Pressure Oxidative Leaching Research Articles

Recovery of Gallium from Smartphones—Part II: Oxidative Alkaline Pressure Leaching of Gallium from Pyrolysis Residue

In this article, we examine the selective hydrometallurgical extraction of gallium from pyrolyzed smartphones. Gallium-enriched pyrolysis residue originating from pyrolyzed smartphones was leached using NaOH and gaseous oxygen at elevated temperatures and pressures. The high content of organic carbon in the material strongly influenced the leaching performance. Oxygen, which is indispensable for the dissolution of gallium, also oxidized the organic carbon in the feed so that CO2 was released, which had a neutralizing effect on the alkaline solution. As a result, the CO2 formation complicated the accurate process control as the leaching temperature increased. The highest gallium yield of 82% was obtained at 180 °C, 5 g/L NaOH and 5 bar oxygen pressure. Decreased temperatures, NaOH concentrations and oxygen pressures resulted in lower leaching yields but with a higher selectivity for Ga. Temperatures higher than 180 °C resulted in extensive carbon oxidation, NaOH consumption and the coextraction of Cu and Ag. We propose that those conditions also facilitated the formation of water-soluble organic compounds, which would also influence the metal dissolution.

On the Assumed Mechanism of the Pressure Oxidation of Pyrrhotine during the Pressure Oxidation Leaching of a Low-Nickel Pyrrhotine Product Using Iron(III) Ions as an Oxidant

Abstract—The influence of the Fe(III) content in the solutions of the sulfuric acid leaching of a pyrrhotine concentrate on the features of the reaction mechanism of its desulfurization is analyzed. Owing to the hydrolysis of ferrisulfate (iron(III) sulfate) in the solution, the active surface of the concentrate is blocked by the formed basic iron hydrosulfate, whose chemical decomposition is accompanied by the oxidation of pyrrhotine sulfur to form elemental sulfur. The organization of the two-stage direct-flow leaching of the pyrrhotine concentrate is shown to be important for extracting nonferrous metals and components of the gangue to the pyrrhotine concentrate and fabricating a concentrated cake containing nickel, copper, and cobalt.

Comprehensive recovery of Sn-Cu bearing residue and preparation of high purity SnO2 and CuSO4·5H2O

In the production of tin, much Sn-Cu bearing residue is yielded, and it is an important secondary resource for Sn and Cu recovery but has the disadvantages of a low metal recovery rate and low-concentration SO2 pollution. In this study, alkaline pressure oxidative leaching (APOL) was proposed to dispose of the Sn-Cu bearing residue. Tin was first dissolved as sodium stannate into the filtrate and used for SnO2 synthesis via a hydrothermal process, while copper sulfides were oxidized to copper oxide, and used for CuSO4·5H2O preparation via acid leaching-evaporative crystallization. Under the optimized leaching conditions of APOL, the leaching efficiencies of Sn, As, and S approached 93.42%, 98.82%, and 97.99%, while the leaching efficiency of copper was only 1.60%, which results in the efficient extraction of tin, desulfurization, dearsenification, and separation of tin and copper. The product SnO2 had a purity higher than 99.86%, a morphology of irregular shape, and a mean size of 2.68 μm. Meanwhile, the purity of prepared CuSO4·5H2O reached 99.98% with a morphology of irregular rods. Both valuable metals are economically and efficiently recycled by this environmentally friendly method.

