Sulfuric Acid Leach Liquor Research Articles

Development of a conceptual direct solvent extraction (DSX) route and a flowsheet to produce purified concentrated cobalt and nickel solutions representing sulfuric acid leach liquor of laterite

Direct solvent extraction (DSX) was applied to produce purified concentrated Ni and Co solutions from a synthetic sulfuric liquor containing Ca, Cu, Mg, Mn, and Zn as impurities, which simulates the solution obtained by the sulfation-roasting-leaching process after precipitation of Fe, Al, and Cr. The commercial extractants Versatic 10, Cyanex 272, D2EHPA, TBP were used in 3 solvent extraction circuits. In the first circuit, operated at 40 °C, Versatic 10 (0.5 M) fully extracted Cu, Zn, Ni, and Co in 3 stages at O/A = 2:1, and pH 6.5, leaving remaining Mn (54% was extracted) and most of the Ca and Mg in the raffinate. The co-extracted Ca and Mg were fully scrubbed off the loaded Versatic 10 in 2 stages at O/A = 5:1, and pH 6.5. The other metals were stripped out of the Versatic 10 extract using a synthetic Ni spent electrolyte (60 g/L Ni, 2 M H2SO4) in 2 stages at O/A = 9:1. This loaded strip liquor was subjected to a second circuit with Cyanex 272 (0.64 M) operated at 50 °C. Three stages were required to fully extract Co, Cu, Mn, and Zn (O/A = 2:1, pH 4), whereas the raffinate containing 82 g/L Ni was deemed suitable for electrowinning. The co-extracted Ni(II) was fully scrubbed off the loaded Cyanex 272 in only 1 stage at O/A = 5:1, and pH 4. All Co, Cu, Mn, and Zn were stripped out from the scrubbed loaded Cyanex 272 in 2 stages at O/A = 10:1 using a synthetic Co spent electrolyte (45 g/L Co, 1 M H2SO4). The raffinate containing 58.4 g/L Co was submitted to a third circuit using a D2EHPA (0.6 M) + TBP (0.73 M) synergistic system operated at 50 °C. Zinc(II) was fully extracted by the D2EHPA + TBP system in 2 stages at pH 2 and O/A = 1:3, while Mn(II) and Cu(II) were fully extracted from the Zn-depleted raffinate in 2 stages at pH 3.5 and O/A = 2:1. The raffinate containing 58.3 g/L Co was deemed suitable for electrowinning. Copper(II), Mn(II), and Zn(II) were stripped out from the loaded D2EHPA + TBP in 3 stages at O/A = 2.5:1 using 1 M H2SO4. The real number of moles of extractants involved in the extractions and apparent equilibrium constants were estimated for all circuits. A flowsheet of the purification conceptual route is presented.

Separation and recovery of scandium from high pressure sulfuric acid leach liquor of limonitic laterite

Extraction of Sc from Sc-rich limonitic laterite without interfering with the main nickel-cobalt production circuit will significantly enhance the economic value of laterite. A process for separation and recovery of Sc from high pressure acid leach (HPAL) liquor of limonitic laterite was proposed in this paper. Based on the chemical composition of the leach liquor, precipitation thermodynamic analysis and experiments were conducted. The results indicated that after a first step to precipitate Fe3+ at pH = 2.5 and 90 °C, Sc could be enriched through a second step precipitate at pH = 4.8 and 60 °C to 267 g/t in Sc-containing precipitate. Then, Sc was re-leached from the Sc-containing precipitate via 1 mol/L sulfuric acid at L:S ratio of 5 mL/g, 80 °C, and 40 min. Meanwhile, gypsum in the Sc-containing precipitate transformed into calcium sulfate hemihydrate during leaching. After that, Sc was synergistically extracted from the leach solution via 5% P204 and 2.5% TBP. Moreover, impurities extracted into the organic phase during the extraction of Sc were scrubbed by HCl. Finally, Sc was stripped by NaOH. By using this process, Sc could be efficiently recovered from limonitic laterite in a facile way.

