Leach Residue Research Articles

Cyclic performance of coke oven gas - Steam reforming with assistance of steel slag derivates for high purity hydrogen production

Steel slag-derived CaO (CaO(SS)) was applied in cyclic sorption enhanced steam reforming of coke oven gas process (cyclic SE-SRCOG) for high-purity hydrogen production. Effects of leaching acid concentration on the cyclic CO2 capture capacity of CaO(SS) were investigated. Using 1 mol/L acetic acid solution as leaching solution, the prepared CaO-based sorbent showed a stable CO2 capture capacity of 0.27 gCO2/gCaO after 35 sorption-desorption cycles. Impacts of operating modes (with different particles contents and loading patterns), operating procedure, and leaching residue (LR) blending on cyclic SE-SRCOG performance were experimentally investigated in detail. Combined catalyst-sorbent by adhesively tableting exhibited the worst stability due to loss of Ni active sites, caused by severe sintering and coverage by CaCO3. With a reforming temperature of 600 °C, cyclic SE-SRCOG with layered packing catalyst-sorbent particles could achieve H2 purity of 95% and CH4 conversion of 87% at the 5th cycle, and they were still maintained to be 87% and 65% at the 10th cycle, respectively. In operating mode using layered packed catalyst-sorbent particles, integration of LR into the sorbent facilitated further improvements of CH4 conversion to 86% at the 10th cycle, mainly due to that LR could act as catalytic promoters of SRCOG and sintering-resistance spacers of CaO(SS).

Reductive transformation and synergistic action mechanism in the process of treating spent lithium-ion batteries with antibiotic bacteria residues

The rapid development of sustainable energy has led to the production of a large number of spent lithium-ion batteries (LIBs), and the toxic metals in spent LIBs will have an adverse impact on the environment and human health. An environmentally friendly and economical recycling method is essential for resource recycling and environmental management. In this work, “treating waste with waste” is innovatively proposed to realize the high-value utilization of different types of solid waste based on the reduction conversion between antibiotic bacteria residues (ABRs) and spent cathode powder (SCP). The reaction products are primarily transformed into Ni, Co, C, MnO, Li2CO3 and CaO under the optimum conditions of ABRs thermal reduction. Then, the valuable metals can be efficiently recovered by low acid leaching without reducing agent. The recovery efficiencies of Ni, Co, Mn, and Li are 99.4%, 99.5%, 99.9% and 99.9% respectively under the optimum leaching condition of 1.0 mol L−1 H2SO4, L/S ratio of 20:1, reaction temperature of 333 K and reaction time of 60 min. In addition, the acid leaching residue can be directly reutilized as LIBs anode materials, which exhibit excellent electrochemical performance. Comparing with the traditional thermal reduction method, this method has advantages of realizing synergistic treatment of two kinds of wastes simultaneously through one-step heat treatment, efficient recovery of valuable metals with low acid and high value utilization of ABRs. The process of treating waste with waste provides a new direction for the recovery of spent LIBs and the treatment of carbonaceous solid wastes.

Process integration of sulphuric acid leach and particle size fractionation for cleaner copper concentrate production by enhancing the radionuclide mobility