Substantiation of a Combined Technology for the Hydrometallurgical Beneficiation of a Pyrrhotine-Containing Charge Based on Pressure Oxidation Leaching Using Ferric Sulfate as a Pyrrhotine Oxidizer

Two series of comparative experiments are carried out to select the most effective version of pressure oxidation technology (POT) for processing a pyrrhotine-containing charge (PCC) consisting of a mixture of ELNPP TCP and NPC NCP (enriched low-nickel pyrrhotine product and nickel-pyrrhotine concentrate). In the first version (basic), a standard oxidizing agent (oxygen) is used for pressure oxidation leaching (POL); in the second version, iron(III) sulfate is used. The use of iron(III) sulfate in comparison with oxygen makes it possible to significantly increase all the main indicators of the hydrometallurgical beneficiation of PCC. The substitution of iron(III) sulfate for oxygen in the POL operation after optimizing the operation parameters of the “head” POT operations is shown to achieve a high quality of the autoclave sulfide concentrate (ASC; 13–14% Ni) and the completeness of extraction of valuable components into it, namely, 96–97% nickel and 93–95% PGMs. An important competitive advantage of POL processes using iron(III) sulfate is that the formation of a bulky precipitate of iron hydroxides is almost completely excluded during the oxidation of pyrrhotine. In the version of “oxygen” leaching of PCC, it is the formation of this precipitate that causes a high level of the PGM losses with waste technological tailings technology and limits the possibility of improving the quality of ASC. A schematic diagram of POT is proposed for the POL-based hydrometallurgical beneficiation of PCC (ELNPP TCF, NPC NCP, SPM TCF, etc.) using iron(III) sulfate. This scheme can be implemented on the basis of the existing fleet of equipment for the hydrometallurgical production of the Nadezhda metallurgical plant at the minimum capital costs

Cyanide-free recovery of metals from polymetallic and refractory gold concentrates by the KellGold process

A patented hydrometallurgical approach (KellGold) for high-yield, cyanide-free recovery of metals from refractory gold and polymetallic concentrates is summarized. Drivers include elimination of atmospheric emissions of SO2, As2O3 and Hg produced by smelters and roasters, as well as cyanide risks arising from transportation, operations and detoxification activities.The Kell process comprises several commercially proven unit operations. Sulphur and other soluble metals are first selectively removed by use of a pressure oxidation leaching step during which the dissolution of precious metals such as Au, Ag and platinum group metals (PGM) is minimized. The residue from pressure oxidation may be subjected to further conditioning by HCl leaching and/or thermal treatment to ensure efficient precious metals recovery by subsequent chlorination leaching from a clean leachate. Subsequent metal recovery steps provide marketable end products close to the mine site.Testwork on a range of concentrate types typically shows >95% recovery of gold, silver, copper, cobalt, zinc, lead, antimony, and, if present in economic quantities, other value elements, including PGM (Pt, Pd, Rh, Ir, Ru). Engineering studies have been successfully completed for two KellGold applications, while KellPGM studies have been completed to definitive feasibility level. Centralized concentrate processing plants are under active consideration in several countries. KellPGM has been demonstrated in a nine-week integrated pilot campaign at 1:1000 scale, supporting a bankable feasibility study for a ~ 1110 ktpa plant now at the financing stage, treating a PGM concentrate at a South African PGM mine site.

Hydrothermal pretreatment of chalcopyrite concentrate with copper sulfate solution

The existing technologies for porphyry copper ores beneficiation, located in deposits in the Urals of Russia, allow the production of chalcopyrite concentrates with a copper content of 19 to 30%. Subsequent technology for the concentrates processing includes autogenous smelting, matte desulfurization, and blister copper refining operations. The current work presents a new hydrometallurgical approach for the copper concentrate enrichment up to a higher grade of 50–55% Cu, which could significantly increase the production capacity and efficiency of smelting and with a process simultaneously avoiding the matte desulphurization operation. The suggested method consists of oxidative pressure leaching followed by hydrothermal pretreatment of chalcopyrite concentrates with copper sulfate solutions. The pressure oxidative leaching of the chalcopyrite concentrates allows Cu extractions up to 99% and produces Cu-rich solution for the hydrothermal pretreatment stage. Optimization of the hydrothermal pretreatment stage was carried out within the following limits: T = 150–230 °C, CuCuSO4/CuCuFeS2 = 1.2–3.3, H2SO4 = 20–120 g/L. Recommended parameters were established as T = 230 °C, CuCuSO4/CuCuFeS2 = 1.8–1.9, H2SO4 = 20–40 g/L, t = 60–100 min since this allows concentrate to be obtained with Cu content of 51–52% and residual Cu concentrations of as low as 0.08–0.13 g/L in the final solution. A parametric model of the hydrothermal pretreatment stage was obtained. The solutions and residues after hydrothermal pretreatment of the material were analyzed for their chemical and phase composition. A technological flowsheet, as well as strategies for by-product processing, were also suggested.

