Selective leaching of lithium from spent lithium-ion batteries using sulfuric acid and oxalic acid

Traditional hydrometallurgical methods for recovering spent lithium-ion batteries(LIBs)involve acid leaching to simultaneously extract all valuable metals into the leachate.These methods usually are followed by a series of separation steps such as precipitation,extraction,and stripping to separate the individual valuable metals.In this study,we present a process for selectively leaching lithium through the synergistic effect of sulfuric and oxalic acids.Under optimal leaching conditions(leaching time of 1.5 h,leaching temperature of 70°C,liquid-solid ratio of 4 mL/g,oxalic acid ratio of 1.3,and sulfuric acid ratio of 1.3),the lithium leaching efficiency reached89.6%,and the leaching efficiencies of Ni,Co,and Mn were 12.8%,6.5%,and 21.7%.X-ray diffraction(XRD)and inductively coupled plasma optical emission spectrometer(ICP-OES)analyses showed that most of the Ni,Co,and Mn in the raw material remained as solid residue oxides and oxalates.This study offers a new approach to enriching the relevant theory for selectively recovering lithium from spent LIBs.

Recovery of Li, Ni, Co and Mn from spent lithium-ion batteries assisted by organic acids: Process optimization and leaching mechanism

The proper recycling of spent lithium-ion batteries(LIBs)can promote the recovery and utilization of valuable resources,while also negative environmental effects resulting from the presence of toxic and hazardous substances.In this study,a new environmentally friendly hydro-metallurgical process was proposed for leaching lithium(Li),nickel(Ni),cobalt(Co),and manganese(Mn)from spent LIBs using sulfuric acid with citric acid as a reductant.The effects of the concentration of sulfuric acid,the leaching temperature,the leaching time,the solid-liquid ratio,and the reducing agent dosage on the leaching behavior of the above elements were investigated.Key parameters were optimized using response surface methodology(RSM)to maximize the recovery of metals from spent LIBs.The maxim-um recovery efficiencies of Li,Ni,Co,and Mn can reach 99.08%,98.76%,98.33%,and 97.63%.under the optimized conditions(the sulfuric acid concentration was 1.16 mol/L,the citric acid dosage was 15wt%,the solid-liquid ratio was 40 g/L,and the temperature was 83℃ for 120 min),respectively.It was found that in the collaborative leaching process of sulfuric acid and citric acid,the citric acid initially provided strong reducing CO_(2)^(-),and the transition metal ions in the high state underwent a reduction reaction to produce transition metal ions in the low state.Additionally,citric acid can also act as a proton donor and chelate with lower-priced transition metal ions,thus speeding up the dissolution process.

Extraction of Valuable Metals from Titanium-bearing Blast Furnace Slag by Acid Leaching

To realize the resource utilization of the valuable metals in the titanium-containing blast furnace slag,the process route of “hydrochloric acid leaching-electrolysis-carbonization and carbon dioxide capture-preparation of calcium carbonate” was proposed.In this study,the influences of process conditions on the leaching rates of calcium,magnesium,aluminum,and iron and the phases of the leaching residue were investigated for the leaching process.The experimental results show that the HCl solution could selectively leach the elements from the titanium-containing blast furnace slag.The better leaching conditions are the HCl solution concentration of 4 mol/L,the leaching time of 30 min,the ratio of liquid volume to solid gas of 10 mL/g,and the stirring paddle speed of 300 r/min.Under the conditions,the leaching rates of calcium,magnesium,aluminum,and iron can reach 85.87%,73.41%,81.35%,and 59.08%,and the leaching rate of titanium is 10.71%.The iron and the aluminum are removed from the leachate to obtain iron-aluminum water purification agents,and the magnesium is removed from the leachate to obtain magnesium hydroxide.The leaching residue phase is dominated by perovskite,followed by magnesium silicate and tricalcium aluminate,and the titaniumrich material could be obtained from the leaching residue by desiliconization.