• A process integration of selective H 2 SO 4 leach and particle size fractionation. • Mobility of long-life radionuclides ( 238 U, 230 Th, 226 Ra, 210 Pb, 210 Po) in process. • The underlying mechanism of radionuclide mobility is unveiled. • A conceptual process is proposed for the cleaner Cu production from IOCGU ores. The growing demand of copper metal in the world continues to push the limit of utilization of complex copper resources, among which the iron oxide associated with copper–gold-uranium ore deposits (IOCG-U) contain economic copper and abundant gangues including uranium-bearing minerals that carry naturally occurring radionuclides and pose a challenge for public health and environment. To support the removal of radionuclides from copper flotation concentrate, the particle size fractionation method is integrated into the H 2 SO 4 leach process to illustrate the deportment of long-life radionuclides ( 238 U, 230 Th, 226 Ra, 210 Pb, 210 Po) and other non-copper elements during the process. H 2 SO 4 leach treatment largely dissolved non-copper elements (Fe, Ca, Mg, Al, K, F, U, Th, etc) from the gangue minerals and promoting the further copper phase transformation from the primary chalcopyrite and bornite towards covellite and copper sulphates. In the leach process, over 98% of 238 U and 230 Th were dissolved into H 2 SO 4 leach solution due to the decomposition of uranium-bearing minerals, while the released 226 Ra, 210 Pb and 210 Po from uranium-bearing minerals were largely remobilized, precipitated as insoluble sulphate salts, and accumulated in the small particle size fractions of the H 2 SO 4 leach residue. The underlying mechanism of this deportment is attributed to the mineralogical changes as an increased amount of covellite, digenite and copper sulphate are associated with the small particle size fractions, and serve as preferential hosts for 226 Ra, 210 Pb and 210 Po. Therefore, a conceptual flowsheet which facilitates further removal of 226 Ra, 210 Pb and 210 Po from the H 2 SO 4 leach residue is discussed based on particle size fractionation of the copper concentrate leach residue where radionuclides have been selectively leached or re-distributed.

Selective separation of Sc(III) and Zr(IV) from the leaching of bauxite residue using trialkylphosphine acids, tertiary amine, tri-butyl phosphate and their mixtures

• It is proposed a scandium and zirconium separation from bauxite residue. • All zirconium was separated using Alamine 336 10% in kerosene. • D2EHPA+TBP mixture reached same separation factor for scandium than Cyanex 923. • Cyanex 923 scrubbing may be carried out with HCl with negligible scandium losses. • The process is strictly connected to the SDGs 7, 8, 9, and 12. The separation of scandium from the solution generated in acid leaching is one of the main challenges for the hydrometallurgical processing of bauxite residue. Several organic extractants were evaluated, being the most prominent phosphine acids, while amine-based compounds have shown great results for separating metallic ions. The present study aims at the separation of scandium and zirconium from the Brazilian bauxite residue. The acid leaching was carried out using 20% of H 2 SO 4 , S/L ratio equals 1/10, for 8 h at 90 °C. Further, solvent extraction experiments were carried out using D2EHPA, Cyanex 923, and Alamine 336 diluted in kerosene, evaluating the effect of pH, temperature, extractant concentration, synergism with TBP and A/O ratio. After optimizing various process parameters, Alamine 336 has obtained separation factors of Zr/Al, Zr/Fe, and Zr/Ti equal to 15150, 45054, and 19713, respectively, at pH 1.0, A/O ratio equals to 1:1 and 25 °C. The stripping rate achieved 92% using Na 2 CO 3 0.25 mol/L. The scandium separation ratio reached higher values for 10% of Cyanex Cyanex 923 than D2EHPA 10% + TBP 5%. Scrubbing for contaminants removal may be carried out using HCl 5 mol/L with 0.1% of scandium losses, and all scandium was stripped by H 3 PO 4 5 mol/L. Despite the works reported in the literature, there a few of them exploring the recovery of scandium and zirconium by the leaching-solvent extraction process. The flowchart proposed is strictly connected to the sustainable development goals 7, 8, 9, and 12.

Integrated carbothermic smelting – Acid baking – Water leaching process for extraction of scandium, aluminum, and iron from bauxite residue

A closed-loop valorization process, combining carbothermic smelting and acid baking – water leaching was developed to recover scandium, aluminum, and iron from bauxite residue, a major side product of aluminum production. This process employs carbothermic smelting to recover 99% of the iron in the starting residue as crude metallic iron, which can be separated from slag that contains aluminum and scandium. Composition of slag was found to be primarily dependent on the smelting temperature and amount of smelting flux added. At high smelting temperatures (1600 °C), sodium was found to evaporate from slag, potentially enabling sodium separation and recovery. The slag was treated by acid baking – water leaching, to convert scandium and aluminum to soluble sulfate species, allowing recovery. The single-pass acid baking – water leaching was optimized to achieve 99% scandium and 47% aluminum extraction by employing sulfuric acid at a ratio of 1.25 gH₂SO₄/g, baking at 150 °C for 2 h, and water leaching at ratio of 7.5 mLH₂O/g for 210 min. The production of solid leaching residue was avoided by employing thermal desulfation at 1400 °C to remove 84% of leaching residue sulfur content thereby potentially allowing regeneration of sulfuric acid for acid baking and producing a recycled smelting flux that would reduce reagent consumption and would allow recapturing the scandium and aluminum not recovered during leaching. This closed-loop process enables efficient and sustainable recovery of scandium, iron, and aluminum from bauxite residue, and constitutes first steps in an integrated process for sustainable valorization of bauxite residue, supporting efforts to develop a zero-waste economy.