Separation and Recovery of Cobalt and Nickel from Sulfuric Acid Leach Liquor of Zinc Smelting Residue with Acidic Extractants

Separation and Recovery of Cobalt and Nickel from Sulfuric Acid Leach Liquor of Zinc Smelting Residue with Acidic Extractants

A novel synergistic extraction system for the recovery of scandium (III) from sulfuric acid medium with mixed Cyanex923 and N1923

High extraction capability and outstanding selectivity are very attractive in the solvent extraction process. Scandium (Sc(III)), as a primary rare earth element (REE), generally lacks independent large-scale deposits and is mainly recovered as a by-product during the processing of mineral. However, the difficulty of stripping has been always a crucial problem to be solved in the scandium extraction system. This study aims to improve stripping performance while enhancing extraction efficiency for scandium. Herein, a novel synergistic extraction system for the recovery of scandium with a mixture of Cyanex923 (A, a mixture of four trialkylphosphine oxides) and N1923 (B, primary amine R1R2CHNH2) in sulfuric acid medium was put forward. The effect of the molar fraction (X) of Cyanex923 on the extraction of scandium was systematically investigated. The experimental results demonstrated that the coupled Cyanex923-N1923 exhibited an obvious synergistic effect on Sc(III). The maximum synergistic enhancement coefficient (Rmax) was calculated to be 2.20 when the molar fraction (X) of Cyanex923 was 0.4, and the H2SO4 concentration was 0.25 mol·L−1. Moreover, the synergistic extraction mechanism was discussed combined with multiple techniques including slope analysis method, FT-IR and NMR characterizations, to which the structure of extracted complexed was deduced to be Sc(SO4)1.5∙(A)0.5(B). Thermodynamic experiment was also explored in detail. More importantly, stripping with only 2 mol·L−1 HNO3 produced a relatively high stripping rate of 95%. The synergistic system still presented excellent reusability even after five treatment cycles. Finally, this study pointed out a potential industrial application value for the separation and recovery of Sc(III) from the sulfuric acid leaching liquor of nickel laterite ores.

Removal of Fe(III) from sulfuric acid leaching solution of phosphate ores with bisphosphonic acids

Iron(III) in phosphate ore leach liquor harms the quality of phosphoric acid and phosphogypsum crystallization. In this paper, the selective separation of iron(III) was studied with two novel extractants, N, N-n-octylamine di(methylene phenylphosphonic acid) (OADMPPA) and N, N-n-hexadetamine di(methylene phenylphosphonic acid) (HADMPPA), from a mixed media of sulfuric acid and phosphoric acid. The effects of temperature, time, diluents, and the concentrations of sulfuric acid and extractant on the separation of iron were investigated. The mechanism of extraction was established by UV–Vis spectroscopy, FT-IR spectroscopy, slope analysis, and the equimolar series method. The results showed that the two new reagents have good extraction ability toward iron(III) even at higher sulfuric acid concentration. Both OADMPPA and HADMPPA can remove more than 99% of iron(III) from the sulfuric acid leach liquor of phosphate ore. The stripping of iron(III) from the organic phase after extraction with acidic phosphine (phosphorus) extractants was difficult. Oxalic acid was found to be a good stripping agent. The single stage stripping efficiency reached 62.8%, and the two-stage stripping efficiency could reach over 99%. The separation and recovery of iron(III) in phosphate ore leach liquor could be realized by our method without changing the traditional phosphoric acid production process, which is beneficial to the purification of phosphoric acid and the recovery of iron.

Aluminum–iron separation in high‐acid leaching solution and high‐purity alumina preparation

AbstractThe demand for bauxite has increased sharply due to the fast development of alumina industry, so new technology must be developed to take full advantage of nontraditional bauxite resources to realize a continuous supply of bauxite. In this experiment, anatase sulfuric acid leaching liquor is taken as the research object to conduct aluminum–iron separation and high‐purity alumina preparation. The leaching solution contains 34.20 g/L of Al2O3, 25.30 g/L of TFe, 7.96 g/L of TiO2, 12.74 mg/L of Sc2O3, and 288.72 g/L of H2SO4. When the temperature is low, aluminum and iron condensation, which are not conducive to the separation and recovery of various elements, may occur easily. In this study, the ammonium aluminum sulfate crystallization–recrystallization and ion exchange–calcination method is adopted to effectively separate aluminum and iron in a strong acid solution comprising high concentrations of aluminum and iron without consuming high residual acid in the leaching liquor. Consequently, successful production of high‐added‐value product 4N (99.99%) level γ‐Al2O3 with ammonium aluminum sulfate is obtained through separation as the raw material. The product has a specific surface area of 7.845 m2/g with a nearly spherical shape and being evenly distributed. Therefore, it realizes high‐value aluminum utilization from nontraditional bauxite.

