Extraction of valuable metals from low-grade nickeliferous laterite ore by reduction roasting-ammonia leaching method

Nickel and cobalt were extracted from low-grade nickeliferous laterite ore using a reduction roasting-ammonia leaching method. The reduction roasting-ammonia leaching experimental tests were chiefly introduced, by which fine coal was used as a reductant. The results show that the optimum process conditions are confirmed as follows: in reduction roasting process, the mass fraction of reductant in the ore is 10%, roasting time is 120 min, roasting temperature is 1 023–1 073 K; in ammonia leaching process, the liquid-to-solid ratio is 4:1(mL/g), leaching temperature is 313 K, leaching time is 120 min, and concentration ratio of NH₃ to CO₂ is 90 g/L:60 g/L. Under the optimum conditions, leaching efficiencies of nickel and cobalt are 86.25% and 60.84%, respectively. Therefore, nickel and cobalt can be effectively reclaimed, and the leaching agent can be also recycled at room temperature and normal pressure.     

Kinetics of nickel leaching from roasting-dissolving residue of spent catalyst with sulfuric acid

Sulfuric acid leaching process was applied to extract nickel from roasting-dissolving residue of a spent catalyst, the effect of different parameters on nickel extraction was investigated by leaching experiments, and the leaching kinetics of nickel was analyzed. The experimental results indicate that the effects of particle size and sulfuric acid concentration on the nickel extraction are remarkable; the effect of reaction temperature is mild; while the effect of stirring speed in the range of 400–1 200 r/min is negligible. Decreasing particle size or increasing sulfuric acid concentration and reaction temperature, the nickel extraction efficiency is improved. 93.5% of nickel in residue is extracted under suitable leaching conditions, including particle size (0.074–0.100) mm, sulfuric acid concentration 30% (mass fraction), temperature 80 °C, reaction time 180 min, mass ratio of liquid to solid 10 and stirring speed 800 r/min. The leaching kinetics analyses shows that the reaction rate of leaching process is controlled by diffusion through the product layer, and the calculated activation energy of 15.8 kJ/mol is characteristic for a diffusion controlled process. Foundation item: Project (50574101) supported by the National Natural Science Foundation of China; Project (2003UDBEA00C020) supported by the Collaborative Project of School and Province of Yunnan Province, China     

New technique of comprehensive utilization of spent Al2O3 -based catalyst

A new technology was developed to recover multiple valuable elements from the spent Al2O3-based catalyst by X-ray phase analysis and exploratory experiments. The experimental results show that in the condition of roasting temperature of 750 ℃ and roasting time of 30 min, molar ratio of Na2O to Al2O3 of 1.2, the leaching rates of alumina, vanadium and molybdenum in the spent catalyst are 97.2%, 95.8% and 98.9%, respectively. Vanadium and molybdenum in sodium aluminate solution can be recovered by precipitators A and B, and the precipitation rates of vanadium and molybdenum are 94. 8% and 92. 6%. Al（OH）3 was prepared from sodium aluminate solution in the carbonation decomposition process, and the purity of Al2O3 is 99. 9% after calcination, the recovery of alumina reaches 90. 6% in the whole process; the Ni-Co concentrate was leached by sulfuric acid, a nickel recovery of 98. 2% and cobalt recovery over 98.5% can be obtained under the experimental condition of 30% H2SO4, 80 ℃, reaction time 4 h, mass ratio of liquid to solid 8, stirring rate 800 r/min.     

Extraction of vanadium from stone coal by modified salt-roasting process

In order to reduce the pollution of Cl₂ and HCl released during extracting vanadium from stone coal by sodium chloride roasting, a modified salt-roasting process was proposed by adding calcined lime in roasting process followed by H₂SO₄ leaching. The effects of parameters including roasting temperature, roasting time, addition mass ratio of NaCl, calcined lime upon leaching rate of vanadium and curing rate of chlorine were investigated, and the effects of leaching time and leaching temperature on leaching rate of vanadium were also studied. The results show that the vanadium leaching rate and the curing rate of chlorine are 67.3% and 51.5% (mass fraction), respectively, at roasting temperature of 750 °C, roasting time of 4 h, 15% sodium chloride and 8% (mass fraction) calcined lime, leaching temperature of room temperature, and leaching time of 4 h.     