Pressure oxidative leaching of chalcopyrite concentrates: influence of the process temperature on cakes cyanidation efficiency

Pressure oxidative leaching of chalcopyrite concentrates: influence of the process temperature on cakes cyanidation efficiency

Direct oxidative pressure leaching of bismuth sulfide concentrate in methanesulfonic acid medium

As the prevalent extraction technology of bismuth sulfide concentrate, the precipitation smelting process has disadvantages of SO2 emission and inefficient recovery of associated metals. Hydrometallurgical alternatives to traditional pyrometallurgical techniques have obtained continuously growing attention in recent years. In this paper, methanesulfonic acid (MSA) was proposed as the reaction medium to extract bismuth from bismuth sulfide concentrate under pressure leaching conditions. It was found that more than 95% of Bi could be leached out in 2.70 mol/L MSA + 28.6 g/L ferrous ion solution at 110 °C and 0.4 MPa oxygen pressure for 2 h. Cu and Fe from chalcopyrite and pyrite can be slightly leached out with ratio of 26.67% and 30.51%, respectively. Scheelite and molybdenite basically remained in the leaching residue. Galena was partially dissolved to generate lead ion and elemental sulfur, which could be further oxidized to sulfate as the leaching temperature and oxygen pressure raised, and then reacted with dissolved lead ion to form insoluble lead sulfate. The produced leaching residue was mainly composed of sulfur, lead sulphate, quartz, molybdenite, and unleached chalcopyrite and pyrite. Compared with the traditional FeCl3, HCl-NaClO3, HNO3 system, the proposed MSA medium is proved to be biodegradable, low toxic, less volatile, and thus has broad application prospects in bismuth green hydrometallurgy.

Autoclave treatment of cakes after pressure oxidation leaching of chalcopyrite concentrates

The existing technologies for copper-porphyry ores enrichment, located in deposits in the Urals of Russia, allow the production of chalcopyrite concentrates of the following composition, %: 21.5 Cu, 24.5 Fe, 26.5 S, 0.4 Pb, 17.6 SiO2, 1.8 CaO, 2–6 Au (ppm), 20– 40 Ag (ppm). A conventional technology for processing such concentrates includes autogenous smelting, matte desulfurization and blister copper refining. Pressure oxidation leaching (POX) is considered the most promising alternative technology for chalcopyrite concentrate processing. The POX of concentrates originated from Mikheevskii GOK allow the production a cake of the following chemical composition, %: 56–65 Fe2O3, 25–30 SiO2, 2.7 Ca, 0.3–1.0 Cu, 2–7 S, 0.6–0.8 Pb, 4–12 Au (ppm), 40–80 Ag (ppm); mass loss was 37–45 %. A standard method of cake cyaniding provides satisfactory indicators of precious metal extraction, but it requires a cumbersome area to be arranged for their processing and offers no solution for residue disposal. In this regard, this paper investigates the method of subsequent cake processing using autoclave treatment (AT) for iron removal. The study shows how the following parameters affect the results of this process: t = 110÷210 °C, H2SO4 = 15÷60 g/dm3, τ = 45÷ ÷120 min. A statistic description of the AT operation is developed. Recommended AT conditions (t = 110 °C, H2SO4 = 60 g/dm3, τ = 60÷100 min) allow to obtain the POX cake yield reduced to 30–35 % of the source material with the following composition, %: 28–33 Fe2O3, 47–53 SiO2, 2–5 Ca, 0.6–2.0 Cu, 0.8–1.5 Pb, 2–8 S. At the same time, the content of precious metals in the cake reaches 12–16 Au (ppm) and 80–120 Ag (ppm). Options for using AT products are proposed.