Interconnected α-Fe2O3 nanoparticles prepared from leaching liquor of tin ore tailings as anode materials for lithium-ion batteries

Interconnected α-Fe2O3 nanoparticles were prepared by a facile chemical precipitation method with the sulfuric acid leaching liquor of tin ore tailings as Fe source. The effect of different precipitants (sodium carbonate, sodium hydroxide, and ammonia) on the microstructure and lithium storage properties of the as-prepared α-Fe2O3 was studied in detail. All the prepared samples are composed of interconnected α-Fe2O3 nanoparticles with particle sizes ranging from 30 to 60 nm. As anode material for lithium-ion batteries, the electrochemical performance of the α-Fe2O3 sample prepared using sodium carbonate as precipitant (FE-1) exhibits the best electrochemical performance, followed by the α-Fe2O3 sample prepared using sodium hydroxide as precipitant (FE-2), and the α-Fe2O3 sample prepared using ammonia as the precipitant (FE-3). For example, the FE-1 electrode exhibits a reversible discharge capacity of 1146 mAh g−1 at 0.5 A g−1 after 300 cycles, higher than the corresponding values of 969 and 834 mAh g−1 for FE-2 and FE-3, respectively. At a high current density of 4 A g−1, the FE-1 electrode can still deliver a reversible discharge capacity of 377 mAh g−1, higher than that of FE-2 (254 mAh g−1) and FE-3 (86 mAh g−1) electrodes. Electrochemical impedance spectroscopy (EIS) analysis reveals that the FE-1 electrode has the smallest electrochemical reaction resistance. Galvanostatic intermittent titration technique (GITT) measurements demonstrate that the Li+ diffusion coefficients of the electrode range from 10−14 to 10−11 cm2 s−1. Cyclic voltammetry (CV) analysis proves that pseudocapacitive behavior significantly contributes to the lithium ion storage in the three α-Fe2O3 sample samples. The results reported in this work provide clues for the high-value use of tin ore tailings, which has good social and economic benefits for the sustainable development of industries.

A sustainable approach for advanced removal of iron from CFA sulfuric acid leach liquor by solvent extraction with P507

The recovery of valuable metals such as lithium (Li), aluminum (Al), gallium (Ga) from coal fly ash (CFA) has attracted considerable attention in recent years. As a main component in CFA, iron (Fe) has adverse effect on the selective recovery of the metals of interest. Removal of Fe is essential for the recovery of these metals. Herein, this work reported a sustainable approach for efficient and advanced removal of Fe(III) from sulfuric acid leach liquor of CFA using 2-ethylhexyl phosphonic acid mono-2-ethylhexyl ester (P507) as extractant. The results showed that more than 97.6% of Fe(III) can be effectively removed from the synthetic solution with a Fe/Al separation factor of 2218 under the optimum conditions. The UV–Vis, 1H NMR, IR spectroscopic results indicated that the extraction mechanism of Fe(III) with P507 was proved to be ion exchange. This technique was subsequently examined to remove Fe from the sulfuric acid leach liquor of CFA. The removal of Fe remained above 95.4% after 8 cycles of extraction-stripping process with H2O2 coupled with H2SO4 as stripping agent, implying the high efficiency and long-term reusability of P507. The study developed a sustainable way for advanced Fe removal from CFA sulfuric acid leach liquor, which could facilitate the recovery of valuable metals from CFA.

Recovery of Fe, Mn, Ni and Co in sulfuric acid leaching liquor of spent lithium ion batteries for synthesis of lithium ion-sieve and NixCoyMn1-x-y(OH)2

A comprehensive utilization process of Fe, Mn, Ni and Co in the acid leaching liquor of spent lithium ion batteries was put forward. Fe and part of Mn were extracted to synthesize Fe-doped lithium ion-sieve. Ni, Co and the rest of Mn were coprecipitated to prepare ternary lithium ion battery cathode material precursor NixCoyMn1-x-y(OH)2. After removing Cu2+ ions in acid leaching liquor by adding MnS, two-stage counter current extraction and one-stage stripping were used to extract all Fe3+ and 40.3% of Mn2+ ions in sulfuric acid leaching solution to obtain a MnSO4 stripping solution containing Fe3+ ions. Iron-doped lithium ion-sieve was synthesized by oxidation of the resulting MnSO4 solution with KMnO4 followed by roasting and acid leaching. Fe incorporates into the lattice of lithium ion-sieve and improves the structure stability. Fe-doped lithium ion-sieve shows Li adsorption capacity of 35.29 mg.g−1 and Mn dissolution loss of 1.66%. The Li+ adsorption in salt lake brine with the Fe-doped ion-sieve follows a pseudo-second-order kinetic model. NixCoyMn1-x-y(OH)2 is regular spherical particles with β-Ni(OH)2-like structure. This process provides a simple method for recovery and utilization of Mn, Fe, Ni and Co in spent lithium ion batteries.

