Preparation of sodium manganate from low-grade pyrolusite by alkaline predesilication–fluidized roasting technique

Low concentration alkaline leaching was used for predesilication treatment of low-grade pyrolusite. The effects of initial NaOH concentration, liquid-to-solid ratio, leaching temperature, leaching time and stirring speed on silica leaching rate were investigated and the kinetics of alkaline leaching process was studied. The results show that silica leaching rate reached 91.2% under the conditions of initial NaOH concentration of 20%, liquid-to-solid ratio of 4:1, leaching temperature of 180 °C, leaching time of 4 h and stirring speed of 300 r/min. Shrinking-core model showed that the leaching process was controlled by the chemical surface reaction with activation energy E_a of 53.31 kJ/mol. The fluidized roasting conditions for preparation of sodium manganate were optimized by the orthogonal experiments using the desiliconized residue. The conversion rate of sodium manganate was obtained to be 89.7% under the conditions of silica leaching rate of 91.2%, NaOH/MnO₂ mass ratio of 3:1, roasting temperature of 500 °C and roasting time of 4 h, and it increased with the increase of silicon leaching rate.     

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 HNO₃ and H₂SO₄. The leaching of selenium includes two stages. The activation energy in the first stage is 103.5 kJ/mol, and the chemical reaction is the rate controlling step. It was almost independent of H₂SO₄ concentration and dependent on HNO₃ concentration since the empirical reaction order with respect to HNO₃ concentration is 0.5613. In the second stage, the activation energy is 30.6 kJ/mol, and the process is controlled by a mixture of diffusion and chemical reaction. The leaching of selenium was almost independent of HNO₃ concentration.     

从锌置换渣中分段浸出锌、铅、镓和锗

采用分段浸出方法从锌置换渣中选择性提取锌、铅、镓和锗.首段浸出过程通过控制合适的硫酸浓度与液固比条件,锌和镓的浸出率达到90％和99％以上,超过92％的锗留在硫酸浸出渣中;二段采用盐酸作为浸出剂选择性提取铅,铅的浸出率为99％,锗损失率小于2％;三段采用1 mol/L氢氧化钠作为浸出剂,通过破坏硅锗固溶体的结构,实现锗...     

Leaching of chalcopyrite with hydrogen peroxide in hydrochloric acid solution

The aim of this work was to investigate the leaching of chalcopyrite concentrate in hydrochloric acid with hydrogen peroxide as a strong oxidizing agent. The effects of the leaching variables on metal extraction, such as stirring speed, solid-to-liquid ratio, temperature and HCl and H₂O₂ concentrations, were studied. The maximum final copper extraction of 33% was attained with 3.0 mol/L H₂O₂ in 0.5 mol/L HCl at room temperature after 180 min of the reaction. The results showed that the copper extraction was increased in the first 60 min of reaction, after which it essentially ceased due to the fast catalytic decomposition of hydrogen peroxide. Further, solid-to-liquid ratio affected the copper extraction significantly and the highest copper extraction was obtained in the most dilute suspension (i.e., S/L ratio of 1:100). The dissolution process was described by the first order kinetics equation. The apparent activation energy of 19.6 kJ/mol suggested that the dissolution process was under diffusion control. The reaction orders for HCl and H₂O₂ were established to be 0.30 and 0.53, respectively. The results of the XRD and SEM/EDS analysis of the leaching residue indicated the generation of the elemental sulphur on mineral surfaces which tended to inhibit the leaching rate.     

