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%.     

Process development for the direct solvent extraction of nickel and cobalt from nitrate solution: aluminum,cobalt, and nickel separation using Cyanex 272

A direct solvent extraction (DSX) process for purifying nickel and cobalt from the nitric acid leach solution of nickel laterite ores was conceived and experimentally probed. The proposed process consists of two solvent extraction (SX) steps but with only one extractant-bis(2,4,4-trimethylpentyl)phosphinic acid (Cyanex® 272)-used in both steps. The first extraction step involved the removal of aluminum and zinc, whereas the second extraction step involved the separation of cobalt along with manganese from nickel. The experimental results showed essentially quantitative removal of aluminum (>97%) and zinc (>99%) in a single extraction stage using 20vol% Cyanex 272 at pH 2.1. Some cobalt (32%) and manganese (55%) were co-extracted but were easily scrubbed out completely from the loaded organic phase using dilute sulfuric acid at pH ≤ 1.38. Cobalt and manganese in the first extraction raffinate were extracted completely in four extraction stages at staggered pH values of 4.0, 4.4, 4.5, and 4.0 in the first, second, third, and fourth stages, respectively, using also 20vol% Cyanex 272. A small amount of nickel (up to 6.6%) was co-extracted but was easily scrubbed out completely with dilute sulfuric acid at pH 2.0. A flow diagram showing the input and output conditions and the metals separated under the deduced optimum conditions is presented.     

用硫酸氧化预处理从难处理金矿中提金的无害化工艺研究

探讨了硫酸作为氧化剂分解含砷难处理金矿的技术可行性,实验结果表明硫酸可以有效地氧化砷黄铁矿而实现金的单体解离,适宜的分解条件为：硫酸与矿粉重量比为３,采用机械搅拌,反应温度２４０～２６０℃,分解时间２～３ｈ,分解渣采用硫脲浸出时浸金率高于９５％浸金条件为：硫脲浓度１％,三价铁作氧化剂,氧化剂与络合剂比值为０．０４～０．０６,浸出时间６ｈ图１,表６,参８     

Simulation of solute transportation within porous particles during the bioleaching process

A mathematical model, accounting for the sulfuric acid and ferric ions diffusion and the copper sulfide mineral leaching process, was developed for an ore particle by considering its porous structure. It was simulated with the simulation tool COMSOL Multiphysics. The simulation results show that the highest acid and ferric concentrations near the particle surface are apparent, while the concentrations in the central particle increase slightly as the less-porous ore core with low permeability prevents the oxidation from penetrating. The extraction of the mineral near the particle surface is the maximum, mainly because of ample sulfuric acid, ferric ions, bacteria, and oxygen available for the leaching process. Because of low oxidation concentration in the central part of the particle, the reaction rate and copper sulphide conversion are small. The simulation shows good agreement with the experimental results.     

墨西哥某氧化铜矿形态学特征对其酸浸试验的影响

通过化学选矿的方法研究了氧化铜矿形态学特征对铜浸出率的影响.以硫酸为浸出剂对某氧化铜矿的两种矿样进行酸浸试验.分别研究浸出剂用量、浸出时间以及料层厚度等影响因素.同时,采用扫描电子显微镜（SEM）观察原矿和尾矿中铜的形态学特征,即A矿样中铜矿物以薄膜的形式覆盖在脉石矿物表面;B矿样中铜矿物呈颗粒状与脉石矿物共生.酸浸试验结果表明：在相同的条件下,A矿样的铜浸出率要远高于B矿样.这是因为A矿样中铜矿物具有更大的比表面积,与硫酸的接触面大,反应更加充分.     