Recovery of gallium from sulfuric acid leach liquor of coal fly ash by stepwise separation using P507 and Cyanex 272

For the ever-growing demand of gallium (Ga) resources in electronic applications, the recovery of Ga from the mining residues or waste has received considerable interest. The present paper is focused on the selective recovery of Ga from the sulfuric acid leach liquor of coal fly ash (CFA) by stepwise separation using acidic extractant. The results show that Ga(III) can be effectively separated by controlling the equilibrium pH values and meanwhile employing appropriate extractant based on the different ionic potential and electronic configuration of metal ions. At equilibrium pH ≤ 0.8, the impurities Ti(IV) and Fe(III) in the liquor can be preferably removed by a two-stage cross current extraction with 1 mol/L P507. Then Ga(III) can be selectively separated against Al(III) using 0.5 mol/L Cyanex 272 at equilibrium pH of 2.4–2.6. The separation factors of Fe/Ga, Fe/Al and Ga/Al can reach 145, 133 and 40, respectively. The extraction mechanism of Ga(III) as well as the Ga(III) complexes with Cyanex 272 were investigated using slope analysis and spectroscopic methods including FT-IR, NMR and ESI-MS, indicating that the extraction of Ga(III) with Cyanex 272 follows an ion exchange mechanism. The Ga(III) complexes with Cyanex 272 (H2L2) contains three coordination compounds, GaL3, Ga(HL)L3 and Ga(HL)3L3. Both the P-O and P = O groups of Cyanex 272 play the roles in the coordination reaction of Ga(III). The study provides a feasible way for the selective recovery of Ga from coal fly ash.

Sustainable Hydrometallurgical Recovery of Valuable Elements from Spent Nickel–Metal Hydride HEV Batteries

In the present study, the recovery of valuable metals from a Panasonic Prismatic Module 6.5 Ah NiMH 7.2 V plastic casing hybrid electric vehicle (HEV) battery has been investigated, processing the anode and cathode electrodes separately. The study focuses on the recovery of the most valuable compounds, i.e., nickel, cobalt and rare earth elements (REE). Most of the REE (La, Ce, Nd, Pr and Y) were found in the anode active material (33% by mass), whereas only a small amount of Y was found in the cathode material. The electrodes were leached in sulfuric acid and in hydrochloric acid, respectively, under different conditions. The results indicated that the dissolution kinetics of nickel could be slow as a result of slow dissolution kinetics of nickel oxide. At leaching in sulfuric acid, light rare earths were found to reprecipitate increasingly with increasing temperature and sulfuric acid concentration. Following the leaching, the separation of REE from the sulfuric acid leach liquor by precipitation as NaREE (SO4)2·H2O and from the hydrochloric acid leach solution as REE2(C2O4)3·xH2O were investigated. By adding sodium ions, the REE could be precipitated as NaREE (SO4)2·H2O with little loss of Co and Ni. By using a stoichiometric oxalic acid excess of 300%, the REE could be precipitated as oxalates while avoiding nickel and cobalt co-precipitation. By using nanofiltration it was possible to recover hydrochloric acid after leaching the anode material.

Recovery of manganese from sulfuric acid leaching liquor of spent lithium-ion batteries and synthesis of lithium ion-sieve

A novel process of recovery of Mn as lithium ion-sieve from H2SO4 leaching liquor of spent lithium-ion batteries was developed. After precipitation of Fe3+ and Cu2+ in the leaching liquor, Mn was recovered with D2EHPA extraction followed by two-step stripping with H2SO4 solution. Lithium ion-sieve was synthesized by oxidation of the stripping solution using KMnO4, then roasting at atmosphere followed by acid leaching. The effects of solvent extraction parameters on Mn recovery rate and Co doping content on the performance of lithium ion-sieve were investigated. The result indicates that through secondary countercurrent solvent extraction, Mn is fully separated from Ni and a MnSO4 solution containing a small amount of Co is obtained. Co can incorporate into the lattice of lithium ion-sieve and improve the structure stability and adsorption capacity. The Co doped lithium ion-sieve shows Li adsorption capacity of 33.1 mg g−1 and Mn dissolution rate of 1.39%. The total recovery efficiency of Mn is 84.6%. The process provides a simple method for recovery and utilization of Mn in the spent lithium-ion batteries.