Leaching of cuprite through NH4OH in basic systems

Cuprite is a difficult oxide to leach under acidic conditions (for the maximum extraction of 50%). In this research, the feasibility of leaching cuprite in an ammoniacal medium was studied. The working conditions addressed here were the liquid/solid ratio (120:1–400:1 mL/g), stirring speed (0–950 r/min), temperature (10–45 °C) and NH₄OH concentration (0.05–0.15 mol/L). In addition, different ammoniacal reagents (NH₄F and (NH₄)₂SO₄) were analyzed. The experiments were performed in a 2 L reactor with a heating mantle and a condenser. The most important results were that the maximum leaching rate was obtained at pH 10.5, 0.10 mol/L NH₄OH, 45 °C, 4 h, 850 r/min and a liquid/solid ratio of 400:1, reaching a copper extraction rate of 82%. This result was related to the non-precipitation of copper in solution by the formation of copper tetra-amine. The liquid/solid ratio and stirring speed were essential for increasing the cuprite leaching. The maximum leaching rate was achieved at higher temperatures; however, significant copper leaching rate occurred at temperatures near the freezing point of water (17.9% over 4 h). Increasing NH₄OH concentration and decreasing particle size increased the cuprite leaching rate. The two ammoniacal reagents (NH₄F and (NH₄)₂SO₄) had low extraction rate of copper compared with NH₄OH. The kinetic model representing cuprite leaching was a chemical reaction on the surface. The order of the reaction with respect to the NH₄OH concentration was 1.8, and it was inversely proportional to the radius of the ore particles. The calculated activation energy was 44.36 kJ/mol in the temperature range of 10–45 °C.     

Leaching performance of Al-bearing spent LiFePO4 cathode powder in H2SO4 aqueous solution

The leaching performance and leaching kinetics of LiFePO₄ (LFP) and Al in Al-bearing spent LFP cathode powder were systematically studied. The effects of temperature (273?368 K), stirring speed (200?950 r/min), reaction time (0?240 min), acid-to-material ratio (0.1:1?1:1 mL/g) and liquid-to-solid ratio (3:1?9:1 mL/g) on the leaching process were investigated. The results show that the concentration of reactants and the temperature have a greater impact on the leaching of Al. Under the optimal conditions, leaching efficiencies of LFP and Al are 91.53% and 15.98%, respectively. The kinetic study shows that the leaching of LFP is kinetically controlled by mixed surface reaction and diffusion, with an activation energy of 22.990 kJ/mol; whereas the leaching of Al is only controlled by surface chemical reaction, with an activation energy of 46.581 kJ/mol. A low leaching temperature can effectively suppress the dissolving of Al during the acid leaching of the spent LFP cathode material.     

Leaching of chromite ore in concentrated KOH by catalytic oxidation using CuO as catalyst

CuO was used as a catalyst in the concentrated KOH solution to enhance the leaching of chromium from the chromite ore. The impacts of temperature, KOH-to-chromite ore mass ratio, CuO-to-chromite ore mass ratio, and gas flow rate on the chromium leaching rate were investigated. The results indicated that CuO played an important role in improving the chromium leaching rate. The leaching rate reached 98% after leaching for 6 h when CuO was applied, whereas it was only 60.8% without CuO under the same reaction conditions: temperature 230 °C, KOH-to-ore mass ratio 6:1, stirring speed 700 r/min, gas flow rate 1 L/min. According to the kinetics study, the catalytic oxidation was controlled by surface chemical reaction and the activation energy was calculated to be 15.79 kJ/mol when the temperature was above 230 °C. In contrast, without CuO, the rate-determining step was external diffusion and the apparent activation energy was 38.01 kJ/mol.     

硝酸溶液中氧化铜矿浸出工艺参数优化及动力学模拟

研究Tunceli孔雀石矿物在硝酸溶液中的溶出行为,以评估各种实验参数的影响.研究为分两个阶段.在第一步中,确定浸出过程的最佳条件,而在第二步中,对该过程进行动力学评估.在优化实验中,以硝酸浓度、温度、搅拌速度和固液比为自变量,采用中心组合设计法(CCD)获得实验数据.确定硝酸浓度、温度、固液比和搅拌速度的最佳值分别为...     