赞比亚低品位难处理铜钴矿湿法冶金新工艺

以赞比亚某低品位难处理铜钴矿石为研究样本,采用全湿法冶金方法,开展新工艺研究.研究可知:矿石Cu,Co和S质量分数分别为1.270%,0.071%,0.022%.矿石中铜矿物主要为假孔雀石和少量的孔雀石.钴矿物主要为钴锰矿和水钴矿,在褐铁矿和黑云母晶体中有少量铜、钴,矿石中铜钴元素赋存状态极其复杂.最佳的浸出条件为粒度小于74μm的矿粒所占比例70%、浸出温度65℃、浸出时间4 h、矿浆质量分数30%、硫酸加入量55 kg·t-1.该条件下铜浸出率可达74.34%左右,钴浸出率可达43.32%左右.充分利用萃余液中的硫酸可降低酸耗,硫酸用量减少20%以上.在搅拌浸出过程中加入适量还原剂Na_2SO_3或FeSO_4,可将钴的浸出率从43%提高到78%.     

稀酸浸出还原焙烧红土矿时铁还原度对浸出的影响

对含镍的红土矿进行了微波加热还原焙烧-稀硫酸浸出的实验研究,考察了活性炭粉加入量、微波功率和加热时间对铁还原程度的影响,并分析了铁还原度对浸出过程中镍和铁浸出率的影响.结果表明红土矿中铁的还原程度随碳粉加入量、微波功率和加热时间的增加而增大,在800 W的微波功率下加热约12.5 min即可完成还原反应.镍的浸出率与铁的还原度近似呈线性关系,但铁的浸出率在铁还原度超过60%后增长迅速.因此铁还原度控制在60%为宜,相应的镍浸出率约为85%,而铁的浸出率不超过30%.     

废催化剂焙烧水浸渣中硫酸浸取钴的动力学研究

以废催化剂处理过程中得到的镍钴渣为研究对象,采用硫酸浸出镍钴渣,使钴和镍得到有效回收,并对硫酸浸出钴的动力学进行探讨。研究结果表明:搅拌速度为400～1200r/min时对钴浸出率的影响非常小,物料粒度、硫酸浓度和反应温度等因素对钴浸出率则有较大影响;当反应温度为80℃,反应时间为180min,原料粒度为(0.074～0.100)mm,H2SO4浓度为6mol/L,搅拌速度为800r/min,固液比为1:10时,钴的浸出率为94.2%,镍的浸出率则为93.5%;硫酸浸出镍钴渣的反应受产物层内扩散控制,表观活化能为16.34kJ/mol。     

A mathematical model for column leaching of ion adsorption-type rare earth ores

Column leaching experiments with ion adsorption-type rare earth ores for different lixiviant concentrations and different column heights were carried out. A mathematical model of column leaching was constructed based on the experimental data. Two parameters(a and b)in the model were determined according to the following methodology: the ore column was divided into several units; each unit was treated with multiple leaching steps. The leaching process was simulated as a series of batch leaching experiments. Parameter a of the model was determined based on the selectivity coefficient of the balanced batch leaching experiment. Further, the influences of ammonium sulfate concentration, rare earth grade, column height, permeability coefficient, and hydrodynamic dispersion coefficient on the extraction were analyzed. Relationships between parameter b, the ammonium sulfate concentration, and the physical and mechanical properties of the ore column, were examined using dimensional analysis. It was determined that the optimal ammonium sulfate concentration for different column heights(2.5, 5.0,7.5, and 10.0 cm) using the mathematical model were 5.9, 6.2, 7.3, and 7.7 g/L, respectively. The mathematical model can be used to estimate the breakthrough curve, leaching rate, and leaching period of rare earth ores, to achieve optimal extraction.     

某高磷铁矿提铁降磷研究

以湖南某地高磷铁矿为原料,采用还原焙烧一磁选一硫酸浸出工艺进行提铁降磷试验研究.对还原焙烧一磁选粗精矿进行硫酸浸出工艺参数优化,对浸出时间、液固比、硫酸用量和搅拌速度等因素对提铁降磷效果的影响进行研究.研究结果表明:对原矿品位为47.28%Fe(质量分数)和磷含量为1.59%的高磷铁矿石经过还原焙烧一磁选得到的粗精矿,在浸出时间为2h、液固比为2.5、硫酸用量为50 kg/t和搅拌速度为500 r/min的条件下进行酸性浸出提铁降磷,最终得到铁精矿品位达62.35%Fe,磷含量为0.20%,铁总回收率为90.54%和脱磷率为87.42%.     