Recovery of yttrium from deep-sea mud

Deep-sea mud rich in rare earth yttrium has received lots of attention from the international community as a new resource for Y. A novel process, which mainly includes acid leaching, solvent extraction, and oxalic acid precipitation-roasting, is proposed for recovery of Y from deep-sea mud. A series of experiments were conducted to inspect the impacts of various factors during the process and the optimum conditions were determined. The results show that the Y of deep-sea mud totally exists in apatite minerals which can be decomposed by hydrochloric acid and sulfuric acid solution. The highest leaching efficiency of Y is 94.53% using hydrochloric acid and 84.38% using sulfuric acid under the conditions of H+ concentration 2.0 mol/L, leaching time 60 min, liquid–solid ratio 4:1 and room temperature 25 °C (only in case of sulfuric acid, when using hydrochloric acid, the leaching temperature should be 60 °C). Because of the much lower leaching temperature, sulfuric acid leaching is preferred. The counter current extraction and stripping tests were simulated by a cascade centrifugal extraction tank device. Using 10 vol% P204, 15 vol% TBP and 75 vol% sulfonated kerosene as extractant, 98.79% Y3+ and 42.60% Fe3+ are extracted from sulfuric acid leaching liquor (adjusted to pH = 2.0) after seven-stage counter current extraction with O/A ratio of 1:1 at room temperature, while other metals ions such as Al3+, Ca2+, Mg2+ and Mn2+ are almost not extracted. The Y3+ in loaded organic can be selectively stripped using 50 g/L sulfuric acid solution and the stripping efficiency reaches 99.86% after seven-stage counter current stripping with O/A ratio of 10:1 at room temperature, while only 2.26% co-extracted Fe3+ is stripped. The Y3+ of loaded strip liquor can be precipitated by oxalic acid to further separate Y3+ and Fe3+. The precipitation efficiency of Y3+ in loaded strip liquor can be 98.56% while Fe3+ is not precipitated under the conditions of oxalic acid solution concentration 200 g/L, quality ratio of oxalic acid to Y of 2, and precipitation time 0.5 h. And the precipitate was roasted at 850 °C for 3 h to obtain the oxide of Y in which the purity of Y2O3/REO is 79.02% and the contents of major non-rare earth impurities are less than 0.21%. Over the whole process included acid leaching, solvent extraction, and oxalic acid precipitation-roasting, the yttrium yield is 82.04%.

Recovery of copper from copper slag using a microbial fuel cell and characterization of its electrogenesis

The microbial fuel cell, which can convert the chemical energy of organic matter into electricity via the catalytic action of microorganisms, is a novel environmentally friendly technology for wastewater treatment and energy generation. The electrical energy generated by the microbial fuel cell can be used as an alternative to a traditional external power source required to extract copper via electrolytic treatment. A dual-chamber microbial fuel cell (DMFC) for the treatment of copper slag sulfuric acid leach liquor was constructed. The electrogenesis performance of the DMFC and its ability to extract copper from the copper slag leachate were investigated. The results demonstrated that the maximum voltage was 540 mV when the DMFC achieved steady-state operation. The removal rate of copper ions was greater than 80.0%, and the maximum value was 92.1%. Moreover, X-ray diffraction and scanning electron microscopy were used to characterize the cathodal products. The results showed that the product deposited onto the cathode was copper and that its morphology was similar to that of the electrolytic copper powder. The DMFC can generate electricity and recover copper from copper slag simultaneously.

