Reductive leaching of low-grade manganese ore with pre-processed cornstalk

Cornstalk is usually directly used as a reductant in reductive leaching manganese. However, low utilization of cornstalk makes low manganese dissolution ratio. In the research, pretreatment for cornstalk was proposed to improve manganese dissolution ratio. Cornstalk was preprocessed by a heated sulfuric acid solution (1.2 M of sulfuric acid concentration) for 10 min at 80°C. Thereafter, both the pretreated solu-tion and the residue were used as a reductant for manganese leaching. This method not only exhibited superior activity for hydrolyzing corn-stalk but also enhanced manganese dissolution. These effects were attributed to an increase in the amount of reductive sugars resulting from lignin hydrolysis. Through acid pretreatment for cornstalk, the manganese dissolution ratio was improved from 50.14%to 83.46%. The pre-sent work demonstrates for the first time the effective acid pretreatment of cornstalk to provide a cost-effective reductant for manganese leaching.     

Preparation of manganese sulfate from low-grade manganese carbonate ores by sulfuric acid leaching

In this study, a method for preparing pure manganese sulfate from low-grade ores with a granule mean size of 0.47 mm by direct acid leaching was developed. The effects of the types of leaching agents, sulfuric acid concentration, reaction temperature, and agitation rate on the leaching efficiency of manganese were investigated. We observed that sulfuric acid used as a leaching agent provides a similar leaching efficiency of manganese and superior selectivity against calcium compared to hydrochloric acid. The optimal leaching conditions in sulfuric acid media were determined; under the optimal conditions, the leaching efficiencies of Mn and Ca were 92.42% and 9.61%, respectively. Moreover, the kinetics of manganese leaching indicated that the leaching follows the diffusion-controlled model with an apparent activation energy of 12.28 kJ·mol-1. The purification conditions of the leaching solution were also discussed. The results show that manganese dioxide is a suitable oxidant of ferrous ions and sodium dimethyldithiocarbamate is an effective precipitant of heavy metals. Finally, through chemical analysis and X-ray diffraction analysis, the obtained product was determined to contain 98% of MnSO₄·H₂O.     

芦丁还原浸出低品位软锰矿的研究

为了研究糖蜜酒精废液中有机物还原浸出软锰矿的机理,采用芦丁为还原剂,在酸性介质中直接浸出低品位软锰矿。通过正交实验和单因素实验,研究了芦丁浓度、硫酸浓度、反应温度、浸出时间等因素对锰浸出率的影响,并对反应过程机理进行了初步探讨。实验结果表明,影响锰浸出率的主要因素依次为反应温度、硫酸浓度、浸出时间和芦丁浓度;当硫酸初始浓度2.35 mol/L,芦丁初始浓度0.041 mol/L,反应温度90℃,浸出时间90 m in时,锰浸出率达94.9%。     

硫酸中香蕉皮还原低品位锰矿的浸出行为及动力学模型

Manganese was leached from a low-grade manganese ore (LGMO) using banana peel as the reductant in a dilute sulfuric acid medium. The effects of banana peel amount, H₂SO₄ concentration, reaction temperature, and time on Mn leaching from the complex LGMO were studied. A leaching efficiency of ~98% was achieved at a leaching time of 2 h, banana peel amount of 4 g, leaching temperature of 120°C, manganese ore amount of 5 g, and sulfuric acid concentration of 15vol%. The phase, microstructural, and chemical analyses of LGMO samples before and after the leaching process confirmed the successful leaching of manganese. Furthermore, the leaching process followed the shrinking core model and the leaching rate was controlled by a surface chemical reaction (1 ? (1 ? x)1/3 = kt) mechanism with an apparent activation energy of 40.19 kJ·mol?1.     

Optimization of reaction conditions for the electroleaching of manganese from low-grade pyrolusite

In the present study, a response surface methodology was used to optimize the electroleaching of Mn from low-grade pyrolusite. Ferrous sulfate heptahydrate was used in this reaction as a reducing agent in sulfuric acid solutions. The effect of six process variables, including the mass ratio of ferrous sulfate heptahydrate to pyrolusite, mass ratio of sulfuric acid to pyrolusite, liquid-to-solid ratio, current density, leaching temperature, and leaching time, as well as their binary interactions, were modeled. The results revealed that the order of these factors with respect to their effects on the leaching efficiency were mass ratio of ferrous sulfate heptahydrate to pyrolusite > leaching time > mass ratio of sulfuric acid to pyrolusite > liquid-to-solid ratio > leaching temperature > current density. The optimum conditions were as follows: 1.10:1 mass ratio of ferrous sulfate heptahydrate to pyrolusite, 0.9:1 mass ratio of sulfuric acid to pyrolusite, liquid-to-solid ratio of 0.7:1, current density of 947 A/m2, leaching time of 180 min, and leaching temperature of 73℃. Under these conditions, the predicted leaching efficiency for Mn was 94.1%; the obtained experimental result was 95.7%, which confirmed the validity of the model.     