Influence of alkaline pre-treatment on acid dissolution of cathode material of 18650 lithium battery

Lithium battery cathodes contain lithium, cobalt, nickel, and manganese. Recycling of spent lithium batteries aims to recover these elements for re-use. Liberation of cathode materials from other metals in the battery such as aluminium, copper, and iron, is essential to obtain a good leaching efficiency in the recovery of valuable metals from end-of-life lithium batteries. This study investigates the behaviour of cathode materials and other metals in spent 18650 lithium batteries during leaching in H2SO4 solution with and without NaOH pre-treatment. Dissolution of aluminium using NaOH is a selective method to separate the metal from other elements. The influence of a 2-hour NaOH pre-treatment on subsequent acid leaching of cathode materials was investigated at both room temperature and 80°C. The extraction of aluminium increased to 75% at a higher temperature. Lithium concentration in the pregnant leach solution from acid leaching also increases with NaOH pre-treatment. The pre-treatment had a negligible effect on nickel, manganese, iron, and copper extraction. However, the cobalt extraction with NaOH pre-treatment was significantly lower. The result was likely due to indirect impact of less hydrogen gas was generated from a lower Al amount. The lattice structure of the leach residue for the sample with NaOH pre-treatment was monoclinic rather than rhombohedral due to stronger delithiation.

Addition of Fe3O4 as electron mediator for enhanced cementation of Cd2+ and Zn2+ on aluminum powder from sulfate solutions and magnetic separation to concentrate cemented metals from cementation products

The zinc leach residues (ZLRs) generated by Zn hydrometallurgical process are disposed of in landfills due to a lack of proper treatment methods although they contain toxic heavy metals like Cd and Zn. To remove these heavy metals, this study investigated cementation using aluminum (Al) together with magnetite (Fe3O4) from sulfate solutions. Cementation using only Al could not remove Cd2+ and Zn2+ because the insulating Al oxide (Al2O3) layer on Al surface was not removed at high pH (5.0) from sulfate solution, and inhibiting electron transfer from Al to metal ions. The addition of Fe3O4, however, enhanced the cementation of Cd2+ and Zn2+ from 0% to 83% and 92%, respectively. The results of SEM-EDS analysis for the cementation products confirmed that Cd and Zn were deposited on Fe3O4 attached on Al surface. Electrochemical impedance spectra for Al electrode in sulfate solutions confirmed that charge transfer resistance decreased when Fe3O4 powder was attached on the electrode surface, suggesting that Fe3O4 acts as a mediator to enhance electron transfer from Al to metal ions in solutions. By applying magnetic separation for the cementation products, the majority of cemented Cd and Zn were recovered together with Fe3O4 in magnetic fraction, and they were separated from Al in non-magnetic fraction.

Leaching of Zinc for Subsequent Recovery by Hydrometallurgical Techniques from Electric Arc Furnace Dusts and Utilisation of the Leaching Process Residues for Ceramic Materials for Construction Purposes

Steel is one of the most widely used materials in the past and today. Various techniques are used to recycle this material, including the electric arc furnace. This process has several advantages, but it also has a major disadvantage, namely, the generation of waste such as electric arc furnace dusts. Electric arc furnace dusts are classified as hazardous waste due to their high percentage of heavy metals, including zinc. Consequently, in the present research, the leaching of zinc for recovery with sulfuric acid solutions at ambient temperature and atmospheric pressure is evaluated, as well as the reuse of the leaching process residue as a raw material for ceramic materials. The sulfuric acid solutions were 0.125, 0.25, 0.5, and 1 molar, using clay for ceramic conforming and percentages of the leaching residue from 0–50%. The results showed that the optimum solution was 1 molar sulfuric acid, recovering all the zinc in the sample in 36 h. Furthermore, it was found that the clay-conformed ceramics with less than 40% leaching residue showed acceptable physical and mechanical properties according to standards. Therefore, this research develops a new environmental hydrometallurgy in which metallic elements of interest are valorized and the production of waste is avoided, reducing the deposition of hazardous waste in landfills and the extraction of raw materials for the manufacture of construction materials.