Effect of Saline Water on Soil Acidity, Alkalinity and Nutrients Leaching in Sandy Loamy Soil in Rwamagana Bella Flower Farm, Rwanda

The necessity to saline and sodic waters is sometimes used for irrigating agricultural activities under certain circumstances, but it is important to note that the use of these waters comes with specific considerations and limitations. One way to decrease undesirable effects of sodic waters on the physical and chemical properties of soils is to apply organic and chemical amendments within the soil. This study aimed to assess the effectiveness of saline water on soil acidity, alkalinity and nutrients leaching in sandy loamy soil at Bella flower farm, in Rwamagana District, Rwanda. The water used was from the Muhazi Lake which is classified as Class I (Saline water quality). Column leaching experiments using treated soils were then conducted under saturated conditions. The soil under experimental was first analyzed for its textural classification, soil properties and is classified as sandy loamy soil. The t-test was taken at 1%, 5% and 10% levels of statistical significance compared to control soil. The results indicated that the application of saline water to soils caused an increase in some soil nutrients like increase of Phosphorus (P), Potassium (K+), Magnesium (Mg2+), Sulphur (S), CN ratio and Sodium (Na+) and decreased soil texture, physical and chemical properties and remained soil nutrients. Consequently, the intensive addition of saline water leachates to soil in PVC pipes led to decreased of soil EC through leaching and a raiser Soluble Sodium Percentage (SSP). The rate of saline water application affected the increase accumulation of SAR and Na% in the top soil layers. The study indicated that saline water is an inefficient amendment for sandy soil with saline water irrigation. The study recommends further studies with similar topic with saline water irrigation, as it accentuated the alkalinity levels.

Process engineering of demineralisation of moderate to high ash Indian coals through NaOH‑HCl leaching and HF leaching

Chemical leaching of coals would be required to produce cleaner coals for some special applications where physical benefi-ciation may not be effective enough.This would also help in recovering Li and rare earth metals besides in the sequestration of CO_(2).About 20 Indian coals having complexly distributed moderate to high ash contents were sequentially treated with various alkali–acid such as NaOH-HCl,HF,HCl,HCl-HF,and NaOH-HCl-HF leaching.This aimed to establish and design the best stepwise sequential process for the highest degree of demineralisation through a chemical leaching process.Kinetics and process intensification studies were carried out.More than 80%demineralisation of Madhaipur and Neemcha coals was observed using the best sequential treatment designed presently.The repeated stepwise treatment of the alkali and the acid was also studied,which was found to significantly enhance the degree of demineralisation of coals.The integrated process of alkali–acid leaching followed by solvent extraction(Organo-refining)and vice versa of the treated coal was also studied for producing cleaner coals.

Leaching behavior of metals from limonitic laterite ore by high pressure acid leaching

The leaching behavior of metals from a nickeliferous limonitic laterite ore was investigated by high pressure acid leaching process for the extraction of nickel and cobalt.The effects of sulfuric acid added,leaching temperature,leaching time and liquid/solid（L/S） ratio on metals extraction were examined.More than 97% Ni,96% Co,93% Mn,95% Mg and less than 1% Fe are extracted under optimum conditions.Analysis of the high pressure acid leaching residue by chemical and XRD analysis indicates that the residual iron and sulfur are mainly present in phases of hematite and alunite,respectively.The high pressure leaching process provides a simple and efficient way for the high recovery of nickel and cobalt from laterite ore,leaving residue as a suitable iron resource.

Preparation of Ti-rich material from titanium slag by activation roasting followed by acid leaching

A method of activation roasting followed by acid leaching using titanium slag was introduced to prepare Ti-rich material. The effects of HaPO4 dosage, roasting temperature, and roasting time on TiO2 grade were investigated. A Ti-rich material containing 88.54% TiO2, 0.42% （CaO＋MgO） was obtained when finely ground titanium slag was roasted with 7.5% H3PO4 at 1000 ℃ for 2 h, followed by a two-stage leaching in boiling dilute sulfuric acid for 2 h. The XRD patterns show that the product is titanium dioxide with a rutile structure. Mechanism studies show that structures of anosovite solid solution and silicate minerals are destroyed in the roasting process. As a result, titanium components in titanium slag are transformed into TiO2 （futile） while impurities are transformed into acid-soluble phosphate and quartz.

