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.

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.

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.

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.

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.

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.

Effect of Na_(2)CO_(3)Roasting Activation on Acid Leaching Property of Fly Ash

By using high-alumina fly ash as raw material,a process was proposed for activating the fly ash with Na_(2)CO_(3)calcination and extracting aluminum from activated clinker with sulfuric acid leaching.The feasibility of roasting process of activated fly ash by Na_(2)CO_(3)was discussed based on thermodynamic analysis.The experimental results showed that Na_(2)CO_(3)gradually reactes with mullite over 700 K to produce NaAlSiO_(4).The optimal process conditions for the activation stage are:a material ratio of 1:1 between sodium carbonate and fly ash,a calcination temperature of 900℃,and a calcination time of 2.5 hours.Under these conditions,the leaching rate of aluminum is 90.3%.By comparing the SEM and XRD analysis of raw and clinker materials,it could be concluded that the mullite phase of fly ash is almost completely destroyed and transformed into sodium aluminosilicate with good acid solubility.

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.

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.

Process mineralogy characteristics of acid leaching residue produced in low-temperature roasting-acid leaching pretreatment process of refractory gold concentrates

To provide a theoretical basis for a suitable process to extract gold from refractory gold concentrates, process mineralogy on the acid leaching residue of gold calcine was studied by chemical composition, X-ray diffraction, scanning electron microscopy-energy spectrum, and mineral dissociation analysis. The results showed that the acid leaching residue contained Au 68.22 g/t, Ag 92.71 g/t, Fe 0.44%, As 0.10%, and S 0.55%. Gold and silver minerals existed as native gold, argentite, and proustite. Quartz, the main gangue mineral, accounted for 78.33 wt/%. The dissociation degree analysis showed that the proportions of monomer and exposed gold in acid leaching residue were 96.66 wt%. The cyanidation results showed that the cyanide gold leaching rate of acid leaching residues was close to 100 wt%. However, the maximum cyanide gold leaching rate of gold calcine was only 85.31 wt%. This suggests that acid leaching can increase the gold dissolution rate in the cyanide process.

Reductive acid leaching of valuable metals from spent lithium-ion batteries using hydrazine sulfate as reductant

Hydrazine sulfate was used as a reducing agent for the leaching of Li,Ni,Co and Mn from spent lithium-ion batteries.The effects of the reaction conditions on the leaching mechanism and kinetics were characterized and examined.97%of the available Li,96%of the available Ni,95%of the available Co,and 86%of the available Mn are extracted under the following optimized conditions:sulfuric acid concentration of 2.0 mol/L,hydrazine sulfate dosage of 30 g/L,solid-to-liquid ratio of 50 g/L,temperature of 80℃,and leaching time of 60 min.The activation energies of the leaching are determined to be 44.32,59.37 and 55.62 k J/mol for Li,Ni and Co,respectively.By performing X-ray diffraction and scanning electron microscopy in conjunction with energy dispersive X-ray spectroscopy,it is confirmed that the main phase in the leaching residue is MnO2.The results show that hydrazine sulfate is an effective reducing agent in the acid leaching process for spent lithium-ion batteries.

Leaching kinetics of selenium from copper anode slimes by nitric acid-sulfuric acid mixture

The leaching kinetics of selenium from copper anode slimes was studied in a nitric acid?sulfuric acid mixture.The effects of main parameters on selenium leaching showed that the leaching rate of selenium was practically independent of stirring speed,while dependent on temperature and the concentrations of HNO3and H2SO4.The leaching of selenium includes two stages.The activation energy in the first stage is103.5kJ/mol,and the chemical reaction is the rate controlling step.It was almost independent of H2SO4concentration and dependent on HNO3concentration since the empirical reaction order with respect to HNO3concentration is0.5613.In the second stage,the activation energy is30.6kJ/mol,and the process is controlled by a mixture of diffusion and chemical reaction.The leaching of selenium was almost independent of HNO3concentration.

An investigation of oxygen pressure acid leaching of Gacun complex Cu-Pb bulk concentrate

The treatment of the Gacun complex Cu-Pb bulk concentrate with high Zn,Ag,etc.,by oxygen pressure acid leaching was studied.The pri-mary copper and leadminerals in the concentrate are tetrahedrite and galena.The treatment of tetrahedrite was rarely studied,and most of silver occurred in themineral too.The optimum operating parameters of oxygen pressure acid leaching were established by conditional tests.Under these parameters,the result of pilot scale test showed that the leaching percentages of copper and zinc were separately as high as 98.9 wt.% and 94.9 wt.%,while lead and silver were transformed into sulfate and sulfide precipitations,respectively.The copper and zinc in lixivium were reclaimed by extraction-electrowinning and purification-electrowinning,respectively,and the lead and silver in the residue were reclaimed separately by carbonate transformation-silicofluoric acid leaching and thiourea leaching.