焙烧氟碳铈矿硫酸浸出稀土的动力学(英文)

研究了硫酸浸出德昌稀土与天青石共伴生矿的焙烧矿过程。考查粒度、搅拌速度、硫酸浓度和温度对稀土浸出率的影响,并对稀土的浸出动力学进行分析。在选定的浸出条件下:粒径0.074～0.100mm、硫酸浓度1.5mol/L、液固比8:1、搅拌速度500r/min,稀土浸出反应受内扩散控制,表观活化能为9.977kJ/mol。     

Complexation extraction of scheelite and transformation behaviour of tungsten-containing phase using H2SO4 solution with H2C2O4 as complexing agent

The extraction of tungsten from scheelite was carried out using a sulfuric acid solution with oxalic acid as the chelating agent. Tungsten was obtained in the form of highly soluble hydrogen aqua oxalato tungstate (H₂[WO₃(C₂O₄)·H₂O]) during the leaching process, while calcium remained in the residue as calcium sulfate dihydrate (CaSO₄·2H₂O). About 99.2% of the tungsten was leached at 70 °C, 1.5 mol/L sulfuric acid, 1 mol/L oxalic acid, a liquid/solid ratio of 25:1 (mL/g), an oxalic acid to sulfuric acid molar ratio of 1:1, a stirring speed of 300 r/min and a leaching time of 2 h. H₂[WO₃(C₂O₄)·H₂O] was thermally decomposed into tungstic acid (H₂WO₄), and tungsten trioxide (WO₃) was directly produced by calcining H₂WO₄ at 700 °C for 2 h. The surface chemical reaction was determined to be the controlling step during tungsten leaching, and the apparent activation energy was calculated to be 51.43 kJ/mol.     

采用Na2SO3溶液从硒渣中选择性浸出Se及其动力学

采用SO2还原沉金后液制得硒渣,再用Na2SO3选择性浸出硒渣,使Se得到有效分离;通过研究浸出过程中Se浸出率随时间的变化,建立该反应的动力学方程,确定Na2SO3溶液浓度、液固比、搅拌速度及反应温度对Se浸出率的影响,并计算相应的表观活化能。结果表明:增加Na2SO3溶液浓度和升高反应温度可以较大幅度提高Se的浸出率,液固比和搅拌速度对浸出Se的影响较小;Na2SO3浸出Se过程为Avrami模型混合控制,其特征参数n和表观活化能E分别为0.235和20.847 kJ/mol,Se的浸出率受反应温度的影响较大。     

Reaction mechanism of roasting Zn2SiO4 using NaOH

The reaction kinetics of roasting zinc silicate using NaOH was investigated. The orthogonal test was employed to optimize the reaction conditions and the optimized reaction conditions were as follows: molar ratio of NaOH to Zn₂SiO₄ of 16:1, reaction temperature of 550 °C, and reaction time of 2.5 h. In order to ascertain the phases transformation and reaction processes of zinc oxide and silica, the XRD phase analysis was used to analyze the phases of these specimens roasted at different temperatures. The final phases of the specimen roasted at 600 °C were Na₂ZnO₂, Na₄SiO₄, Na₂ZnSiO₄ and NaOH. The reaction kinetic equation of roasting was determined by the shrinking unreacted core model. Aiming to investigate the reaction mechanism, two control models of reaction rate were applied: chemical reaction at the particle surface and diffusion through the product layer. The results indicated that the diffusion through the product layer model described the reaction process well. The apparent activation energy of the roasting was 19.77 kJ/mol.     

Optimization of operating parameters and kinetics for chloride leaching of lead from melting furnace slag

The feasibility and kinetics of lead recovery from the slag of traditional lead melting furnace using chloride leaching were investigated. The effects of operating parameters such as leaching time, NaCl concentration, FeCl₃ concentration, liquid/solid ratio, stirring rate, temperature, and particle size on recovery of lead were studied and the optimization was done through the response surface methodology (RSM) based on central composite design (CCD) model. The optimum conditions were achieved as follows: leaching time 60 min, 80 °C, stirring rate 800 r/min, NaCl concentration 200 g/L, FeCl₃ concentration 80 g/L, liquid/solid ratio 16, and particle size less than 106 μm. More than 96% of lead was effectively recovered in optimum condition. Based on analysis of variance, the reaction temperature, liquid/solid ratio, and NaCl concentration were determined as the most effective parameters on leaching process, respectively. Kinetics study revealed that chloride leaching of galena is a first-order reaction and the diffusion through solid reaction product and chemical reaction control the mechanism. The activation energy of chloride leaching of galena was determined using Arrhenius model as 27.9 kJ/mol.     