Kinetics of titanium leaching with citric acid in sulfuric acid from red mud

The recovery of titanium with citric acid in sulfuric acid from red mud was put forward to strengthen acid leaching efficiency.The main factors on the recovery of titanium such as citric acid addition,sulfuric acid concentration,leaching temperature,time and liquid-to-solid ratio were studied.The kinetics analysis of titanium leaching from red mud was deeply investigated.The results show that the citric acid could increase the recovery of titanium and decrease the consumption of sulfuric acid.The recovery of titanium was increased from 65% to 82% and the consumption of sulfuric acid was decreased by about 30% with using 5% citric acid.The dissolution of perovskite,brookite,and hematite in red mud could easily be dissolved using citric acid.The acid leaching process was controlled by internal diffusion of shrinking core model（SCM） and the correlation coefficient was above 0.98.The apparent rate constant was increased from 0.0012 to 0.0019 with 5% citric acid at 90 °C.The apparent activation energy of titanium leaching decreased from 39.77 k J/mol to 34.61 k J/mol with 5% citric acid.

Acid leaching kinetics of zinc plant purification residue

The leaching processes of zinc plant purification residue in sulfuric acid solution were investigated with respect to the effects of sulfuric acid concentration, reaction temperature, and particle size. A particle size of 75？80 μm was required to leach 99.8%cobalt and 91.97%zinc at 70 °C for 20 min when the sulfuric acid concentration was 100 g/L and the ratio of liquid to solid was 50？1 （mL/g）. The leaching kinetics of zinc plant purification residue in sulfuric acid solution system conformed well to the shrinking core model, and the dissolution rates of cobalt and zinc were found to be controlled by diffusion through a porous product layer. The apparent activation energy values of cobalt and zinc reaction were calculated to be 11.6931 kJ/mol and 6.6894 kJ/mol, respectively, according to the Arrhenius formula equation. The results show that diffusion through the inert particle pores is the leaching kinetics rate-controlling step.

Thermodynamics study on leaching process of gibbsitic bauxite by hydrochloric acid

For the low-grade gibbsitic bauxite,the leaching rate of alumina is very low during the Bayer process.The acid leaching method is attracting more attention,and the hydrochloric acid leaching was developed rapidly.The mineral composition and chemical composition were investigated by X-ray diffraction analysis and semi-quantitative analysis.The thermodynamics of leaching process was analyzed.The results show that the major minerals in the bauxite are gibbsite,secondly goethite and quartz,anatase and so on.The acid leaching reactions of the bauxite would be thermodynamically easy and completed.Under the conditions that ore granularity is less than-55 μm,the L/S ratio is 100：7,and the leaching temperature is 373-383 K,the leaching time is 120 min and the concentration of HCl is 10%,both the leaching rates of Al and Fe are over 95%.The main composition of leaching slag is SiO2 which is easy for comprehensive utilization.

Reductive acid leaching of cadmium from zinc neutral leaching residue using hydrazine sulfate

Zinc neutral leaching residue（ZNLR） from hydrometallurgical zinc smelting processing can be determined as hazardous intermediate containing considerable amounts of Cd and Zn which have great threats to the environment. The ZNLR contained approximately 35.99% Zn, 15.93% Fe and 0.26% Cd, and Cd mainly existed as ferrites in the ZNLR in this research. Reductive acid leaching of ZNLR was investigated. The effects of hydrazine sulfate concentration, initial sulfuric acid concentration, temperature, duration and liquid-to-solid ratio on the extraction of Cd, Zn and Fe were examined. The extraction efficiencies of Cd, Zn and Fe reached 90.81%, 95.83% and 94.19%, respectively when the leaching parameters were fixed as follows： hydrazine sulfate concentration, 33.3 g/L; sulfuric acid concentration, 80 g/L; temperature, 95 °C; duration of leaching, 120 min; liquid-to-solid ratio, 10 m L/g and agitation, 400 r/min. XRD and SEM-EDS analyses of the leaching residue confirmed that lead sulfate（Pb SO4） and hydrazinium zinc sulfate（（N2H5）2Zn（SO4）2） were the main phases remaining in the reductive leaching residue.

