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).     

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.     

Reductive leaching of manganese from low-grade pyrolusite ore in sulfuric acid using pyrolysis-pretreated sawdust as a reductant

Manganese (Mn) leaching and recovery from low-grade pyrolusite ore were studied using sulfuric acid (H₂SO₄) as a leachant and pyrolysis-pretreated sawdust as a reductant. The effects of the dosage of pyrolysis-pretreated sawdust to pyrolusite ore, the concentration of sulfuric acid, the liquid/solid ratio, the leaching temperature, and the leaching time on manganese and iron leaching efficiencies were investigated. Analysis of manganese and iron leaching efficiencies revealed that a high manganese leaching efficiency was achieved with low iron extraction. The optimal leaching efficiency was determined to be 20wt% pyrolysis-pretreated sawdust and 3.0 mol/L H₂SO₄ using a liquid/ solid ratio of 6.0 mL/g for 90 min at 90℃. Other low-grade pyrolusite ores were tested, and the results showed that they responded well with manganese leaching efficiencies greater than 98%.     

Reductive atmospheric acid leaching of spent alkaline batteries in H_2SO_4/Na_2SO_3 solutions

This work studies the optimum reductive leaching process for manganese and zinc recovery from spent alkaline battery paste. The effects of reducing agents, acid concentration, pulp density, reaction temperature, and leaching time on the dissolution of manganese and zinc were investigated in detail. Manganese dissolution by reductive acidic media is an intermediate-controlled process with an activation energy of 12.28 kJ ·mol-1. After being leached, manganese and zinc were selectively precipitated with sodium hydroxide. The zinc was entirely converted into zincate(Zn(OH)2-4) ions and thus did not co-precipitate with manganese hydroxide during this treatment(2.0 M NaO H, 90 min, 200 r/min, pH 13). After the manganese was removed from the solution, the Zn(OH)2- 4 was precipitated as zinc sulfate in the presence of sulfuric acid. The results indicated that this process could be effective in recovering manganese and zinc from alkaline batteries.     

Kinetics of the removal of magnesium from phosphate ore with diluted acid

The kinetics of magnesium removal from phosphate ore with diluted sulfuric acid was studied. The experimental results showed that hydrogen ion concentration, temperature, reaction time and particle size distributions of phosphate ore have different effect on the phosphorus removal and magnesium removal. A kinetic database of magnesium and phosphorus removals from ore was set up. The three-dimensional graphs of phosphorus and magnesium removals versus time and reaction conditions were drawn by a computer. The kinetic equations were developed.     

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.     

Hydrometallurgical application for treating a Nigerian chalcopyrite ore in chloride medium: Part I. Dissolution kinetics assessment

The dissolution kinetics of a Nigerian chalcopyrite ore in hydrochloric acid was studied in this article. Acid concentration, reaction temperature, and ore particle size were chosen as experimental parameters. The chemical and morphological studies of the ore before and after leaching at optimal conditions were carried out by X-ray diffraction (XRD) and scanning electron microscopy (SEM). It is revealed that increasing the acid concentration and system temperature and decreasing the ore particle size greatly enhances the dissolution rate. The dissolution kinetics was found to follow the shrinking core model for the diffusion control mechanism where the activation energy (E_a) of 32.92 kJ·mol?1 was obtained for the process and supported by morphological changes at a higher dissolution of 91.33%.     

Extracting copper from copper oxide ore by a zwitterionic reagent and dissolution kinetics

Sulfamic acid (SA), which possesses a zwitterionic structure, was applied as a leaching reagent for the first time for extracting copper from copper oxide ore. The effects of reaction time, temperature, particle size, reagent concentration, and stirring speed on this leaching were studied. The dissolution kinetics of malachite was illustrated with a three-dimensional diffusion model. A novel leaching effect of SA on malachite was eventually demonstrated. The leaching rate increased with decreasing particle size and increasing concentration, reaction temperature and stirring speed. The activation energy for SA leaching malachite was 33.23 kJ/mol. Furthermore, the effectiveness of SA as a new reagent for extracting copper from copper oxide ore was confirmed by experiment. This approach may provide a solution suitable for subsequent electrowinning. In addition, results reported herein may provide basic data that enable the leaching of other carbonate minerals of copper, zinc, cobalt and so on in an SA system.     