Reductive recovery of manganese from low-grade manganese dioxide ore using toxic nitrocellulose acid wastewater as reductant

The hydrometallurgical strategy of extracting Mn from low-grade Mn ores has attracted attention for the production of electrolytic manganese metal (EMM). In this work, the reductive dissolution of low-grade MnO₂ ores using toxic nitrocellulose acidic wastewater (NAW) as a reductant was investigated for the first time. Under the optimized conditions of an MnO₂ ore dosage of 100 g·L-1, an ore particle size of -200 mesh, concentrated H₂SO₄-to-NAW volume ratio of 0.12, reaction temperature of 90℃, stirring speed at 160 r·min-1, and a contact time of 120 min, the reductive leaching efficiency of Mn and the total organic carbon (TOC) removal efficiency of NAW reached 97.4% and 98.5%, respectively. The residual TOC of 31.6 mg·L-1 did not adversely affect the preparation of EMM. The current process offers a feasible route for the concurrent realization of the reductive leaching of Mn and the treatment of toxic wastewater via a simple one-step process.     

玉米秸秆和稻草-硫酸还原浸出低品位软锰矿

研究了以低品位软锰矿为原料,用玉米秸秆和稻草为混合还原剂,在硫酸介质中还原浸出锰的效果.通过单因素实验确定了玉米秸秆和稻草-硫酸还原浸锰的适宜工艺条件,该适宜工艺条件如下:软锰矿粒径为109μm,玉米秸秆和稻草的质量比为1∶1.5,玉米秸秆和稻草还原剂与软锰矿的质量比为3∶10,硫酸浓度为3.0 mol/L,液固比为10 m L/g,反应温度为90℃,反应时间为80 min.在适宜的浸出工艺条件下,软锰矿中锰的浸出率可达97%以上,锰的浸出效果较好.     

用草酸在硫酸溶液中从中性浸出渣中还原浸出铟

The present study evaluates the reductive leaching of indium from indium-bearing zinc ferrite using oxalic acid as a reducer in sulfuric acid solution. The effect of main factors affecting the process rate, including the oxalic-acid-to-sulfuric-acid ratio, stirring rate, grain size, temperature, and the initial concentration of synergic acid, was precisely evaluated. The results confirmed the acceptable efficiency of dissolving indium in the presence of oxalic acid. The shrinking-core model with a chemical-reaction-controlled step can correctly describe the kinetics of indium dissolution. On the basis of an apparent activation energy of 44.55 kJ/mol and a reaction order with respect to the acid concentration of 1.14, the presence of oxalic acid was found to reduce the sensitivity to temperature changes and to increase the effect of changes in acid concentration. Finally, the equation of the kinetic model based on the factors under study is presented.     

响应面法优化葡萄糖还原浸出半氧化锰矿的研究

采用响应面法优化葡萄糖还原浸出半氧化锰矿的工艺,以锰浸出率为响应值,对硫酸用量/矿量、葡萄糖用量/矿量、氟化铵用量/矿量和浸取时间四因素进行建模。结果表明,这些因素与锰浸出率的关系符合二次模型,各个因素对锰浸出率影响均极其显著,各个因素的二次方以及硫酸用量/矿量与葡萄糖用量/矿量的交互作用对锰浸出率影响较显著。最优工艺条件为：硫酸用量/矿量0.880 g/g、葡萄糖用量/矿量0.059 g/g、氟化铵用量/矿量0.04 g/g、浸取时间6 h。在此条件下,锰浸出率为93.09%,与模型预测值93.18%无显著差异,而相同条件下未加氟化铵的锰浸出率为82.37%。     

Hydrometallurgical leaching and kinetic modeling of low-grade manganese ore with banana peel in sulfuric acid

Manganese was leached from a low-grade manganese ore(LGMO) using banana peel as the reductant in a dilute sulfuric acid medium. The effects of banana peel amount, H_2SO_4 concentration, reaction temperature, and time on Mn leaching from the complex LGMO were studied. A leaching efficiency of ～98% was achieved at a leaching time of 2 h, banana peel amount of 4 g, leaching temperature of 120°C,manganese ore amount of 5 g, and sulfuric acid concentration of 15 vol%. The phase, microstructural, and chemical analyses of LGMO samples before and after the leaching process confirmed the successful leaching of manganese. Furthermore, the leaching process followed the shrinking core model and the leaching rate was controlled by a surface chemical reaction(1-(1-x)~(1/3)=kt) mechanism with an apparent activation energy of 40.19 k J·mol~(-1).     