An integrated process for separation and recovery of tantalum from spent TaCrTi sputtering targets by hydrometallurgical methods

In order to recover valued metals from spent TaCrTi sputtering targets, an integrated hydrometallurgical method including leaching, solvent extraction and precipitation was proposed. Most of impurities were firstly removed by H2SO4, Ta and Ti then enriched in the leaching residues were dissolved by H2SO4 mixed with HF. 99% of Ta dissolved in the solution was separated by extraction with tri-butyl-phosphate (TBP) and stripped by 3.9 M NH4F solution. Ta2O5 powders were obtained in the Ta stripping solution by adding aqueous ammonia following calcination in the tube furnace at 900 °C. K2TiF6 powders was prepared by adding the KCl in the Ta free raffinate. An integrated process for recovery Ta and Ti from the spent TaCrTi sputtering targets was proposed.

Effect of polyvinyl pyrrolidone on chalcopyrite bioleaching with Acidithiobacillus ferrooxidans

Chalcopyrite (CuFeS2) is the most abundant copper-bearing mineral in nature. Large amounts of low-grade chalcopyrite ores cannot easily be utilized due to technical and economic limitations, which lead to an enormous waste of resources. Bioleaching is one of the most common methods to treat low-grade copper ores. However, the low leaching efficiency limits the industrial application of bioleaching chalcopyrite. In this study, polyvinyl pyrrolidone (PVP) was used to intensify the bioleaching of chalcopyrite with Acidithiobacillus ferrooxidans. The bioleaching results showed that PVP increased the bioleaching efficiency and the maximal effect was achieved by adding PVP after 6 days of experimental time. The solid leach residues were investigated using X-ray diffraction (XRD), Raman spectroscopy, and X-ray photoelectron spectrometry (XPS) analyses to elucidate the role of PVP. Elemental sulfur and jarosite were detected on the chalcopyrite surface. These compounds may have formed a “passivation” layer, thereby hindering bioleaching as described in the literature. Moreover, PVP inhibited the generation of elemental sulfur and weakened the stability of the “passivation” layer. Additionally, PVP may have improved the surface hydrophilicity of chalcopyrite, which probably lead to an enhanced solid-liquid mass transfer.

Preparation of Zinc Oxide and Zinc Ferrite from Zinc Hypoxide by Wet Process and Electrochemistry

With the increase of zinc resource consumption, the recovery and utilization of zinc resources in zinc suboxide has become one of the current research hotspots. In this study, the electrochemical method was used to remove the impurities in the zinc leaching night and enrich the zinc ferrite in the ammonia leaching residue for the solution and ammonia leaching slag after the ammonia leaching of zinc hypoxide, in order to realize the comprehensive utilization of the essence of zinc immersion night and new resources. The results showed that the reduction potentials of copper, lead, cadmium, and zinc in the ammonia leaching solution were analyzed by electrochemical testing methods to be −0.76 V, −0.82 V, −0.94 V, and −1.3 V, respectively. Through constant potential electrodeposition, the removal rate of copper, lead, cadmium. The removal rate of cadmium is 98.73%, and the removal rate of lead and copper is more than 99%. The purified ammonia leaching solution is evaporated at 90 °C for 25 min to obtain basic zinc carbonate. The purity of ZnO obtained after calcination at 500 °C for 120 min is 96.31%. The ammonia leaching residue was pickled with 3 mol·L−1 acetic acid for 30 min to effectively remove PbCO3, and then magnetic separation was carried out with a current intensity of 1.4 A. The final zinc ferrite content was 83.83%.