Kinetics of copper extraction from copper smelting slag by pressure oxidative leaching with sulfuric acid

The accumulation of copper smelting slag not only pollutes the environment but also wastes resources, using oxygen pressure acid leaching process to treat copper slag has incomparable advantages over other processes, but there are few researches reported at home and abroad. In this paper, copper slag was treated by this process. Under the premise of fully studying the influence of various factors on copper dissolution, the selective leaching behavior and kinetics of copper oxy-pressure acid system in copper slag were expounded. The experimental results showed that the dissolution kinetics analysis of the experimental data based on the shrinking core model for various conditions indicated that the reaction rate of leaching was mainly controlled by the chemical reaction during its early stages, then switched to be controlled by mixed chemical-reaction and product-layer diffusion which could be ignored because of its short lifetime, and finally was controlled solely by diffusion through a surface product layer in the later stage. The activation energy was calculated to be 47.03 kJ·mol/L in the early surface chemical reaction controlling stage and 11.35 kJ·mol/L in the later diffusion controlling stage, respectively. The leaching rate of copper is mainly determined by the chemical reaction control stage, in which the reaction orders of sulfuric acid concentration, oxygen partial pressure and particle size are 0.83, 0.92 and −1, respectively.

Pressure oxidative leaching of copper cake from nickel refining plant in sulfuric acid media

Pressure oxidative leaching of copper cake from nickel refining plant in sulfuric acid media

Pressure Leaching of Chalcopyrite Concentrate: Ultra-Fine Milling as Process Acceleration Method

Ural enrichment plants processing copper-porphyry deposits are produced chalcopyrite concentrates of the following chemical composition, %: 21.5 Cu, 0.1 Zn, 26.59 S, 24.52 Fe, 0.05 Pb, 0.04 Ni, 16.28 SiO2 [1]. According to previous studies [2, 3, 21], such concentrates can be effectively treated using pressure oxidative leaching (POX). At temperature of 190–200 °C, partial oxygen pressure 4–6 bar and [H2SO4] = 15–30 g / L during 100-120 min about 98 % Cu was extracted. In this paper, an inluence of chalcopyrite concentrate preliminary grinding on efficiency of the POX stage was studied. It was found that even the finest grinding of the particle size class P80 = 13 μm does not lead to significant intensification of process kinetics.

Sodium Lignosulfonate and Sodium Dodecyl-Sulfate Mixtures Influence on Zinc Concentrate Pressure Leaching and Zinc Electro-Winning

This paper is describing an investigation of sodium lingo-sulfonate and sodium dodecyl-sulfate mixtures influence on zinc concentrates high temperature oxidative pressure leaching and zinc electro-winning. For this purpose, surfactants concentration at leaching tests was varied from 200 to 800 mg∙l-1. It was established that the maximum zinc extraction (99 %) at leaching was achieved in the presence of mixture containing 800 mg∙l-1 lignosulfonate and 200 mg∙l-1 sodium dodecyl-sulfate. Therefore, this mixture can be recommended for high temperature oxidative pressure leaching of zinc concentrates. Sulfur-sulfide pellets formation also was observed at a low lingo-sulfonate concentration (200 mg∙l-1) in a mixture with sodium dodecyl-sulfate. This phenomenon can lead to emergency shut down of autoclave. It was observed that the mixture usage of 800 mg∙l-1 lignosulfonate and 200 mg∙l-1 sodium dodecyl-sulfate had no significant impact on zinc current efficiency, it was in the rage of 92-93 %. The mixture usage of 200 mg∙l-1 lignosulfonate and 600 mg∙l-1 sodium dodecyl-sulfate allowed to increase current efficiency up to 95 %. Increasing sodium dodecyl-sulfate concentration in mixtures with lignosulfonates leads to decrease of current efficiency, to formation of deep pores and defects on cathode zinc surface.

