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.     

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.     

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.     

Extraction of vanadium from high calcium vanadium slag using direct roasting and soda leaching

The extraction of vanadium from high calcium vanadium slag was attempted by direct roasting and soda leaching. The oxidation process of the vanadium slag at different temperatures was investigated by X-ray diffraction (XRD), scanning electron microscopy (SEM), and energy dispersive spectroscopy (EDS). The effects of roasting temperature, roasting time, Na₂CO₃ concentration, leaching temperature, leaching time, and liquid to solid ratio on the extraction of vanadium were studied. The results showed that olivine phases and spinel phases in the vanadium slag were completely decomposed at 500 and 800℃, respectively. Vanadium-rich phases were formed at above 850℃. The leaching rate of vanadium reached above 90% under the optimum conditions:roasting temperature of 850℃, roasting time of 60 min, Na₂CO₃ concentration of 160 g/L, leaching temperature of 95℃, leaching time of 150 min, and liquid to solid ratio of 10:1 mL/g. The main impurities were Si and P in the leach liquor.     

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%.     

Use of ionic liquid in leaching process of brass wastes for copper and zinc recovery

Brass ash from the industrial brass manufacturer in Turkey was leached using the solutions of ionic liquid (IL) 1-butyl-3-methyl-imidazolium hydrogen sulfate ([bmim]HSO₄) at ambient pressure in the presence of hydrogen peroxide (H₂O₂) and potassium peroxymonosulfate (oxone) as the oxidants. Parameters affecting leaching efficiency, such as dissolution time, IL concentration, and oxidizing agent addition, were investigated. The results show that [bmim]HSO₄ is an efficient IL for the brass ash leaching, providing the dissolution efficiencies of 99% for Zn and 24.82% for Cu at a concentration of 50vol% [bmim]HSO₄ in the aqueous solution without any oxidant. However, more than 99% of zinc and 82% of copper are leached by the addition of 50vol% H₂O₂ to the [bmim]HSO₄ solution. Nevertheless, the oxone does not show the promising oxidant behavior in leaching using [bmim]HSO₄.     

A novel hydrothermal method for zinc extraction and separation from zinc ferrite and electric arc furnace dust

A novel hydrothermal process was developed to extract zinc from pure zinc ferrite (ZnFe₂O₄) nanopowder and zinc-containing electric arc furnace (EAF) dust using hexahydrated ferric chloride (FeCl₃·6H₂O) as a decomposing agent. The effects of solid FeCl₃·6H₂O to ZnFe₂O₄ ratio by mass (R_F/Z), hydrothermal reaction temperature, and time on zinc extraction were systematically investigated. In the results, when the hydrothermal reaction is conducted at 150℃ for 2 h with R_F/Z of 15:20, the efficiency of zinc extraction from ZnFe₂O₄ reaches 97.2%, and the concentration of ferric ions (Fe3+) in the leaching solution is nearly zero, indicating a high selectivity for zinc. In addition, the zinc extraction efficiency from the EAF dust reaches 94.5% in the case of the hydrothermal reaction performed at 200℃ for 10 h with the solid FeCl₃·6H₂O to EAF dust ratio by mass (R_{F/EAF dust}) of 15:10. Zinc and iron separation is achieved by adjusting the pH value of the leaching solution according to the different precipitation pH values of metal hydroxides.     

Alkaline digestion behavior and alumina extraction from sodium aluminosilicate generated in pyrometallurgical process

In pyrometallurgical process, Al-and Si-bearing minerals in iron and aluminum ores are easily transformed into sodium aluminosilicates in the presence of Na₂O constituents, which alters the leaching behaviors of Al₂O₃ and SiO₂. It was confirmed that sodium aluminosilicates with different phase compositions synthesized at various roasting conditions were effectively digested in the alkaline digestion process. Under the optimum conditions at temperature of 100-120℃, liquid-to-solid ratio (L/S) of 10:2 mL/g, caustic ratio of 4, and Na₂O concentration of 240 g/L, the actual and relative digestion ratio of Al₂O₃ from the synthesized sodium aluminosilicates reached maximums of about 65% and 95%, respectively, while SiO₂ was barely leached out. To validate the superior digestion property of sodium aluminosilicate generated via an actual process, the Bayer digestion of an Al₂O₃-rich material derived from reductive roasting of bauxite and comprising Na_1.75Al_1.75Si_0.25O₄ was conducted; the relative digestion ratio of Al₂O₃ attained 90% at 200℃.     