Extraction of copper from complex carbonaceous sulfide ore by direct high-pressure leaching

The increase of impurities and complexity of copper ores are among the recent challenges in the mining industry. Complex carbonaceous sulfide ores are extremely difficult to treat due to their mineralogical complexity and impurities of organic carbon and carbonates. This study focuses on the development of a hydrometallurgical process for efficient copper extraction from complex carbonaceous sulfide ore which contains chalcopyrite, carbonates (dolomite and calcite), and carbonaceous gangue minerals. Characterization of the ore sample and leach residues was conducted using XRD and EPMA analysis, while ICP-OES was used for the determination of total dissolved metals in solution. High-pressure leaching of complex carbonaceous sulfide ore in oxygenated sulfuric acid solution was performed and the influence of leaching parameters such as sulfuric acid concentration, temperature, total pressure, and pulp density was studied. The extraction of copper increased with increasing temperature, sulfuric acid concentration, and total pressure. On the other hand, an increase in pulp density resulted in a decline in copper extraction due to an increased slurry viscosity and resistance in the diffusive mass transfer of reactants. Selective dissolution of copper from iron can be achieved by controlling free acidity in the pregnant leach solution (PLS). Under these leaching conditions: 100 g/L, 1 M H2SO4, 160 °C, 1.0 MPa total pressure, the highest copper and iron extractions achieved were 97.55% and 95.37%, respectively. Precipitation of copper from the PLS by NaSH sulfidization was investigated and more than 99.9% of copper was recovered at a Cu: NaSH molar ratio of 1:1.8.

Multi-step purification of electrolytic manganese residue leachate using hydroxide sedimentation, struvite precipitation, chlorination and coagulation: Advanced removal of manganese, ammonium, and phosphate

Water pollution caused by the release of manganese (Mn2+) and ammonia nitrogen (NH4+-N) from electrolytic manganese residue (EMR) generated from industrial activities poses a serious threat to ecosystems and human health. In this study, an integrated process consisting sequentially of hydroxide sedimentation, struvite precipitation, breakpoint chlorination, and ferric chloride coagulation was optimized to remove Mn2+ and NH4+-N from EMR leachate, and to address the issue of residual orthophosphate caused by struvite precipitation. The precipitates were characterized using X-ray diffraction, scanning electron microscopy, and thermogravimetric analyses. Results show that Mn2+ ions and the resulting chemical oxygen demand (COD) were mainly removed using hydroxide precipitation at a sedimentation pH of 10.2, with poor-crystalline manganese hydroxide as the main precipitate. NH4+-N was primarily removed and recovered using struvite precipitation with well crystalline struvite as the main product, and then further eliminated using breakpoint chlorination. The residual orthophosphate introduced by struvite precipitation is successfully removed with ferric coagulation, and the effluent pH (7.5) is also lowered to discharge limits by means of hydrolysis of ferric coagulant. The concentration of COD, Mn2+, NH4+-N, and orthophosphate concentrations in the final effluent were 30.52 ± 9.38, 0.026 ± 0.013, 0.87 ± 0.01, and 0.06 ± 0.002 mg/L, respectively, meeting all local discharge standards. This combined process has robust pollutant removal efficiency, high resource recovery potential and few environmental constraints; thus, it is recommended as a potential solution for the treatment of Mn2+- and NH4+-N-rich acid mine drainage.

Sequential-Anaerobic and Sequential-Aerobic Bioleaching of Metals (Ni, Mo, Al and V) from Spent Petroleum Catalyst in Stirred Tank Batch Reactor: A Comparative Study.

The online version contains supplementary material available at 10.1007/s12088-021-00978-8.