Effect of desliming on the magnetic separation of low-grade ferruginous manganese ore

In the present investigation, magnetic separation studies using an induced roll magnetic separator were conducted to beneficiate low-grade ferruginous manganese ore. The feed ore was assayed to contain 22.4% Mn and 35.9% SiO₂, with a manganese-to-iron mass ratio (Mn:Fe ratio) of 1.6. This ore was characterized in detail using different techniques, including quantitative evaluation of minerals by scanning electron microscopy, which revealed that the ore is extremely siliceous in nature and that the associated gangue minerals are more or less evenly distributed in almost all of the size fractions in major proportion. Magnetic separation studies were conducted on both the as-received ore fines and the classified fines to enrich their manganese content and Mn:Fe ratio. The results indicated that the efficiency of separation for deslimed fines was better than that for the treated unclassified bulk sample. On the basis of these results, we proposed a process flow sheet for the beneficiation of low-grade manganese ore fines using a Floatex density separator as a pre-concentrator followed by two-stage magnetic separation. The overall recovery of manganese in the final product from the proposed flow sheet is 44.7% with an assay value of 45.8% and the Mn:Fe ratio of 3.1.     

A primary study of simultaneous leaching of silver containing manganese ore and a sphalerite concentrate

It is demonstrated that a low grade, silver containing manganese deposit can be exploited by reaction with a ZnS concentrate. In the process, Mn（IV） is reduced to Mn（II） and the ZnS converted to ZnSO4. In the process, Mn and Zn are solubilized. By studying the effects of quantity of reducing agent and sulfuric acid added and of time, temperature and liquid-to-solid ratio on the leaching process, the optimum technological conditions have been achieved un- der which the amount of manganese leached was 98 %. The silver can be removed from remaining solids by adding ammonia liquor. Then by adding hydrazine hydrate, sponge silver with purity up to 99 % can be obtained with a recovery of 87 %.     

Reductive leaching of indium from the neutral leaching residue using oxalic acid in sulfuric acid solution

The present study evaluates the reductive leaching of indium from indium-bearing zinc ferrite using oxalic acid as a reducer in sulfuric acid solution. The effect of main factors affecting the process rate, including the oxalic-acid-to-sulfuric-acid ratio, stirring rate, grain size,temperature, and the initial concentration of synergic acid, was precisely evaluated. The results confirmed the acceptable efficiency of dissolving indium in the presence of oxalic acid. The shrinking-core model with a chemical-reaction-controlled step can correctly describe the kinetics of indium dissolution. On the basis of an apparent activation energy of 44.55 k J/mol and a reaction order with respect to the acid concentration of 1.14, the presence of oxalic acid was found to reduce the sensitivity to temperature changes and to increase the effect of changes in acid concentration. Finally, the equation of the kinetic model based on the factors under study is presented.     

改善原地溶浸工艺中矿石破碎质量的力学机理

通过改善矿石的破碎质量,提高矿堆的渗透性,使原地溶浸法开采稳固性和渗透性极差的低品位铜镍矿石成为可能.依据矿石中缺陷存在的情况,应用断裂动力学和损伤力学的观点,着重分析了爆炸应力波能量加载对矿石破裂、损伤行为产生的影响;对矿石块体中贮存爆破动能加载,致使块体产生破裂或损伤的行为进行了探讨.此外,对浸蚀裂解与作用在爆堆上的地压能在破裂或损伤过程中所起的辅助作用进行了讨论.给出了应用中深孔挤压爆破技术,原地崩落低品位铜镍矿石,形成溶浸矿堆的一些相应相关技术参数.分析结果表明采用合理的爆破技术,可以改善应力波和爆破动能的加载条件,能提高矿石块体的破碎质量及累积损伤程度;借助于浸蚀裂解与地压能的作用,可进一步改善堆浸矿石的块体分布,提高溶浸液与有用组分的接触和被浸出的机会.     