Selective leaching of vanadium from V-Ti magnetite concentrates by pellet calcification roasting-H_2SO_4 leaching process

A novel method of pellet calcification roasting-H_2 SO_4 leaching was proposed to efficiently separate and extract vanadium(V) from vanadium-titanium(V-Ti) magnetite concentrates.The leaching rate of V is as high as 88.98%,while the leaching rate of impurity iron is only 1.79%.Moreover,the leached pellets can be used as raw materials for blast furnace ironmaking after secondary roasting.X-ray photoelectron spectroscopy(XPS) and scanning electron microscopy with energy dispersive X-ray spectrometry(SEMEDS) analyses showed that V~(3+) was oxidized to V~(5+) after roasting at 1200℃,and V~(5+) was then leached by H_2 SO_4.X-ray diffraction(XRD) analyses and single factor experiment revealed a minimal amount of dissolved Fe_2 O_3 during H_2 SO_4 leaching.Therefore,a high separation degree of V and iron(Fe) from V-Ti magnetite concentrate was achieved through H_2 SO_4 leaching.Compared with the traditional roastingleaching process,this process can achieve a high selectivity of V and Fe,and has excellent prospects for industrial production.     

含钒灰渣酸浸提钒工艺酸浸条件的试验研究

含钒灰渣中钒的酸浸效率直接影响了整个提钒工艺中钒的总回收率,同时,酸浸条件对钒的酸浸效率有着显著的影响作用.为获取含钒灰渣酸浸提钒工艺酸浸阶段的工艺条件,及含钒灰渣的酸浸特性,在对灰渣中钒的赋存状态进行全面分析的基础上,针对硫酸浓度、酸浸温度、酸浸时间和液固比等酸浸影响因素,在实验室分别进行了酸浸条件试验研究;分析酸浸液、酸浸残渣的成分,计算得到V2O5酸浸效率.试验结果表明:酸浸温度和硫酸浓度对酸浸效率起主要影响作用.合理工艺条件为:硫酸浓度为5.0～6.0mol/L,酸浸温度为115～120℃,酸浸时间为6h左右,液固比为2.5∶1～3∶1.在此条件下,V2O5酸浸效率达到85%以上.     

Process optimization of reaction of acid leaching residue of asbestos tailing and sodium hydroxide aqueous solution

Silica is the major component of the acid leaching residue of asbestos tailing. The waterglass solution can be prepared by the reaction of the residue with sodium hydroxide aqueous solution. Compared to the high temperature reaction method, this process is environmental friendly and low cost. In this paper, the reaction process of the residue and the sodium hydroxide aqueous solution is optimized. The optimum reaction process parameters are as follows: the usage of sodium hydroxide is 26.4 g/100 g acid leaching residue, the reaction temperature is 90℃, the reaction time is 1 h, and the ratio of the liquid/solid is 2.0. The significance sequence of the process parameters to the alkali leaching reaction effect is the usage of sodium hydroxide > the ratio of the liquid/solid > the reaction time > the reaction temperature. The significance sequence to the leaching ratio of SiO2 is the ratio of the liquid/solid > the usage of sodium hydroxide > the reaction time > the reaction temperature. The significance sequence to the modulus of the sodium silicate is the ratio of the liquid/solid > the usage of sodium hydroxide > the reaction time > the reaction temperature. Under the optimum conditions, the leaching ratio of the SiO2 is 77.5%, and the modulus of the sodium silicate is 3.15. The XRD analysis result indicates that the major components of the alkali leaching residue are serpentine, talc, quartz and some albite.     