Extraction of gold from typical Carlin gold concentrate by pressure oxidation pretreatment - Sodium jarosite decomposition and polysulfide leaching

Most of the gold is encapsulated in sulfides and arsenopyrite in Carlin type gold concentrate; the gold leaching efficiency is less than 8% without pretreatment. Cyanide process is a traditional gold leaching method with the disadvantages of several environmental pollution incidents occurred due the breach of tailings disposal. In view of the difficulties in gold extraction and serious pollution incidents of typical Carlin type gold concentrate, this paper proposes pressurized oxidation - sodium jarosite decomposition - polysulfide leaching method to recover gold from typical Carlin type gold concentrate. Standard characterization/analytical methods of XRD, ICP, MLA, XPS, and SEM-EDS were employed in the experiment. In the pressurized oxidation pretreatment process with high temperature and ferric sulfate concentration, jarosite minerals with stable properties were easily formed due to the appropriate amounts of K+ introduced in the flotation of the ore and a small amount of mica minerals containing Na+. The results show that pressurized oxidation pretreatment can efficiently liberate gold from the mineral, but causes secondary inclusion of sodium jarosite. The one-step sodium jarosite decomposition treatment of the residue produced by oxygen pressure leach, can effectively liberate gold from the sodium jarosite pHase. Thus, the influence of factors on extracting gold from sodium jarosite decomposition residue using polysulfide were systematically studied. Under the most suitable conditions, the leaching efficiency of gold increased up to 90.2%, which was 8.4% higher than that of the direct cyanidation of oxygen pressure leach residue. This study provides an efficient, clean, and environmentally friendly way for the resource utilization of refractory gold concentrate, with a potential industrial application.

Recovery of Al, Co, Cu, Fe, Mn, and Ni from Spent LIBs after Li Selective Separation by the COOL‐Process. Part 1: Leaching of Solid Residue from COOL‐Process

AbstractLithium recycling from spent LIBs along the COOL‐process produces a Li‐free metal rich black mass, which still contains the entire fraction of valuable metal such as Co, Ni, and Mn. A recycling process was developed, which allows for mobilizing these metals. The first process stage is the selective leaching of Li via COOL‐process. Subsequently, two inorganic (H2SO4 and HCl) and two organics acids (citric and gluconic acid) were tested for mobilizing the metals from the solid residue. To optimize this process stage, acid concentration as well as addition of H2O2 as a reduction reagent were investigated. The optimal conditions to dissolve valuable metals such as Co, Cu, Fe, Mn, and Ni was identified as 2 N H2SO4 and 2 vol % H2O2.

Leaching Basalt Rock using Averrhoa Bilimbi Extract

<p>This research was conducted to determine the effect of solid/liquid ratio (s/v), temperature, and stirring Speed on per cent extraction of iron using oxalic acid from Averrhoa bilimbi. The variation of the solid/liquid (s/v) is 10%, 15% and 20 (%), the temperature variations used are 30, 60 and 90 <sup>o</sup>C, while the variation of the stirring Speed uses 200, 400 and 600 rpm. Some basalt rock samples are put into the reactor and leached using an oxalic acid solution from Averrhoa bilimbi. Within a certain period of time, the filtrate was taken and analyzed for Fe content using Atomic Absorption Spectroscopy. The residue was analyzed using XRF and XRD. The results showed that the highest percentage of iron extraction for basalt rock leaching experiments using oxalic acid from Averrhoa bilimbi was achieved at a temperature of 90<sup>o</sup>C, 10% of solid/liquid ratio, 400 rpm stirring speed and 120 minutes of leaching time with a percent iron extraction of 65.90%. XRD analysis results showed that the compounds formed in the residual leaching were Anorthite, Ferrosilite, Cristobalite and quart low.</p>

Pretreatment to Leaching for a Primary Copper Sulphide Ore in Chloride Media

The dissolution of copper sulphide ores continues to be a challenge for the copper industry. Several media and leaching alternatives have been proposed to improve the dissolution of these minerals, especially for the leaching of chalcopyrite. Among the alternatives, pretreatment prior to leaching was proposed as an option that increases the dissolution of copper from sulphide ores. In this study, a mineral sample from a copper mining company was used. The copper grade of the sample was 0.79%, and its main contributor was chalcopyrite (84%). The effect of curing time (as pretreatment) in a chloride media on copper sulphide ore was evaluated at various temperatures: 25, 50, 70 and 90 °C. The pretreated sample and leaching residues were characterized by X-ray diffraction, scanning electron microscopy, and reflected light microscopy. Pretreatment products such as CuSO4, NaFe3(SO4)2(OH)6, and S0 were identified although with difficulty, due to the low presence of chalcopyrite in the initial sample (1.99%). Under the conditions of 15 kg/t of H2SO4, 25 kg/t of NaCl, and 15 days of curing time, a copper extraction of 93.1% was obtained at 90 °C with 50 g/L of Cl− and 0.2 M of H2SO4.