碱性氧化铜矿产氨菌浸出特性

为了研究碱性氧化铜矿产氨菌浸出特性,分析了产氨菌浸矿过程对矿石的作用.将云南某矿的碱性氧化铜矿置于含菌培养液、去菌培养液和氨水等5种浸矿溶液中,在同一条件下进行摇瓶浸矿实验,剖析溶液中各可能的浸矿因子.研究结果表明:产氨菌产氨能力较强,尿素培养液中氨质量浓度最大达8.93 g·L-1;产氨量与细菌含量呈正相关关系,细菌含量越高,产氨量越大;产氨菌主要通过产氨间接浸矿,此外产氨菌和其代谢产物都能直接作用于矿石,浸矿能力细菌产氨>细菌>细菌代谢产物,三者比值约为12:5:4.     

硫氧混合铅锌矿中锌的氧化氨浸动力学

常压低温条件下在NH3-(NH4)2 SO4体系中使用过硫酸铵作为氧化剂对硫氧混合铅锌矿中的锌进行浸出实验,系统研究了搅拌速度、浸出剂浓度、氧化剂浓度与温度对于锌浸出率的影响.结果表明,在最优条件下锌的浸出率可达93.2%,且浸出过程中几乎没有其他离子进入溶液,实现了锌的选择性高效浸出,从而简化了后续的浸出液净化与材料制备过程.动力学研究表明,硫氧混合铅锌矿中锌的氧化氨浸过程遵循固体产物层扩散控制的未反应核收缩模型,浸出反应的表观活化能为17.89 kJ·mol-1.     

铁闪锌矿在硫酸浸出过程中的电化学行为

为研究铁闪锌矿氧化浸出机理,以人工合成铁闪锌矿的悬浮矿浆为研究对象,采用三电极体系对其在硫酸浸出过程中的电化学行为进行了研究.结果表明:在铁闪锌矿浸出过程中有H2S中间产物生成;通氧、元素硫分散剂以及氯离子引入浸出体系均能明显促进铁闪锌矿的电化学氧化;随硫酸浓度升高,铁闪锌矿的电化学氧化不断加强而闪锌矿的电化学氧化受抑.与闪锌矿相比,铁闪锌矿更易发生电化学氧化.在铁闪锌矿硫酸浸出过程中,其电化学氧化作用不容忽视.     

蛇纹石硫酸浸出过程工艺条件的优化研究

硫酸浸出过程是蛇纹石综合利用过程中的关键因素之一 ,在单因素条件实验的基础上 ,采用 L16( 4 4)正交实验法对蛇纹石硫酸浸出过程的优化工艺条件进行了研究。结果表明 :当以矿粒粒径 <75μm、搅拌速率 >2 0 0 r/min、活化剂 FS-1加入量为 1.5 %为固定因素时 ,其优化工艺条件为 :浸出温度 10 0℃ ,浸出时间 3h,硫酸质量浓度为 45 % ,液固比 4.0。     

电解锰渣固结体中重金属浸出毒性及其在模拟酸雨下的淋溶特性分析

以一种硅铝基胶凝材料为固结剂,对某电解锰厂的锰渣进行固化处理,分析在不同养护周期下锰渣固结体中重金属的浸出毒性,通过动态淋溶试验,研究养护周期和模拟酸雨pH值对固结体淋溶特性的影响,并通过SEM和XRD手段,分析酸雨淋溶前后固结体的形貌和成分,探讨固结体淋溶过程中重金属的溶出机理。结果表明,固化处理可有效固定锰渣中的重金属,固结体的养护周期对其在酸雨淋溶过程中Mn的溶出影响较大,其中养护28天的锰渣固结体在不同pH值酸雨淋溶条件下,淋溶液中Mn和Zn的最高浓度分别为0.189、0.116mg·L~(-1),其它重金属含量也较低,均低于国家标准浓度限值;酸雨淋溶过程中,固结体的晶体结构没有受到破坏,大部分的重金属被包裹在晶体内部或以更稳定的化合态存在,不易被溶出。     

Sulfuric acid leaching kinetics of South African chromite

The sulfuric acid leaching kinetics of South African chromite was investigated. The negative influence of a solid product layer constituted of a silicon-rich phase and chromium-rich sulfate was eliminated by crushing the chromite and by selecting proper leaching conditions. The dimensionless change in specific surface area and the conversion rate of the chromite were observed to exhibit a proportional relationship. A modified shrinking particle model was developed to account for the change in reactive surface area, and the model was fitted to experimental data. The resulting model was observed to describe experimental findings very well. Kinetics analysis revealed that the leaching process is controlled by a chemical reaction under the employed experimental conditions and the activation energy of the reaction is 48 kJ·mol-1.