Pressure oxidation leaching of the chalcopyrite concentrate from Mikheevsky Mining and Processing Plant in sulphuric acid media

Pressure oxidation leaching of the chalcopyrite concentrate from Mikheevsky Mining and Processing Plant in sulphuric acid media

Oxredmetry as a method to ensure optimum distribution of precious metals in pressure oxidation leaching of pyrrhotite material depending on the type of surfactant used

Oxredmetry as a method to ensure optimum distribution of precious metals in pressure oxidation leaching of pyrrhotite material depending on the type of surfactant used

Syncytiotrophoblast basal membrane-derived exosomes increase NO generation in human umbilical vein endothelial cells

Rhenium and molybdenum are valuable metals present in the earth's crust in very low concentration and are increasingly used in industry thanks to their multiple properties, but due to the harmful aspects of the production process, there is a need to develop systems of concentration or purification of salts of these elements. For this work a PEG/Na2MoO4 aqueous two phase system (ATPS) was used for the partition of ReO4- anion. By using a central composite design, the effects of PEG concentration, sodium molybdate concentration, temperature, pH and their possible interaction on the partition coefficient were studied, also to estimate the optimal configuration of the factors that influence the process. At the same time, experimental density, sound velocity and isentropic compressibility were also obtained for top and bottom phases. The results showed that the optimal conditions for the partition of ReO4− anion using the ATPS formed by PEG and Na2MoO4 were 14.7 wt% PEG 4000, 16.3 wt% Na2MoO4, 35 °C and pH 6. For these conditions, the partition coefficient (K) obtained was up to 17.8 ± 0.78. The order of factors with high significance on the partition coefficient was: Na2MoO4 concentration > PEG concentration > temperature > pH > interaction between Na2MoO4 and PEG concentrations; the effect of these parameters mainly is related with the speciation, salting-out effect and hydrophobicity. The influence of these factors on the density, and sound velocity of the top phase was minimal, unlike its influence on the properties of the bottom phase, not isentropic compressibility and effect of PEG and Na2MoO4 concentrations on sound velocity. The evaluation of the partitioning behavior was made by experimental design, applying statistical regression analysis and analysis of variance (ANOVA). The statistical models for partition coefficient and for phase densities (and the other properties) as function of parameters with significant effects were obtained, and also a good correlation between experimental and correlated data was observed (R-Squared = 0.9379, ReO4− partition coefficient; 0.9882, bottom phase; and 0.9836 upper phase). This work is oriented to the possible application in stripping of a top phase provided for a co-extraction of rhenium and molybdenum by ATPS in acidic media or in solutions with high contend of Mo such as those provided by pressure oxidative leaching of molybdenite in basic media.

Color of Oxidative Pressure Leaching Liquor: A Case of Sphalerite

In lab-scale experiments of oxidative pressure leaching of sphalerite, an interesting phenomenon is that the leaching liquor looks three main colors: black, red and green. In on-going work, semi-quantitative study is carried out to investigate such phenomenon, including drawing of predominance area color diagram, effect of sulfur surfactant on color of leaching liquor, color and stability of oxidation of leaching liquor. Experimental results show that color of the leaching liquor depends on leaching condition mainly including temperature, dissolved oxygen concentration, that experimental results show sulfur surfactant has no effect on the color of leaching liquor, that, with increased molar ratio of elemental sulfur and sodium sulfide, , x = 1 ∼ 8 is more stable, and that in oxidation of aqueous sulfide anion, to achieve reasonable fast oxidation kinetic, temperature of 150 °C and oxygen pressure of 2 × 105 Pa is the minimum required conditions, which is coincide with such conditions to achieve reasonable fast oxidation kinetic in oxidation of aqueous ferrous iron. In this paper, different reaction mechanisms in given leaching condition were discussed.