Effects of mechanical activation and oxidation-reduction on hydrochloric acid leaching of Panxi ilmenite concentration

The effects of oxidation-reduction treatment and mechanical activation on the hydrochloric acid leaching performance of Panxi ilmenite concentration were investigated.The results show that both of oxidation-reduction treatment and mechanical activation significantly accelerate the extraction of Fe,Ca and Mg from Panxi ilmenite concentration;however,the CaO and MgO contents of the calcined residues obtained from oxidized-reduced ilmenite concentration are higher than the standard values required by chlorination process.The Ca and Mg in oxidized-reduced ilmenite concentration can be leached much faster after mechanical activation,yielding a synthetic rutile which meets the requirements of chlorination process containing 90.50% TiO2 and 1.37% total iron as well as combined CaO and MgO of 1.00%.The optimum oxidation and reduction conditions are as follows：oxidization at 900 ℃ in the presence of oxygen for 15 min and reduction at 750 ℃ by hydrogen for 30 min.

φ-pH diagram of V-Ti-H_2O system during pressure acid leaching of converter slag containing vanadium and titanium

To analyze the thermodynamic characteristics of leaching process of converter slag, φ-pH diagram of V-Ti-H2O system at oxygen partial pressure of 0.5 MPa, ionic mass concentration of 0.1 mol/kg and temperatures ranging from 60 to 200 ℃ was obtained by recently published critically assessed standard Gibbs energies and activity coefficients of various species. When pH2, stable regions of V3＋, VO2＋ and VO2＋ exist in the stable region of TiO2. The pH values of stable regions of vanadium and titanium decrease and redox potentials become more positive with the temperature increasing. Vanadium and titanium could be separated by one-step leaching based on thermodynamics. The experiment results of pressure acid leaching of converter slag show that leaching rates of vanadium and titanium are 96.87% and 8.76% respectively, at 140 ℃ of temperature, 0.5 MPa of oxygen partial pressure, 0.055-0.075mm of particle size, 15：1 of liquid to solid ratio, 120 min of leaching time, 500 r/min of stirring speed and 200 g/L of initial acid concentration. Vanadium and titanium could be selectively separated in the pressure acid leaching process, and the experiment result is in agreement with thermodynamic calculation result.

Leaching behavior of heavy metals with hydrochloric acid fromfly ash generated in municipal waste incineration plants

The extraction behavior of heavy metals from municipal waste incineration （MWI） fly ash was investigated systematically. The extraction process includes two steps, namely, fly ash was firstly washed with water, and then subjected to hydrochloric acid leaching. The main parameters for water washing process were tested, and under the optimal conditions, about 86% Na, 70% K and 12% Ca were removed from fly ash, respectively. Hydrochloric acid was used for the extraction of valuable elements from the water-washed fly ash, and the optimal extraction was achieved for each heavy metal as follows： 86% for Pb, 98% for Zn, 82% for Fe, 96% for Cd, 62% for Cu, 80% for Al, respectively. And the main compositions of the finally obtained solid residue are Ca2PbO4, CaSi2Os, PbsSiO7, Ca3A12Si3012 and SiO2.

Rapid Determination of Cd in Squilla by Mixed Acid Leaching-flame Atomic Absorption Method

[Objective] This study was conducted to find a method for rapid determi-nation of Cd in squil a. [Method] Cd in squil a was determined by mixed acid leach-ing-flame atomic absorption method by digesting samples with HNO3-HSO4 （5：1） and determining Cd in the samples by a flame atomic absorption spectrophotometer. [Results] The Cd contents in squil a were 1.0-2.0 mg/kg, and the recoveries of Cd were 86%-109%, and the relative standard deviations were 0-6%. [Conclusion] The method is simple and rapid with strong operability, good reproducibility and high ac-curacy.

Extraction of molybdenum and nickel from roasted Ni-Mo ore by hydrochloric acid leaching, sulphation roasting and water leaching

To extract molybdenum and nickel from the roasted Ni-Mo ore, a process of hydrochloric acid leaching, sulphation roasting and water leaching was investigated. The results showed that this process could get a high leaching rate of Mo and Ni. Under the optimum conditions of hydrochloric acid leaching （roasted Ni-Mo ore leached with 0.219 mL/g hydrochloric acid addition at 65 ℃ for 30 min with a L/S ratio of 3 mL/g）, sulphation roasting （51.9% sulfiaric acid addition, roasting temperature 240 ℃ for 1 h）, followed by leaching with the first stage hydrochloric acid leaching solution at 95 ℃ for 2 h, the leaching rates of Mo and Ni reached 95.8% and 91.3%, respectively.