Study on hydrometallurgical process and kinetics of manganese extraction from low-grade manganese carbonate ores

In order to utilize low-grade manganese ore resources effectively, a hydrometallurgical process was developed for manganese extraction in dilute sulfuric acid medium, and the kinetics of leaching manganese was also investigated. At room temperature, manganese from low-grade manganese carbonate ores was extracted by sulfuric acid leaching without reductants. During the extracting process, single-factor analysis method was used to evaluate the effects of grinding fineness, sulfuric acid concentration, liquid-to-solid ratio, agitation rate and leaching time on the leaching efficiencies of Mn and Fe. The optimal leaching conditions are determined as coarse particles of below 2 mm size (without ball-milling), sulfuric acid concentration of 0.86 mol/L, liquid-to-solid ratio of 5:1, agitation rate of 150 r/min and leaching for 180 min at room temperature. Under the optimal conditions, the leaching efficiencies of Mn and Fe are 96.21% and 13.35%, respectively. In addition, through the experiments at different temperatures, it is found that the leaching process follows the shrinking core model under the conditions of changing acid concentration and intermittent reaction device. Moreover, the apparent activations of effective diffusion and chemical reaction in the kinetic model are calculated to be 18.83 and 27.15 kJ/mol, respectively.     

Recovery of copper sulfate after treating As-containing wastewater by precipitation method

The solid sodium hydroxide neutralized acidic As-containing wastewater till pH value was 6. Green copper arsenite was prepared after copper sulfate was added into the neutralized wastewater when the molar ratio of Cu to As was 2:1 and pH value of the neutralized wastewater was adjusted to 8.0 by sodium hydroxide. The arsenious acid solution and red residue were produced after copper arsenite mixed with water according to the ratio of liquid to solid of 4:1 and copper arsenite was reduced by SO₂ at 60 °C for 1 h. The white powder was gained after the arsenious acid solution was evaporated and cooled. Copper sulfate solution was obtained after the red residue was leached by H₂SO₄ solution under the action of air. The results show that red residue is Cu₃(SO₃)₂·2H₂O and the white powder is As₂O₃. The leaching rate of Cu reaches 99.00% when the leaching time is 1.5 h, molar ratio of H₂SO₄ to Cu is 1.70, H₂SO₄ concentration is 24% and the leaching temperature is 80 °C. The direct recovery rate of copper sulfate is 79.11% and the content of CuSO₄·5H₂O is up to 98.33% in the product after evaporating and cooling the copper sulfate solution.     

用硫酸盐法从锂云母中浸取锂

以锂云母精矿为原料,考察了用硫酸盐法浸取锂时的煅烧温度、煅烧时间、酸矿质量比、酸化焙烧时间等影响因素。结果表明：适当升高煅烧温度、增加煅烧时间、增大酸矿质量比、增加焙烧时间均有利于锂的溶出。最佳反应条件为：950℃煅烧2h,浓硫酸酸化焙烧3h,酸矿质量比为1．5：1,在此条件下锂的溶出率可达到96．95％。     

电镀污泥中铬的无害化处理及动力学分析

以某表面处理工业园电镀废水处理污泥为研究对象,以铬浸出率为指标,通过对重金属的浸出,分步回收达到无害化、资源化的目的.将污泥干燥、研磨,在不同浓度硫酸溶液中浸出,控制浸出时间、浸出温度和搅拌速率;浸出完成后抽滤使浸出液与残渣分离.采用正交试验法,确定对铬浸出效果影响因素的顺序为:硫酸浓度>搅拌速度>浸出时间>固液比.通过单因素优化试验,结果显示:当浸出温度为25 ℃、固液比为1∶15、浸出时间为20 min、搅拌速率为800 r/min、硫酸体积分数为30%时,铬的浸出率最高.最后用黄钠铁矾法除铁,用焦亚硫酸钠还原六价铬,用氢氧化钠分步沉淀铬、镍重金属,锌则继续留在溶液中.电镀污泥的浸铬实验的浸出动力学研究结果表明硫酸作为浸出剂的反应级数为1,反应的速率常数为:k=0.053 2e-4.52/RT.     

Process mineralogy approach to optimize curing-leaching in vanadium-bearing stone coal processing plants

The purpose of this study is to apply process mineralogy as a practical tool to further understand and analyze the reasons for low leaching rates in the curing-leaching process of vanadium-bearing stone coal and to find a solution or improvement to optimize the leaching index. Using vanadium-bearing stone coal with the V₂O₅ mass fraction of 0.88% as the research object, the effects of particle size, mineral composition, and sulfuric acid curing on the feed, intermediate, an...     