Recycling of Electric Vehicle Lithium Ion Batteries: Development of a Novel Method to Recover Lithium from Leached Cathodes Based on Donnan Dialysis

Lithium ion batteries (LIBs) are the battery of choice for electric vehicles (EVs) at present and for the foreseeable future. Adoption of EVs at a rate as high as possible is highly desirable since EVs have no direct emissions of green-house gases, smog causing chemicals, or particulate matter. The IEA has developed a roadmap to achieve 50% electrification of passenger vehicles by 2050. Unfortunately the literature shows that by the mid 2020’s lithium producers will not be able to keep up with the demand forecasted by the roadmap without recycling of EV LIBs [1]. To avoid a bottleneck, we calculated that depleated EV LIBs should be recycled and the lithium contained therein be used to make new LIBs at a rate of 28% and 75% in 2025 and 2040 (if the batteries last 5 years) or 40% and 96% in 2025 and 2035 respectively (if the batteries last 10 years). Currently, LIB cathodes are recycled by leaching the cathodes in acid, recovering transition metals one at a time by solvent extraction, and recovering lithium as lithium carbonate by precipitation with sodium carbonate. Leaching efficiencies of ~100% have been reported, but precipitation efficiencies for lithium have stagnated at 80% for the last 17 years [2], [3]. Additionally the highest purity lithium carbonate purity reported is 99.18% [3]. The purity required for EV LIBs is 99.9 to 99.99%. We are developing a novel process to recover lithium from leachate (see the figure below) based on Donnan dialysis with cation exchange membranes (CEM). Lithium ions are extracted first with a monovalent CEM ({1} in the figure) then ions of the other cathode metals (Co, Mn, and Ni) with a polyvalent CEM ({2} in the figure). Lithium is recovered by: neutralizing the acidic Li solution obtained from the first monovalent CEM ({1} in the figure); exchanging Li ions with K ions from a potassium bicarbonate solution pressurized with CO2 with another monovalent CEM ({3} in the figure); depressurizing the Li enriched KHCO3 solution, which results in the precipitation of lithium carbonate (the desired product); adding KHCO3 to make up the K ions exchanged; re-pressurising; and recycling the entire solution past the monovalent cation exchange membrane ({3} in the figure). The transition metal salts recovered in step {2} may be separated by solvent extraction or crystalized and used mixed in making new cathodes with the same chemistry as the leached cathodes. The work we have performed so far is equilibrium calculations on each of the stages and preliminary experiments to determine the mass transfer coefficients across the membranes. For steps {1} and {2} we assumed three counter-current equilibrium stages and stripping acid [H+] =0.90 M. We assumed a leach solution concentration of [Li+] = 0.285 M, [Co++] = [Ni++] = [Mn++] = 0.095 M for step {1} and [Co++] = [Ni++] = [Mn++] = 0.095 M for step {2}. We obtained a theoretical lithium recovery of 97.6% in step {1}, and a Co, Ni, and Mn theoretical recoveries of 99.1% in step {2}. For step {3} we assumed an incoming [Li+] = 0.278 M, a stripping solution of [KHCO3] =2 M and [Li+] =0.0935 M (this is the solubility limit of LiHCO3 prior to re-pressurization with CO2). The theoretical lithium recovery using three counter-current equilibrium stages was calculated to be 95.3%. Thus the theoretical overall maximum lithium recovery rate given by the process is 93.0%, which is superior to values presented in literature and is achievable in a process free of sodium, where leaching solution is recyclable, and where there is no heating. Our preliminary experiments reveal that using a Neosepta CMS monovalent cation exchange membrane the transmembrane mass transfer coefficients of Mn++, Co++, Ni++, when these ions are present at concentrations between 0.005 and 0.095 M and when the stripping acid [H+] is between 0.10 and 0.90 M, are less than 1% of the coefficient of lithium. Thus, the Neosepta CMS monovalent CEM provides an excellent separation of lithium from cobalt, nickel, and manganese. [1] H. Vikström, S. Davidsson, and M. Höök, “Lithium availability and future production outlooks,” Appl. Energy, vol. 110, pp. 252–266, Oct. 2013. [2] P. Zhang, T. Yokoyama, O. Itabashi, Y. Wakui, T. M. Suzuki, and K. Inoue, “Hydrometallurgical process for recovery of metal values from spent lithium-ion secondary batteries,” Hydrometallurgy, vol. 47, no. 2–3, pp. 259–271, 1998. [3] X. Chen, Y. Chen, T. Zhou, D. Liu, H. Hu, and S. Fan, “Hydrometallurgical recovery of metal values from sulfuric acid leaching liquor of spent lithium-ion batteries,” Waste Manag., vol. 38, pp. 349–356, 2015. Figure 1

Hydrometallurgical recovery of metal values from sulfuric acid leaching liquor of spent lithium-ion batteries