Leaching of vanadium from stone coal with sulfuric acid

The effects of roasting, mass ratio of H2SO4 to stone coal, leaching temperature, liquid-to-solid ratio, grinding fineness of stone coal, and two-stage counter-current leaching on the vanadium leaching ratio were studied. The results show that the vanadium leaching ratio of roasted stone coal through two-stage counter-current leaching can reach 65.1% at the mass ratio of H2SO4 to stone coal of 20%, leaching temperature of 95℃, leaching time of 30 h, liquid-to-solid ratio of 1.1 mL·g^-1, and grinding fineness of 0.1 mm, which can serve as an experimental basis for the production of vanadium from stone coal.

Vanadium recovery from clay vanadium mineral using an acid leaching method

A technique including direct acid leaching, vanadium precipitation with alkaline, sodium hydroxide releaching, impurity removing by adjusting pH value, precipitation vanadium with ammonium chloride, and vanadium pentoxide by roasting steps was proposed according to the characteristic of Xichuan clay vanadium mineral. The factors influencing leaching vanadium such as temperature and the concentration of sulfuric acid were investigated and optimized. The experimental results indicate that the extract ratios of V205 can reach 94% and 92% at a sodium chlorate ratio of 3% and a manganese dioxide ratio of 3%, respectivdy. A completely chemical precipitation method was adopted to decontaminate and enrich the vanadium in the acid leaching solution. The X-ray diffraction （XRD） pattern and the purity analysis of vanadium pentoxide indicate that the purity of final vanadium pentoxide can reach 99% and meet the standard specifications. The total recovery can reach about 75%. The technique has the characteristics of simplicity, less investment, and more environment safety as compared with the traditional salt roasting method.

Extracting vanadium from stone-coal by oxygen pressure acid leaching and solvent extraction

Vanadium extraction from stone-coal was investigated by oxygen pressure acid leaching and solvent extraction.The mineralogy of the stone-coal from Tongren City of Guizhou Province,China,was investigated by various determination methods. The effects of leaching time,leaching temperature,leaching agent concentration,leaching L/S ratio,granularity of material,additive consumption were investigated based on the mineralogy.The results show that under the conditions of leaching time of 3-4 h, temperature of 150℃,sulfuric acid consumption of 25%?30%,ratio of liquid to solid of 1.2:1,the granularity less than 0.074 mm, additive consumption of 3%-5%,and oxygen pressure of 1.2 MPa,and the vanadium leaching rate can be more than 92%by the method of two-step pressurized acid leaching.The powdery V2O5 product with 99.52%in V2O5 content is obtained by the flowsheet of acid recovery,removing iron by reduction process,solvent extraction,precipitating vanadium with ammonium water,and pyrolysis from the stone-coal oxygen pressure acid-leaching solution.The total recovery efficiency of vanadium is above 85%,which is more than 20%higher than that obtained in the conventional process.Furthermore,the new process does not cause air pollution since no HCl or Cl2 is released by calcination of the raw material.

Extraction of vanadium from vanadium slag by high pressure oxidative acid leaching

To extract vanadium in an environment friendly manner, this study focuses on the process of leaching vanadium from vanadium slag by high pressure oxidative acid leaching. Characterizations of the raw slag, mineralogy transformation, and the form of leach residues were made by X-ray diffraction, scanning electron microscopy, and energy dispersive X-ray spectroscopy. The result shows that the vana-dium slag is composed of major phases of fayalite, titanomagnetite, and spinel. During the high pressure oxidative acid leaching process, the fayalite and spinel phases are gradually decomposed by sulfuric acid, causing the release of vanadium and iron in the solution. Meanwhile, unreacted silicon and titanium are enriched in the leach residues. With the initial concentration of sulfuric acid at 250 g·L^-1, a leaching tem-perature of 140℃, a leaching time of 50 min, a liquid-solid ratio of 10：1 mL·g^-1, and oxygen pressure at 0.2 MPa, the leaching rate of vana-dium reaches 97.69%.