Nickel extraction from low grade laterite by agitation leaching at atmospheric pressure

This research work deals with the extraction of nickel from a low grade nickel laterite ore, taken from a deposit located in southwestern of Iran, through agitation leaching at atmospheric pressure. The assaying and mineralogical studies carried out on the nickel laterite sample, showed the 0.88% Ni, and principally consisted of oxide and silicate crystalline phases i.e. dolomite, quartz, magnetite, and goethite. Among numerous factors affecting such process, four major parameters i.e. temperature, agitator speed (r/min), leaching agents and their concentration were considered in a two-level full factorial experimental design. The agitation leach tests showed that the ore could be leached at atmospheric pressure with sulfuric acid while citric acid was almost unpromising. Analysis of variance (ANOVA) using DX7 software was employed to identify effective parameters. Sulfuric acid concentration and temperature were the most effective parameters on Ni extraction. Furthermore, the factorial models for experiment responses were developed. The results showed 83% Ni extraction after 4 h leaching, under optimized conditions i.e. temperature at 95 °C, acid concentration at 5 N and agitator speed at 1000 r/min. This study revealed that factorial experimental design can be implemented to identify effective parameters on the agitation leaching process of nickel laterite.     

Technological conditions and kinetics of leaching copper from complex copper oxide ore

The kinetic behavior of leaching copper from low grade copper oxide ore was investigated. The effects of leaching temperature, H₂SO₄ concentration, particle size of crude ore and agitation rate on the leaching efficiency of copper were also evaluated. And the kinetic equations of the leaching process were obtained. The results show that the leaching process can be described with a reaction model of shrinking core. The reaction can be divided into three stages. The first stage is the dissolution of free copper oxide and copper oxide wrapped by hematite-limonite ore. At this stage, the leaching efficiency is very fast (leaching efficiency is larger than 60%). The second stage is the leaching of diffluent copper oxides, whose apparent activation energy is 43.26 kJ/mol. During this process, the chemical reaction is the control step, and the reaction order of H₂SO₄ is 0.433 84. The third stage is the leaching of copper oxide wrapped by hematite-limonite and silicate ore with apparent activation energy of 16.08 kJ/mol, which belongs to the mixed control.     

Extraction of metals from boiler slag by sintering

A new method of recycling aluminum and iron in boiler slag derived from plants that use coal as fuel was presented. The new method can integrate efficient extraction and reuse of the leached pellets together. An elemental analysis of aqueous solutions leached by sulfuric acid was conducted by the EDTA-Na₂-ZnCl₂ titration method, and the components and microstructures of the samples were examined by means of XRF, XRD and SEM. An aluminum extraction efficiency of 86.50% was achieved when the sintered pellets were leached using 4 mol·L⁻¹ H₂SO₄ with solid/liquid ratio(m/V) of 1:5 at 80 °C for 24 h. An iron extraction efficiency of 94.60% was achieved under the same condition for the maximum extraction efficiency of Al. The extraction efficiency of Al and Fe increased with temperature, leaching time and acidity. The concentration of alumina and iron hydroxide in the final product was determined to be 99.12% and 92.20% respectively. This product of alumina would be used directly for the production of metallic aluminum.     

Dissolution kinetics of malachite in ammonia/ammonium sulphate solution

The dissolution kinetics of malachite was investigated in ammonia/ammonium sulphate solution. The effects of ammonia and ammonium sulphate concentration, pH, leaching time, reaction temperature, and particle size were determined. The results show that the optimum leaching conditions for malachite ore with a copper extraction more than 96.8% are ammonia/ammonium concentration 3.0 mol/L NH₄OH + 1.5 mol/L (NH₄)₂SO₄, liquid-to-solid ratio 25:1 mL/g, leaching time 120 min, stirring speed 500 r/min, reaction temperature 25 °C and particle size finer than 0.045 mm. The dissolution process of malachite with an activation energy of 26.75 kJ/mol is controlled by the interface transfer and diffusion across the product layer. A semi-empirical rate equation is obtained to describe the leaching process and the reaction orders with respect to concentration of ammonia and ammonium sulphate are 2.983 0 and 0.941 1, respectively.     