Process optimization and kinetics for leaching of cerium,lanthanum and neodymium elements from iron ore waste's apatite by nitric acid

The leaching of rare earth elements (REEs) including cerium, lanthanum and neodymium from apatite concentrate obtained from iron ore wastes by nitric acid was studied. The effects of nitric acid concentration, solid to liquid ratio and leaching time on the recoveries of Ce, La and Nd were investigated using response surface methodology. The results showed that the acid concentration and solid to liquid ratio have significant effect on the leaching recoveries while the time has a little effect. The maximum REE leaching recoveries of 66.1%, 56.8% and 51.7% for Ce, La and Nd, respectively were achieved at the optimum leaching condition with 18% nitric acid concentration, 0.06 solid to liquid ratio and 38 min leaching time. The kinetics of cerium leaching was investigated using shrinking core model. It was observed that the leaching is composed of two stages. In the first stage a sharp increase in cerium leaching recovery was observed and at the longer time the leaching became slower. It was found that in the first stage the diffusion of reactants from ash layer is the rate controlling mechanism with an apparent activation energy of 6.54 kJ/mol, while in the second stage the mass transfer in the solution is the controlling mechanism.     

Dechromization and dealumination kinetics in process of Na2CO3-roasting pretreatment of laterite ores

A novel process was proposed for the activation pretreatment of limonitic laterite ores by Na₂CO₃ roasting. Dechromization and dealumination kinetics of the laterite ores and the effect of particle size, Na₂CO₃-ore mass ratio, and roasting temperature on Cr and Al extraction were studied. Experimental results indicate that the extraction rates of Cr and Al are up to 99% and 82%, respectively, under the optimal particle size of 44–74 μm, Na₂CO₃-to-ore mass ratio of 0.6:1, and temperature of 1000 °C. Dechromization within the range of 600–800 °C is controlled by the diffusion through the product layer with an apparent activation energy of 3.9 kJ/mol, and that it is controlled by the chemical reaction at the surface within the range of 900–1100 °C with an apparent activation energy of 54.3 kJ/mol. Besides, the Avrami diffusion controlled model with on apparent activation energy of 16.4 kJ/mol is most applicable for dealumination. Furthermore, 96.8% Ni and 95.6% Co could be extracted from the alkali-roasting residues in the subsequent pressure acid leaching process.     

Leaching process of germanium oxide dust

A thermodynamic analysis on the acid leaching process of germanium oxide dust and discussion on the behaviors of main substances of the dust in the leaching process were carried out. The effects of temperature, acid concentration, leaching time and stirring speed on the leaching rate of germanium were investigated. Based on the characteristic of the dust, the kinetics and reactive mechanism of acid leaching were stud- ied. The results show that the leaching of the dust by acid belonged to "the unreacted core shrinking model" of producing solid outgrowth layer. The chemical reaction was controlled by inner diffusion process. The apparent activation energy of leaching process was 12.60 kJ/mol. The leaching reaction of germanium was determined to be mainly second order reaction. The optimum conditions were the reaction tem- perature of 363 K, the leaching time of 2.5 h, the stirring speed of 120 r/min, the solid-to-liquid ratio of 1/8 and the acid concentration of 120 g/L. Under these conditions, the leaching rate of germanium can come up to more than 87%.     