Kaolin-based geopolymers with various NaOH concentrations

Kaolin geopolymers were produced by the alkali-activation of kaolin with an activator solution (a mixture of NaOH and sodium silicate solutions). The NaOH solution was prepared at a concentration of 6–14 mol/L and was mixed with the sodium silicate solution at a Na₂SiO₃/NaOH mass ratio of 0.24 to prepare an activator solution. The kaolin-to-activator solution mass ratio used was 0.80. This paper aimed to analyze the effect of NaOH concentration on the compressive strength of kaolin geopolymers at 80℃ for 1, 2, and 3 d. Kaolin geopolymers were stable in water, and strength results showed that the kaolin binder had adequate compressive strength with 12 mol/L of NaOH concentration. When the NaOH concentration increased, the SiO₂/Na₂O decreased. The increased Na₂O content enhanced the dissolution of kaolin as shown in X-ray diffraction (XRD) and Fourier transform infrared spectroscopy (FTIR) analyses. However, excess in this content was not beneficial for the strength development of kaolin geopolymers. In addition, there was the formation of more geopolymeric gel in 12 mol/L samples. The XRD pattern of the samples showed a higher amorphous content and a more geopolymer bonding existed as proved by FTIR analysis.     

Isothermal reduction of titanomagnetite concentrates containing coal

The isothermal reduction of the Panzhihua titanomagnetite concentrates (PTC) briquette containing coal under argon atmosphere was investigated by thermogravimetry in an electric resistance furnace within the temperature range of 1250–1350℃. The samples reduced in argon at 1350℃ for different time were examined by X-ray diffraction (XRD) analysis. Model-fitting and model-free methods were used to evaluate the apparent activation energy of the reduction reaction. It is found that the reduction rate is very fast at the early stage, and then, at a later stage, the reduction rate becomes slow and decreases gradually to the end of the reduction. It is also observed that the reduction of PTC by coal depends greatly on the temperature. At high temperatures, the reduction degree reaches high values faster and the final value achieved is higher than at low temperatures. The final phase composition of the reduced PTC-coal briquette consists in iron and ferrous-pseudobrookite (FeTi₂O₅), while Fe_2.75Ti_0.25O₄, Fe_2.5Ti_0.5O₄, Fe_2.25Ti_0.75O₄, ilmenite (FeTiO₃) and wustite (FeO) are intermediate products. The reaction rate is controlled by the phase boundary reaction for reduction degree less than 0.2 with an apparent activation energy of about 68 kJ·mol?1 and by three-dimensional diffusion for reduction degree greater than 0.75 with an apparent activation energy of about 134 kJ·mol?1. For the reduction degree in the range of 0.2–0.75, the reaction rate is under mixed control, and the activation energy increases with the increase of the reduction degree.     

臭氧-过氧化氢联合浸出方铅矿

在盐酸溶液中,以臭氧和过氧化氢为氧化剂、三氯化铁为助浸剂联合浸出方铅矿精矿制备氯化铅,考察各种操作参数对铅浸出率的影响.实验结果表明:搅拌速率为500r/min、氯化钠初始质量浓度为250g/L、反应温度为90℃、反应时间为180min、三氯化铁初始质量浓度为259/L、过氧化氢初始质量浓度为6.669/L、盐酸初始浓度为0.3mol/L、臭氧进口氧气流量为1.0L/min是臭氧-过氧化氢联合浸出方铅矿的最佳操作参数,此时铅的浸出率达99.5％,产物氯化铅纯度达99.6％.     

Photoluminescence of ZnO nanorods grown by hydrothermal method on Si substrate

ZnO nanorods were grown on Si substrate by hydrothermal method under various reaction time (12, 24, and 36 h), Zn²⁺ concentrations (0.01, 0.02, and 0.05 mol/L) and reaction temperatures (70–95 °C). Their photoluminescence (PL) tests were taken at room temperature. Nanorods grown under longer reaction times or higher temperatures can emit stronger UV light and depressed green light, suggesting better crystallization of ZnO nanorods. Higher Zn²⁺ concentration results in stronger green band emitting, whereas the UV peak is depressed with the Zn²⁺ concentration over 0.02 mol/L. This indicates that excessive interstitial Zn defects may be introduced into the nanorods in Zn-rich environment.     