Environmentally hazardous substances contained in spent Li-ion batteries, such as heavy metals and nocuous organics, will pose a threat to the environment and human health. On the other hand, the sustainable recycling of spent lithium-ion batteries may bring about environmental and economic benefits. In this study, a hydrometallurgical process was adopted for the comprehensive recovery of nickel, manganese, cobalt and lithium from sulfuric acid leaching liquor from waste cathode materials of spent lithium-ion batteries. First, nickel ions were selectively precipitated and recovered using dimethylglyoxime reagent. Recycled dimethylglyoxime could be re-used as precipitant for nickel and revealed similar precipitation performance compared with fresh dimethylglyoxime. Then the separation of manganese and cobalt was conducted by solvent extraction method using cobalt loaded D2EHPA. And McCabe–Thiele isotherm was employed for the prediction of the degree of separation and the number of extraction stages needed at specific experimental conditions. Finally, cobalt and lithium were sequentially precipitated and recovered as CoC2O4⋅2H2O and Li2CO3 using ammonium oxalate solution and saturated sodium carbonate solution, respectively. Recovery efficiencies could be attained as follows: 98.7% for Ni; 97.1% for Mn, 98.2% for Co and 81.0% for Li under optimized experimental conditions. This hydrometallurgical process may promise a candidate for the effective separation and recovery of metal values from the sulfuric acid leaching liquor.

Recovery of manganese and nickel from polymetallic manganese nodule using commercial extractants

The paper describes a process for the recovery of manganese and nickel from the sulfuric acid leach liquor of Indian Ocean manganese nodules using starch as the reductant. The leach liquor contained 22.85g/L manganese, 6.38g/L iron, 1.01g/L copper, 0.023g/L zinc, 0.09g/L cobalt and 1.44g/L nickel. From this leach liquor, Fe was first precipitated out with Ca(OH)2 at pH3.8 followed by the extraction of copper using LIX 84I. From the almost Fe- and Cu-free leach liquor, Zn was extracted out as an impurity with D2EHPA in kerosene, after which the extraction of Mn was carried out with NaD2EHPA. The extraction efficiency of manganese using NaD2EHPA was 99.93% in 2-stages at A:O ratio of 3:4. The co-extraction of cobalt with manganese was 43mg/L which was subsequently removed by scrubbing in two stages with MnSO4 solution. The stripping of manganese was achieved with 4% H2SO4 in 2-stages at A:O ratio of 1:1. After extraction of Mn, Ni was extracted with NaD2EHPA. The extraction isotherm of nickel reads 2-stages at A:O ratio of 3:2. The co-extraction of cobalt with nickel was 47mg/L for which the Ni-LO was scrubbed out with NiSO4 solution. The stripping efficiency of nickel was 99.93% with 0.5% H2SO4 in 2-stages at A:O ratio of 1:1.

Successive LEM Separation of Copper and Uranium From Abu Zeneima Mineralized Dolostone Using LIX-973N and Aliquat-336

We processed a technological sample of a mineralized dolostone rock facies of SW Sinai, assaying 5000 ppm copper and 750 ppm uranium for the purpose of recovering of its metal values. The prepared sulfuric acid leach liquor was successfully subjected to two successive emulsion liquid membrane procedures using LIX-973N encapsulating sulfuric acid followed by Aliquat-336 encapsulating sodium carbonate solution for copper and uranium extraction, respectively. A prior iron separation step from the leach liquor was found necessary in order to avoid or minimize its concentration. The relevant extraction factors in both circuits have been studied for proper optimization.

Thermodynamic Behavior and Applications of KMnO4 in Cerium Separation Process Using Cyanex 923

In this work, studies were carried out on the extraction properties of Mn(II) and MnO4- in sulfuric acid medium using Cyanex 923. Effects of different variables on the extraction of Mn(II) and MnO4-, such as the concentrations of acid, the extractant, and the temperature, were investigated. Results indicated that Mn(II) was extracted weakly by Cyanex 923; however, MnO4- could be strongly extracted into the organic phase. The extraction mechanism of MnO4- was proposed, and the influence of MnO4- on the extraction of cerium was identified when KMnO4 as oxidizer added into the bastnasite sulfuric acid leaching liquor. As MnO4- was easier to be extracted into the organic phase than Ce(IV) and then lost its ability for oxidization, a new device was designed to realize sufficient oxidization of cerium from III to IV, and which has been applied to industrialization.