Leaching kinetics of antimony-bearing complex sulfides ore in hydrochloric acid solution with ozone

The leaching kinetics of Sb and Fe from antimony-bearing complex sulfides ore was investigated in HCl solution by oxidation–leaching with ozone. The effects of temperature, HCl concentration, stirring speed and particle size on the process were explored. It is found that the recoveries of Sb and Fe reach 86.1% and 28.8%, respectively, when the reaction conditions are 4.0 mol/L HCl, 900 r/min stirring speed at 85 °C with <0.074 mm particle size after 50 min leaching. XRD analysis indicates that no new solid product forms in the leaching residue and the leaching process can be described by shrinking core model. The leaching of Sb corresponds to diffusion-controlled model at low temperature (15–45 °C) and mixed-controlled model at high temperature (45–85 °C), and the apparent activation energies are 6.91 and 17.93 kJ/mol, respectively. The leaching of Fe corresponds to diffusion-controlled model, and the apparent activation energy is 1.99 kJ/mol. Three semi-empirical rate equations are obtained to describe the leaching process.     

用碱性Na_2EDTA溶液从次氧化锌烟灰中选择性回收铅(英文)

采用碱性Na2EDTA溶液从次氧化锌烟灰中回收铅。探讨温度、浸出时间、Na2EDTA浓度和起始NaOH浓度对铅、锌浸出率的影响。得到最优实验条件如下:液固比5:1 mL/g、搅拌速度650 r/min、Na2EDTA浓度0.12mol/L、NaOH初始浓度0.5 mol/L、温度70°C、浸出时间120 min。在最优实验条件下,铅、锌、氟和氯的平均浸出率分别为89.92%、0.94%、62.84%和90.02%。浸出液用于电沉积铅粉。在温度为60°C、电流密度为200A/m2、H3PO4浓度为1.5 g/L、铅离子浓度不低于5 g/L时,电沉积铅粉平均电流效率大约为93%,阴极铅纯度高于98%。电沉积1 kg铅粉大约消耗0.218 kg Na2EDTA和0.958 kW·h电能。     

Cyclic metallurgical process for extracting V and Cr from vanadium slag: Part II. Separation and recovery of Cr from vanadium precipitated solution

Cyclic metallurgical process for separation and recovery of Cr from vanadium precipitated solution by precipitation with PbCO₃ and leaching with Na₂CO₃ was investigated. The concentration of Cr residue in the solution decreases from 2.360 to 0.001 g/L by adding PbCO₃ into vanadium precipitated solution according to Pb/Cr molar ratio of 2.5, adjusting the pH to 3.0 and stirring for 180 min at 30 °C. Then, the precipitates were leached with hot Na₂CO₃ solution to obtain leaching solution containing Na₂CrO₄ and leaching residue containing PbCO₃. The leaching efficiency of Cr reaches 96.43% by adding the precipitates into 0.5 mol/L Na₂CO₃ solution with the mass ratio of liquid to solid (L/S) of 10:1 mL/g and stirring for 60 min under pH 9.5 at 70 °C. After filtration, leaching residue is reused in Cr precipitation and leaching solution is used to circularly leach the Cr precipitates until Na₂CrO₄ approaches the saturation. Finally, the product of Na₂CrO₄·4H₂O is obtained by evaporation and crystallization of leaching solution.     

氯化浸取钢渣中钒的动力学研究

为了提高湿法浸出低钒钢渣中钒的浸出率,并为湿法浸出低钒钢渣中钒提供理论依据,从动力学角度分析整个浸出过程,并考察温度、液/固比、浸出时间和搅拌速度对浸出过程的影响。结果表明,在90℃,液/固比为10:1以及4.0mol/L盐酸,过氧化氢8.0mL,浸取90min条件下,低钒钢渣中钒的浸出率可达到98.8%。通过正交实验和动力学推导,得到描述浸出过程的经验方程。低钒钢渣湿法浸出钒的动力学模型为未反应收缩核模型,浸出过程的表观活化能为7.21kJ/mol。该模型表明浸出过程中的控制步骤取决于边界层的扩散速度。提高温度、液/固比和浸出时间,均可增加钒的浸出速度,提高钒的浸出率。