Alkaline pressure oxidative leaching of bismuth-rich and arsenic-rich lead anode slime

A new alkaline pressure oxidative leaching process (with NaNO3 as the oxidant and NaOH as the alkaline reagent) is proposed herein to remove arsenic, antimony, and lead from bismuth-rich and arsenic-rich lead anode slime for bismuth, gold, and silver enrichment. The effects of the temperature, liquid-to-solid ratio, leaching time, and reagent concentration on the leaching ratios of arsenic, antimony, and lead were investigated to identify the optimum leaching conditions. The experimental results under optimized conditions indicate that the average leaching ratios of arsenic, antimony and lead are 95.36%, 79.98%, 63.08%, respectively. X-ray diffraction analysis indicated that the leaching residue is composed of Bi, Bi2O3, Pb2Sb2O7, and trace amounts of NaSb(OH)6. Arsenic, antimony, and lead are thus separated from lead anode slime as Na3AsO4· 10H2O and Pb2Sb2O7. Scanning electron microscopy and energy-dispersive spectrometry imaging revealed that the samples undergo appreciable changes in their surface morphology during leaching and that the majority of arsenic, lead, and antimony is removed. X-ray photoelectron spectroscopy was used to demonstrate the variation in the valence states of the arsenic, lead, and antimony. The Pb(IV) and Sb(V) content was found to increase substantially with the addition of NaNO3.

Selective Separation of Arsenic from Lead Smelter Flue Dust by Alkaline Pressure Oxidative Leaching

This study investigated the feasibility of using an alkaline pressure oxidative leaching process to treat lead smelter flue dust containing extremely high levels of arsenic with the aim of achieving the selective separation of arsenic. The effects of different parameters including NaOH concentration, oxygen partial pressure, liquid-to-solid ratio, temperature, and time for the extraction of arsenic were investigated based on thermodynamic calculation. The results indicated that the leaching efficiency of arsenic reached 95.6% under the optimized leaching conditions: 80 g/L of NaOH concentration, 1.0 MPa of oxygen partial pressure, 8 mL/g of liquid-to-solid ratio, 120 °C of temperature, 2.0 h of time. Meanwhile, the leaching efficiencies of antimony, cadmium, indium and lead were less than 4.0%, basically achieving the selective separation of arsenic from lead smelter flue dust. More than 99.0% of arsenic was converted into calcium arsenate product and thus separated from the leach solution by a causticization process with CaO after other metal impurities were removed from the solution with the addition of Na2S. The optimized causticization conditions were established as: 4.0 of the mole ratio of calcium to arsenic, temperature of 80 °C, reaction time of 2.0 h. The resulting product of calcium arsenate may be used for producing metallic arsenic.

Arsenic and antimony extraction from high arsenic smelter ash with alkaline pressure oxidative leaching followed by Na2S leaching

In this study, a combined process of alkaline pressure oxidative leaching (APOL) and Na2S leaching was proposed to extract arsenic and antimony from high arsenic smelter ash. The arsenic was selectively dissolved into the leachate during APOL, while the antimony was separated into the residue as insoluble NaSb(OH)6. The effects of different process parameters on the extraction of arsenic and antimony were investigated based on thermodynamic calculation. The results indicated that the extraction of arsenic was as high as 96% after APOL process under the conditions: 2 mol/L of NaOH concentration, 95 °C of leaching temperature, 90 min of leaching time, 1.5 MPa of oxygen partial pressure, 10:1 of liquid-to-solid ratio. The NaSb(OH)6 generated after APOL could be converted to Na3SbS4·10H2O and thus was leached out by Na2S leaching. The optimum conditions were established as: 2.75 mol/L of Na2S concentration, 95 °C of leaching temperature, 120 min of leaching time and 10:1 of liquid-to-solid ratio, under which the extraction of antimony was as high as 97.01%. A series of tests indicated that the leaching toxicity of the leaching residue met the Chinese identification standards.