Separation and recovery of copper in Cu-As-bearing copper electrorefining black slime by oxidation acid leaching and sulfide precipitation

A new hydrometallurgical route for separation and recovery of Cu from Cu-As-bearing copper electrorefining black slime was developed. The proposed process comprised oxidation acid leaching of Cu-As-bearing slime and selective sulfide precipitation of Cu from the leachate. The effects of various process parameters on the leaching and precipitation of Cu and As were investigated. At the first stage, Cu extraction of 95.2% and As extraction of 97.6% were obtained at 80 ℃ after 4 h with initial H2 SO4 concentration of 1.0 mol/L and liquid-to-solid ratio of 10 mL/g. In addition, the leaching kinetics of Cu and As was successfully reproduced by the Avrami model, and the apparent activation energies were found to be 33.6 and 35.1 kJ/mol for the Cu and As leaching reaction, respectively, suggesting a combination of chemical reaction and diffusion control. During the selective sulfide precipitation, about 99.4% Cu was recovered as CuS, while only 0.1% As was precipitated under the optimal conditions using sulfide-to-copper ratio of 2.4:1, time of 1.5 h and temperature of 25 ℃.

Kinetics of leaching lithium from α-spodumene in enhanced acid treatment using HF/H_2SO_4 as medium

An enhanced leaching of Li fromα-spodumene was carried out using a mixture of hydrofluoric and sulfuric acid(HF/H2SO4)as the medium.Based on the optimized leaching conditions,the leaching kinetics of Li was investigated in an ore/HF/H2SO4 ratio of 1:3:2 g:mL:mL with leaching temperature ranging from 50 to 100°C.The results indicate that the leaching kinetics of Li fitted well with a model based on the shrinking core model.In addition,the leaching rate of Li was controlled by chemical reactions and diffusion through the product layers.The apparent activation energy Ea was calculated to be 32.68 kJ/mol.Solid films were formed because of the generation of insoluble products such as cryolithionite(Na3Li2Al2F12),cryolite(Na3AlF6),calcium fluoride(CaF2),potassium cryolite(K2AlF5),aluminum fluoride(AlF3),and fluorosilicates(Na2SiF6 or KNaSiF6).Furthermore,the effects of the ore/HF ratio and leaching temperature on the leaching behavior of Li,Al and Si were investigated.The results indicate that the ore/HF ratio and leaching temperature could clearly affect the distribution of HF molecules on the leaching of Li,Al and Si,which are important for the selective leaching of Li over Al and Si with this fluorine-based chemical method.

Sulfuric acid leaching of high iron-bearing zinc calcine

Sulfuric acid leaching of high iron-bearing zinc calcine was investigated to assess the effects of sulfuric acid concentration, liquid-to-solid ratio, leaching time, leaching temperature, and the stirring speed on the leaching rates of zinc and iron. The results showed that the sulfuric acid concentration, liquid-to-solid ratio, leaching time, and leaching temperature strongly influenced the leaching of zinc and iron, whereas stirring speed had little influence. Zinc was mainly leached and the leaching rate of iron was low when the sulfuric acid concentration was less than 100 g/L. At sulfuric acid concentrations higher than 100 g/L, the leaching rate of iron increased quickly with increasing sulfuric acid concentration. This behavior is attributed to iron-bearing minerals such as zinc ferrite in zinc calcine dissolving at high temperatures and high sulfuric acid concentrations but not at low temperatures and low sulfuric acid concentrations.

Kinetic study of high-pressure acid leaching of Mg and Ni from serpentine

Inner Mongolian serpentine ore was subjected to sulfuric acid leaching tests,and the effects of the leaching process parameters on the leaching efficiency of different metals were investigated.The leaching efficiency of Mg,Fe,Al,Ni,and Co reaches 93.98%,60.09%,82.08%,90.58%,and 94.06%,respectively,under the leaching conditions of 5 mol/L H_(2)SO_(4),liquid/solid ratio of 4 mL/g,and leaching temperature 100℃.Hence,the valuable metals in serpentine were effectively recovered by sulfuric acid leaching.The leaching behaviors of Mg,Fe,and a small amount of Al were analyzed using X-ray diffraction.The results show that the unreacted Mg and Fe remained as MgFe_(2)O_(4),and Al formed Al_(2)Si_(2)O_(5)(OH)_(4) in the leaching residue.The kinetics of Mg and Ni in the leaching process was studied respectively.The leaching kinetics of Mg conformed to the shrinking core model with an activation energy of 16.95 kJ/mol,which was controlled by the combination of the diffusion and chemical reaction.The leaching kinetics of Ni accorded with the Avrami equation with an activation energy of 11.57 kJ/mol,which was controlled by diffusion.In the study,the valuable metal elements were extracted from serpentine minerals with high efficiency and low cost,which possessed important practical values.

Leaching kinetics of acid-soluble Cr(Ⅵ) from chromite ore processing residue with hydrofluoric acid

Leaching kinetics of acid-soluble Cr（VI） in chromite ore processing residue （COPR） using hydrofluoric （HF） acid solution as a leaching agent was investigated for potential remediation of COPR with industrial waste water containing HF. The results show that HF can effectively destabilize the Cr（VI）-bearing minerals, resulting in the mobilization of Cr（VI） from COPR into the leachate. Particle size significantly influences the leaching of acid-soluble Cr（VI） from COPR, followed by leaching time, whereas the effects of HF concentration and leaching temperature are slight and the influence of stirring rate is negligible. The leaching process of acid-soluble Cr（VI） from COPR is controlled by the diffusion through the product layer. The apparent activation energy is 8.696 kJ/mol and the reaction orders with respect to HF concentration and particle size is 0.493 8 and -2.013 3, respectively.