用硫酸亚铁浸出同时沉淀铁矾法处理低品位锰矿

研究了用硫酸亚铁浸出同时沉淀黄钠铁矾的方法处理低品位软锰矿的过程．在该过程中,软锰矿中的ＭｎＯ２被还原成ＭｎＳＯ４同时ＦｅＳＯ４被氧化并以黄钠铁矾的形态沉淀．沉淀产生的酸可直接用于ＭｎＯ２浸出．考察了硫酸、硫酸亚铁和硫酸钠的加入量及温度等参数对锰浸出和沉铁效率的影响．讨论了过程动力学．实验结果表明,锰浸出率和沉铁效率（质量分数）在最佳条件下可分别达到９６％和９２％．     

Investigation on leaching of malachite by citric acid

Leaching of an oxidized copper ore containing malachite, as a new approach, was investigated by an organic reagent, citric acid. Sulfuric acid is the most common reagent in the leaching of oxide copper ores, but it has several side effects such as severe adverse impact on the environment. In this investigation, the effects of particle size, acid concentration, leaching time, solid/liquid ratio, temperature, and stirring speed were optimized. According to the experimental results, malachite leaching by citric acid was technically feasible. Optimum leaching conditions were found as follows:the range of particle size, 105-150 μm; acid concentration, 0.2 M; leaching time, 30 min; solid/liquid ratio, 1:20 g/mL; temperature, 40℃; and stirring speed, 200 r/min. Under the optimum conditions, 91.61% of copper was extracted.     

废旧镍氢电池正极材料中镍和钴的回收

研究了在硫酸体系中回收废旧镍氢电池正极材料中的金属镍和钴. 用正交实验方法考察了浸出温度、浸出时间、硫酸初始浓度以及氧化剂用量对镍、钴浸出率的影响. 实验结果表明,各因素对镍和钴浸出率的影响程度排序均为:氧化剂用量>浸出时间>温度>硫酸初始浓度. 在实验得出的最佳浸出条件下,Co的浸出率为99.7%,Ni的浸出率为99.1%.     

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.     

Nickel solvent extraction from cold purification filter cakes of Angouran mine concentrate using LIX984N

Cold purification filter cakes generated in the hydrometallurgical processing of Angouran mine zinc concentrate commonly contain significant amounts of Zn, Cd, and Ni ions and thus are valuable resources for metal recovery. In this research, a nickel containing solution that was obtained from sulfuric acid leaching of the filter cake following cadmium and zinc removal was subjected to solvent extraction experiments using 10vol% LIX984N diluted in kerosene. Under optimum experimental conditions (pH 5.3, volume ratio of organic/aqueous (O:A) = 2:1, and contact time = 5 min), more than 97.1% of nickel was extracted. Nickel was stripped from the loaded organic by contacting with a 200 g/L sulfuric acid solution, from which 77.7% of nickel was recovered in a single contact at the optimum conditions (pH 1–1.5, O:A = 5:1, and contact time = 15 min).     

Extraction and separation of nickel and cobalt from saprolite laterite ore by microwave-assisted hydrothermal leaching and chemical deposition

Extraction and separation of nickel and cobalt from saprolite laterite ore were studied by using a method of microwave-assisted hydrothermal leaching and chemical deposition. The effects of leaching temperature and time on the extraction efficiencies of Ni2+ and Co2+ were investigated in detail under microwave conditions. It is shown that the extraction efficiencies of Ni2+ and Co2+ from the ore pre-roasted at 300℃ for 5 h were 89.19% and 61.89% when the leaching temperature and time were about 70℃ and 60 min, respectively. For the separation process of Ni and Co, the separation of main chemical components was performed by adjusting the pH values of sulfuric leaching solutions using a NaOH solution based on the different pH values of precipitation for metal hydroxides. The final separation efficiencies of Ni and Co were 77.29% and 65.87%, respectively. Furthermore, the separation efficiencies of Fe of 95.36% and Mg of 92.2% were also achieved at the same time.     

