hydroxamic acid moiety introduces strong interactions at metal-atom surface sites and hinders subsurface-cation stabilization at vacancy-defect sites, in calcite especially. Resulting from hydrogen-bond-induced interactions, SHA lies flat on the bastnäsite surface and shows a tendency for multilayer formation at high coverages. In this conformation, SHA complexation with bastnäsite metal ions is stabilized, leading to advanced flotation performance. In contrast, SHA lies perpendicular to the calcite surface due to a difference in cationic spacing. We anticipate that these insights will motivate rational design and selection of future collector molecules for enhanced ore beneficiation., Graphical Abstract, Highlights • Salicylhydroxamic acid (SHA) is an effective collector for flotation of bastnäsite • DFT calculations show that the phenol group anchors SHA flat to bastnäsite surface • Spectroscopic results suggest SHA multilayers on bastnäsite at high coverage • This peculiar adsorption explains efficacy of SHA in bastnäsite flotation from ores, Chemical Engineering; Spectroscopy; Physical Inorganic Chemistry; Surface Chemistry

Disclosed herein is a treated ore solid comprising a reduced amount of a contaminant, for example arsenic, compared to the ore solid prior to treatment. Also disclosed are temperature and pressure modifications, parameters, and methods for treating an ore solid by pressure oxidation leaching of enargite concentrates. The disclosed methods and processes may be applied to copper sulfide orebodies and concentrates containing arsenic. In some cases, the disclosed methods and systems extract, remove, or reduce contaminants, for example arsenic, from an ore containing solution at moderately increased temperature, pressure, and oxygen concentration, and in the presence of an acid.

Today, with a stringent economic and environmental climate prevailing in the copper business, there is increased interest in evaluating new processing alternatives for production. Hydrometallurgical pressure oxidation of copper concentrates is one of the more viable approaches, and several technological candidates have emerged. Of these, an overlooked but, ironically, the first industrially proven methodology utilized nitrogen species catalyzation in the oxidizing pressure-leach system to produce copper via solvent extraction/electrowinning. Given its advantages, this may prove to be a feasible process alternative for the future. In this article, the history of the system and its application to copper concentrates and ores will be outlined. In particular, a non-cyanide methodology for effective recovery of precious metals from chalcopyrite concentrates will be discussed.

Copper sulfides are generally treated by flotation followed by smelting. However, concerns over air pollution have made governments tighten their regulations regarding the emission of sulfur dioxide. Thus, hydrometallurgical techniques seem to be the most convenient methods to selectively extract non-ferrous metals. In this study, the leaching kinetics of copper from low-grade concentrate by nitrite salt was investigated. The probabilities of several chemical reactions and the products formed were explained based on literature data and thermodynamic analysis. The copper concentrate was obtained from a flotation plant in the Kastamonu region of Turkey. Leaching experiments were conducted in sulfuric acid electrolyte (0.5 M-2.0 M) containing nitrite compounds (0.025 M-0.15 M) at a moderate temperature (80-120 degrees C). A high level of copper recovery (96%) from low-grade chalcopyrite concentrate was obtained under a total pressure of 6 atm at 120 degrees C within 2 h when using small amounts of nitrite species due to their autocatalytic behavior in acidic solutions. Elemental sulfur was the primary leaching product on the mineral surface, as confirmed using XRD and SEM/EDX analysis, and the formation of the sulfur layer inhibited leaching rate at later stage of dissolution process. (c) 2013 Elsevier B.V. All rights reserved.

A new process for smelting copper, Looping Sulfide OxidationTM (LSOTM) significantly and positively alters the smelting energy balance. This process uses copper oxide in conjunction with air for the desulfurization of copper and chalcopyrite concentrates. Copper oxide is regenerated in a flash type furnace and then “looped” back for the oxidation of the copper concentrate. Heat and material balances show that the Looping Sulfide OxidationTM process requires significantly lower net energy than conventional technology. This process produces anode copper and higher SO2 content off-gases for efficient formation of sulfuric acid. Additionally, the process yields slags with lower copper solubility as compared to the conventional technology.

Since 1984, nitrogen-species catalyzed pressure leaching has been utilized at the Sunshine Mining and Refining Company for the recovery of precious metals from a refractory sulfide concentrate. The process operates without the use of cyanide. The plant is now the only industrial-scale commercial application of this technology. Since the inception of the process, changes have been implemented to improve efficiencies and throughputs. Two of these changes, the replacement of nitric acid by sodium nitrite anda concentrate regrind, have recently been implemented. This paper will discuss the outcomes of these recently patented plant-process changes. In particular, a comparison between the previously used method and the new leach method is made. Also discussed is the potential applicability of the process to other materials. A companion publication (Anderson, Harrison and Krys, 1996) indicates possible mechanisms responsible for the success of these process changes.

Since 1984, nitrogen-species catalyzed pressure leaching has been utilized at Sunshine Mining and Refining Company for the recovery of precious metals from a refractory sulfide concentrate. The process operates without the use of cyanide. The plant is now the only industrial-scale commercial application of this technology. Since the inception of the process, changes have been implemented to improve efficiencies and throughputs. Two of these changes, the replacement of nitric acid by sodium nitrite and a concentrate regrind, have recently been implemented. This paper discusses the development and potential mechanisms involved in these recently patented plant-process changes. A companion publication (Anderson et al., 1996) covers the plant integration and potential applicability of the technology.

The use of nitric acid in metal sulfide oxidation is not new. Many derivations of the technology have been researched and piloted for applications such as copper, nickel and precious metals. Again, as noted in Table 1, and as specifically applied to copper concentrates and precious metals, only the nitrogen species catalyzed (NSC) acid pressure leach has ever been built and operated successfully on an industrial scale [1].