废旧印刷线路板两步浸出有价金属

为回收废旧印刷线路板中的有价金属,采用两步浸出的方法对其进行处理.先用双氧水-硫酸浸出贱金属,再用王水浸出贱金属浸出渣中的金.10 g废旧印刷线路板贱金属最佳浸出条件为双氧水20 mL,固液比1:5,硫酸浓度5 mol/L,反应温度60℃,反应时间90 min,原料溶损率达90.0%;金最佳浸出条件为反应温度40℃,反应时间30 min,浸出率达97.5%,研究证明两步浸出法能有效处理废旧印刷线路板,金溶出过程受扩散控制。     

硼铁矿中硼镁铁的硫酸法共浸出研究

提出一种采用硫酸酸浸硼铁矿使其中的硼、镁和铁元素共同浸出的方法．硫酸酸浸硼铁矿时,主要是与矿物中的硼镁石[Mg（BO2）（OH）]、磁铁矿[Fe3O4]、蛇纹石[Mg3Si2O5（OH）4]反应．通过热力学分析,验证采用硫酸共溶硼铁矿中的硼、镁和铁元素的可行性,并考察硫酸浓度、液固比、酸浸时间和酸浸温度对酸浸的影响,确定硫酸酸浸硼铁矿的最佳工艺条件：硫酸的质量分数30％,液固比（质量比）8：1,浸出温度90℃,浸出时间120min,搅拌速度大约100r／min．在最佳浸出条件下,硼铁矿中的硼、镁和铁元素的浸出率分别达到99．0％,91．0％,92．9％以上,达到了硼铁矿中硼、镁和铁元素共浸出的目的．     

Study on solubility of Nb2O5 in KOH solution and alkali leaching of niobite

Solubility of Nb₂O₅ and leaching behaviors of Nb and Ta from niobite in KOH solution have been investigated in order to develop an alkali hydrothermal leaching process of Nb and Ta. The solubility of Nb₂O₅ was measured in the range of 40 °C to 200 °C at various molar ratios of K₂O to Nb₂O₅(n(K₂O)/n(Nb₂O₅)). It has been found that Nb₂O₅ shows the maximum solubility at the solution composition of n(K₂O)/n(Nb₂O₅)=4/3 at a given temperature; the rise of temperature increases the solubility of Nb₂O₅ below 120 °C, but decreases it above 120 °C. The leaching behaviors of Nb and Ta were studied in the range of 150 °C to 250 °C and 0.1 MPa to 5 MPa. With the rise of temperature, the leaching degree increases when the leaching temperature is below 200 °C, but it decreases when the leaching temperature is above 200 °C. The maximum leaching degree is about 90% at 200 °C. It was proved that the alkali hydrothermal leaching process is effective for the recovery of Nb and Ta from niobite concentrate. Foundation item: The Key Project of Science and Technology Agency of Japan, 1994 Biography of the first author: ZHOU Kang-gen, doctor of engineering, professor, born in 1963, majoring in extractive metallurgy of rare metals and application of membrane separation technology.     

新疆石煤料团流化床焙烧提钒试验研究

在一台小型流化床燃烧试验台上对新疆石煤料团进行了焙烧特性的试验,着重考察了焙烧温度、焙烧时间、流化风速、添加剂种类对焙烧成球率的影响,并对飞灰、底渣、床料进行收集采样,利用水浸、质量分数为2%的Na2CO3溶液、6%的H2SO4溶液、10%的H2SO4溶液对各种样品浸取提钒,研究了焙烧温度、焙烧时间、浸取方式对转浸率的影响.结果表明:采用水泥为添加剂,温度为930℃,焙烧时间为90min,采用质量分数为10%的H2SO4溶液酸浸,可得较高焙烧成球率和转浸率,钒总回收率约为55.1%,同时可有效回收石煤热能,用于产汽发电.