An integrated and sustainable hydrometallurgical process for enrichment of precious metals and selective separation of copper, zinc, and lead from a roasted sand

The developed hydrometallurgical process for recovering copper, zinc, and lead from the roasted sand of WPCB-SA. • A hydrometallurgical process for recovering valuable elements was proposed. • Copper and zinc were selectively recovered by water leaching. • Lead was efficiently recovered by high-chloride leaching. • The leaching kinetics and conversion mechanism of lead were analyzed. • Gold and silver have been successfully enriched in the leached residue. This study developed an efficient and sustainable hydrometallurgical process for the enrichment of gold and silver and the stepwise separation of copper, zinc, and lead from sulfated roasted sand of waste printed circuit board smelting ash. Selective separation of copper and zinc was achieved by water leaching, and silver dispersion was reduced by controlling the amount of NaCl added during the leaching process. The results of the water leaching showed that the copper and zinc leaching rates were 99.85% and 99.47%, respectively, whereas the loss rate of silver was 2.1% with optimal leaching parameters. The high-chloride-complex method was used to study the efficient conversion and separation of lead from the leached residue, and the leaching kinetics and conversion mechanism of lead were discussed. The results showed that under the optimal conditions, the leaching rate of lead was 99.79%. Leaching kinetics analysis showed that lead leaching in the high - chlorine system was controlled by a chemical reaction; the apparent activation energy was 53.63 kJ/mol. After the leaching of copper, zinc, and lead, 1.66% Ag and 213 g/t Au were enriched in the leached residue; and the precious metal enrichment goal was reached. The chlorinated leachate showed good recycling performance, and a lead leaching rate of 97.93% was obtained after three circulations. After cooling, crystallization, and purification, lead chloride with a purity of 99.89% and high economic and industrial value was obtained from the lead-rich leachate. This process has favorable and sustainable industrial application prospects.

Leaching kinetics of low nickel matte in an aqueous solution of sulfuric acid saturated with FeSO4 and NiSO4

In the previous study of our group, the low nickel matte was leached directly using H2SO4 under atmospheric pressure. As long as the acidity of the solution was high enough, the high-efficiency leaching of iron, cobalt and nickel can be realized. Moreover, the cuprous sulfide and precious metals cannot be decomposed by H2SO4 and were enriched in the leach residue. Due to the high concentration acid used in the leaching process of low nickel matte, the residual acid in leaching solution may need to be neutralized firstly before the subsequent metal separation. However, a lot of alkali will be consumed, resulting in the formation of harmful inorganic salts. In the experiments, we found that sulfate in the leaching solution could reach saturated and continuously crystallized out under the condition of high concentration of sulfuric acid. Based on the above considerations, the crystallization method may be a choice. Upon combination of the leaching and crystallization process, a new way to extract valuable metals and effectively recycle the sulfuric acid was proposed. However, the dissolution of low nickel matte will be also affected when the solution is saturated. Therefore, it is necessary to explore the leaching behavior of low nickel matte in the presence of solutions saturated with sulfate. In this study, the leaching kinetics of low nickel matte in H2SO4 solutions saturated with FeSO4 and NiSO4 was investigated. Kinetic data of the dissolution of nickel for various parameters showed best fit to the kinetic model governed by diffusion in the inert porous media. The results shown that the reaction order of H2SO4 was 3.02 and the activation energy of the system was found to be approximately 48.7 kJ·mol−1. Compared with sulfuric acid leaching, the reaction order of this study increased greatly, which indicated that the influence of acid concentration on leaching process increased. In addition, the decrease of activation energy showed that the oversaturated sulfate solution leaching process was consistent with the internal diffusion control, which may be because the existence of crystal had an effect on the diffusion of H+ ions in pores.