Isothermal oxidation behavior and mechanism of a nickel-based superalloy at 1000℃

The oxidation behavior of a nickel-based superalloy at 1000°C in air was investigated through X-ray diffraction, scanning electron microscopy, andenergy-dispersive spectroscopy analysis. A series of oxides, including external oxide scales (Cr2O3, (TiO2 + MnCr2O4)) and internal oxides (Al2O3,TiN), were formed on the surface or sub-surface of the substrate at 1000°C in experimental still air. The oxidation re-sistance of the alloy was dependent on the stability of the surface oxide layer. The continuity and density of the protective Cr2O3 scale were affected by minor alloying elements such as Ti and Mn. The outermost oxide scale was composed of TiO2 rutile and MnCr2O4 spinel, and the growth of TiO2particles was controlled by the outer diffusion of Ti ions through the pre-existing oxide layer. Severe internal oxidation oc-curred beneath the external oxide scale, consuming Al and Ti of the strength phaseγ′ (Ni3(Al,Ti)) and thereby severely deteriorating the sur-face mechanical properties. The depth of the internal oxidation region was approximately 35μm after exposure to experimental air at 1000°C for 80 h.     

Dissolution kinetics of iron oxides in clay in oxalic acid solutions

Clay samples containing 8.15% iron oxides and 27.49% alumina were leached in oxalic acid. Leaching experiments were performed in aqueous solutions of oxalic acid of 0.2-2 mol/L at 40-80℃ for up to 90 min. The mixed kinetic mechanism, i.e., t/τ=[(1-2X/3)-(1-X)2/3]+b[1-(1-X)1/3], seemed to be the most appropriate one to fit the kinetic data of leaching iron oxides contained in clay in the aqueous oxalic acid solutions. The Arrhenius activation energy for leaching in the 1.8 mol/L oxalic acid was found to be 41.035 kJ/mol.     

盐酸体系中二氧化锰浸出方铅矿的研究

在盐酸溶液中,用二氧化锰浸出方铅矿精矿制备氯化铅,并考察操作参数对铅浸出率的影响。研究结果表明盐酸体系中二氧化锰浸出方铅矿的最佳操作条件如下:搅拌速率为500 r/min,反应体系中总液体与总固体质量比为10,二氧化锰与方铅矿精矿质量比为1.3,反应时间为60 min,反应温度为80℃,盐酸浓度为3 mol/L,氯化钠质量浓度为250 g/L;在此最佳操作条件下,方铅矿精矿中的铅浸出率在99.5%以上,氯化铅产物纯度在99.6%以上。     

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.     

漂浮阳极泥中铋的提取与三氧化二铋的制备

漂浮阳极泥经过盐酸浸出、稀释水解、氢氧化钠中和得到氯氧铋,氯氧铋经过氢氧化钠转化制备得到三氧化二铋.当盐酸浓度为6mol/L,固液比为1∶5,反应温度为80℃,反应时间为1h时,漂浮阳极泥中锑和铋浸出率分别达到99.17％和99.08％.当稀释比为8∶1时,盐酸浸出液中锑水解率为98.13％,铋水解率仅为8.8％.稀释后液中加入氢氧化钠溶液,当pH为1.5时,铋水解率达到99.5％,水解产物氯氧铋(BiOCl)中铋、氧、氯的质量分数分别为54.23％,19.30％和14.61％.氯氧铋再次经过盐酸浸出,稀释水解,氢氧化钠沉淀得到氯氧铋.除杂后氯氧铋经过硫酸洗涤、氢氧化钠转化,当氢氧化钠浓度为6mol/L,液固比为3∶1,反应温度为80℃时,反应2h后过滤,用0.5mol/L盐酸洗涤得到形貌为纤维状、晶型为单斜的α-Bi2O3,氧化铋纯度达到99.81％.     

不同气氛下氟醚橡胶热分解动力学的对比研究

利用热失重分析方法(TGA),研究了氟醚橡胶生胶在空气和氮气中的热分解规律,并探讨了五种升温速率下的热分解动力学.应用Flynn-Wall-Ozawa法计算获得了其热解过程的动力学参数,并利用Popescu法推断得到了热解过程的反应机理函数.研究结果表明:氟醚橡胶生胶只呈现出一个主要的热失重峰;升温速率越大,热分解温度越高;氮气中更稳定且氧气对其热分解有一定的促进作用;空气中的热分解活化能平均值为172.5kJ/mol,氮气中的为260.9kJ/mol;两种气氛下,热解过程均不能由单一的机理函数来描述;空气中,340℃~370℃阶段机理符合相边界反应,球形对称,370℃~380℃阶段机理函数符合Ginstling-Brounshtein(G-B)方程,380℃~400℃阶段符合Zhuralev(Zh)方程;氮气中,400℃~430℃阶段机理函数符合Ginstling-Brounshtein(G-B)方程,430℃~460℃阶段机理符合相边界反应,球形对称.