Recovery of valuable metals from a low-grade nickel ore using an ammonium sulfate roasting-leaching process

Metal leaching from a low-grade nickel ore was investigated using an ammonium sulfate roasting-water leaching process. The nickel ore was mixed with ammonium sulfate, followed by roasting and finally leaching with water. During the process the effects of the amount of ammonium sulfate, roasting temperature, and roasting time on the leaching recovery of metal elements were analyzed. The optimum technological parameters were determined as follows:ammonium sulfate/ore ratio, 0.8 g/g; roasting temperature, 400℃; and roasting time, 2 h. Under the optimum condition the leaching recoveries of Ni, Cu, Fe, and Mg were 83.48%, 76.24%, 56.43%, and 62.15%, respectively. Furthermore, the dissolution kinetics of Ni and Mg from the nickel ore was studied. The apparent activation energies for the leaching reaction of Ni and Mg were 18.782 and 10.038 kJ·mol-1, which were consistent with the values of diffusion control reactions. Therefore, the results demonstrated that the leaching recoveries of Ni and Mg were controlled by diffusion.     

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.     

玉米秸秆还原浸出高铁低品位锰矿研究

采用"玉米秸秆硫酸预处理—浸出"工艺处理含Mn 9.63%(质量分数)的朝阳锰矿.通过试验考察用硫酸对玉米秸秆进行预处理的时间、秸秆用量、温度、硫酸浓度和浸出温度对锰浸出率的影响.试验表明,在秸秆用量2.5 g、时间10 min、硫酸浓度1.2 mol/L、温度80℃时预处理秸秆,浸出温度为90℃的条件下,锰的浸出率达92%,杂质铁溶出率仅为20%.玉米秸秆处理前后FTIR分析结果表明,对秸秆进行硫酸预处理能够破坏难降解的木质素结构,有利于提高锰的浸出效果.     

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%。     

铀矿堆浸过程的动力学模型及实验研究

通过浸出率与浸出时间的关系曲线,推导出铀矿硫酸堆浸过程的控制模式和化学反应常数。实验结果表明浸出率在小于60％以内,与浸出时间呈直线关系,得出浸出过程控制模式为化学反应控制的结论。实验数据计算得出在该实验中化学反应浸出速率常数为1．5×10^-7s^-1,液固相化学反应速率系数为3．6×10^-8m／s。实验结果在未加氧化剂的酸堆浸铀矿石中,最多只有40％的四价铀能氧化成六价铀,要想进一步提高铀的浸出率,必须加入氧化剂,提高氧化能力。     

不同化学结合剂对铜氧化物的团聚和浸出行为

The chemical binder is one of the critical factors affecting ore agglomeration behavior and leaching efficiency. In this study, we investigated the effect of the type of binder and mass fraction of the H₂SO₄ solution used on the curing, soaking, and leaching behavior of agglomerations. The results revealed that Portland cement (3CaO·SiO₂, 2CaO·SiO₂, and 3CaO·Al₂O₃) was the optimal binder for obtaining a well-shaped, stable agglomeration structure. A higher extraction rate was achieved when using Portland cement than that obtained using sodium silicate, gypsum, or acid-proof cement. An excessive geometric mean size is not conducive to obtaining well-shaped agglomerations and desirable porosity. Using computed tomography (CT) and MATLAB, the porosity of two-dimensional CT images in sample concentrations L1–L3 was observed to increase at least 4.5vol% after acid leaching. Ore agglomerations began to be heavily destroyed and even to disintegrate when the sulfuric acid solution concentration was higher than 30 g/L, which was caused by the excessive accumulation of reaction products and residuals.     

斑铜矿在酸性溶液中的浸出动力学行为

以斑铜矿为研究对象,在H₂SO₄酸性体系中,以NaS₂O₈为氧化剂,详细考察浸出时间、温度、矿物尺寸、液固比、H₂SO₄浓度和NaS₂O₈浓度对铜浸出率的影响.浸出行为表明,斑铜矿浸出动力学行为符合固体膜层的界面传质和扩散的混合控制,表观反应活化能为33.97 kJ/mol,浸出动力学方程为（ln(1-x)）/3-1+(1-x)^-1/3=k_mt.     

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 vanadium 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 temperature 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 vanadium reaches 97.69%.     

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 %.