红土镍矿微波水热法浸提镍钴

采用微波水热盐酸浸出方法对腐泥土型红土镍矿提取镍钴进行了研究,详细探讨了焙烧预处理、微波水热浸出温度和浸出时间对镍钴浸出率的影响.对于300℃焙烧预处理后的红土镍矿,微波水热温度为50℃,浸出时间为1 h时,镍的浸出率高达93.65%,钴的浸出率为87.86%.红土镍矿的微波水热浸出体系与普通水热浸出体系相比,镍和钴的浸出效果更好.研究表明,扩散过程是镍、钴浸出过程的主要限制环节.     

Accelerating leaching of copper ore with surfactant and the analysis of reaction kinetics

To solve the problem of low permeability and lower extraction rates of high-mud ores, a surfactant was added as a penetrant to the pregnant leaching solution during column leaching tests. On the basis of the theories of physical chemistry and seepage flow mechanics, the mechanism by which seepage is enhanced under the effects of the surfactant was analyzed. The results show that the action modes of the surfactant were divided into four aspects: changing the wettability of the ore, reducing the viscosity of the leaching solution, adsorbing onto the surface of ore, and enhancing the permeability effect. The findings of column leaching tests demonstrated that permeability was substantially improved by the surfactant. In the later period of leaching, the permeability coefficient was two times higher than that of the control group. Meanwhile, the ore extraction rate increased by approximately 10%. During the leaching process, the surface tension of the solution did not substantially change, and that of the solution with surfactant increased slightly. The kinetics analysis of ore column leaching illustrated that the leaching processes were controlled by both internal diffusion(principal factor) and chemical reaction.     

低品位氧化锌矿的浓硫酸焙烧

采用浓硫酸焙烧新工艺,提取低品位氧化锌矿中的锌、铁、铝.考查了反应温度、反应时间及酸矿摩尔比对低品位氧化锌矿中锌、铁、铝提取率的影响.利用扫描电镜(SEM)和X射线衍射分析仪(XRD)对熟料和尾渣的微观形貌进行了表征并分析了生成产物成分.在此基础上,通过正交试验,确定最佳反应条件.研究结果表明:最佳反应条件为反应温度6735K,反应时间120min,酸矿摩尔比16∶1.此时,锌的提取率可达9902%,铁、铝的提取率分别为2378%与7768%,硅、钙、铅等元素富集在渣中.     

Innovative method for boron extraction from iron ore containing boron

A novel process for boron enrichment and extraction from ludwigite based on iron nugget technology was proposed. The key steps of this novel process, which include boron and iron separation, crystallization of boron-rich slag, and elucidation of the boron extraction behavior of boron-rich slag by acid leaching, were performed at the laboratory. The results indicated that 95.7% of the total boron could be enriched into the slag phase, thereby forming a boron-rich slag during the iron and slag melting separation process. Suanite and kotoite were observed to be the boron-containing crystalline phases, and the boron extraction properties of the boron-rich slag depended on the amounts and grain sizes of these minerals. When the boron-rich slag was slowly cooled to 1100℃, the slag crystallized well and the efficiency of extraction of boron (EEB) of the slag was the highest observed in the present study. The boron extraction property of the slow-cooled boron-rich slag obtained in this study was much better than that of szaibelyite ore under the conditions of 80% of theoretical sulfuric acid amount, leaching time of 30 min, leaching temperature of 40℃, and liquid-to-solid ratio of 8 mL/g.