Leaching kinetics of low grade zinc oxide ore in NH3-NH4Cl-H2O system

The leaching kinetics of low grade zinc oxide ore in NH3-NH4Cl-H2O system was studied. The effects of ore particle size, reaction temperature and the sum concentration of ammonium ion and ammonia on the leaching efficiency of zinc were examined. The leaching kinetics of low-grade zinc oxide ore in NH3-NH4Cl-H2O system follows the kinetic law of shrinking-core model. The results show that diffusion through the inert particle pores is the leaching kinetics rate controlling step. The calculated apparent activation energy of the process is about 7.057 kJ/mol. The leaching efficiency of zinc is 92.1% under the conditions of ore particle size of 69μm, holding at 80℃ for 60 min, sum ammonia concentration of 7.5 mol/L, the molar ratio of ammonium to ammonia being 2：1, and the ratio （g/mL） of solid to liquid being 1：10.     

Leaching kinetics of low grade zinc oxide ore in NH3-NH4Cl-H2O system

The leaching kinetics of low grade zinc oxide ore in NH3-NH4Cl-H2O system was studied. The effects of ore particle size,reaction temperature and the sum concentration of ammonium ion and ammonia on the leaching efficiency of zinc were examined.The leaching kinetics of low-grade zinc oxide ore in NH3-NH4Cl-H2O system follows the kinetic law of shrinking-core model. The results show that diffusion through the inert particle pores is the leaching kinetics rate controlling step. The calculated apparent activation energy of the process is about 7.057kJ/mol. The leaching efficiency of zinc is 92.1% under the conditions of ore particle size of 69μm, holding at 80℃ for 60min, sum ammonia concentration of 7.5mol/L, the molar ratio of ammonium to ammonia being 2:1, and the ratio (g/mL) of solid to liquid being 1:10.     

Leaching kinetics of low-grade copper ore with high-alkality gangues in ammonia-ammonium sulphate solution

The leaching kinetics of low-grade copper ore with high-alkality gangues was studied in ammonia-ammonium sulphate solution. The main parameters, such as ammonia and ammonium sulphate concentrations, particle size, solid-to-liquid ratio and reaction temperature, were chosen in the experiments. The results show that the increase of temperature, concentrations of ammonia and ammonium sulphate is propitious to the leaching rate of copper ore. The leaching rate increases with the decrease of particle size and solid-to-liquid ratio. The leaching rate is controlled by the diffusion through the ash layer and the activation energy is determined to be 25.54 kJ/mol. A semi-empirical equation was proposed to describe the leaching kinetics.     

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.     

冶金污泥中铜锌镉铅的氨水和碳酸铵选择性浸出分离

为了将冶金污泥中的有价金属资源化并合理回收与利用,用采用氨水和铵盐复配对冶金污泥中的铜、锌、镉和铅进行选择性浸出,并且通过单因素条件试验分别考察铵盐种类、氨水和铵盐复配体系作为浸出剂中氨水和铵盐各自的浓度、浸出反应的温度和浸出反应时间等因素对铜、锌、镉和铅浸出率的影响,从而通过对比得出最佳的金属浸出工艺条件:复配体系中铵盐选择为碳酸铵,浓度为0.5mol/L,缓冲剂氨水浓度为3mol/L,反应温度为50℃,浸出反应时间时间为120min.在此条件下,铜、锌和镉的浸出率分别为91.51%、85.13%和89.51%,而铅的浸出率则低于6.00%,本工艺可以实现铜、锌和镉的回收同时可以实现对铅元素的基本分离.     

Facile and controlled synthesis of silica sol nanospheres through a modified sol-gel process

An effective and reproducible preparation of silica sol nanospheres via a modified sol-gel process has been described. Monodisperse and stable silica sol nanospheres with uniformsize were successfully obtained through the optimized synthesis in which the mixture of tetraethyl orthosilicate （TEOS） and ethanol was followed by the addition of water and ammonium hydroxide （NH3） separately, and the size of silica sol spheres was strictly controlled in the range of 25-119 nm with a narrow size distribution by fine adjustment of several reaction parameters. Results showed that in the presence of low concentration of TEOS, spheres size rose first and reached maximum when H2O concentration was up to 66 g/L. However, the diameter of silica sol spheres decreased above 66 g/L of H2O concentration. Furthermore, it was also found that the size and size distribution of silica sol nanospheres were affected by NH3 concentration. As NH3 concentration increased from 15 to 35 g/L, the diameter declined from 83 to 64 nm. Nevertheless, higher NH3 concentration would result in relatively broad size distribution, and gelation occurred when NH3 concentration reached 44 g/L. In addition, the effect of the different feed rates ofNH3 on the size growth of silica sol nanospheres was also discussed.     

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.     

氨络合物体系中镍阴极电沉积的电化学行为

通过极化曲线测量，对氨络合物体系中镍阴极电沉积电化学行为进行研究，系统探讨了溶液中总镍离子浓度、氨水浓度、氯化铵浓度、阴离子及温度等工艺条件对镍阴极还原的影响。研究结果表明：镍放电电流随着总镍离子浓度的上升而上升，随氨水浓度的升高而降低；在1—4mol／L氯化铵浓度范围内，镍放电电流随其浓度的降低而升高，而当氯化铵浓度低于1mol／L时，镍放电电流出现下降的现象；氯盐氨络合物体系中镍阴极放电电流明显高于硫酸盐氨络合物体系镍放电电流，镍放电电流随温度的升高而升高。根据实验现象，进一步分析了镍阴极电沉积电化学行为变化的原因。     

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     

Octacalcium phosphate fiber synthesized by homogeneous precipitation method

Calcium phosphate fiber was synthesized by homogeneous precipitation method using urea as precipitation agent. Effects of the reactant concentration and hydrothermal temperature on the calcium phosphate morphology and composition were studied using SEM, FTIR and XRD. It is found that fine octacalcium phosphate(OCP)fiber can be synthesized when the reactant concentrations are 0.167 mol/L for Ca(NO3)2, 0.1 mol/L for (NH4)2HPO4, and 0.6-0.7 mol/L for (NH2)2CO, respectively, with the initial reactant solution pH value around 2.2, hydrothermal temperature 90 ℃ and termination reaction at pH value around 5.0. At very low urea concentration, the product size distribution is highly inhomogeneous, however, at an excessively high concentration the product becomes larger and shorter and a mixture of OCP and hydroxyapatite (HA). Increasing the hydrothermal temperature is favorable to the fast precipitation of OCP, higher productivity and finer product.     

Preparation and characteristic research of anhydrous magnesium chloride with dehydrated ammonium carnallite

Taking the saline lake bischofite and NH4Cl that was removed with the ammonia method and continuously followed by filtration as raw materials with a molar ratio of 1 ： 1 of MgCl2 to NH4Cl, ammonium carnallite was synthesized. And then the ammonium carnallite was dehydrated to some extent at 160 ℃ for 4 h. Ammonium carnallite reacted with ammonia at 240 ℃ for 150 min and the ammonation ammonium carnallite was produced. Finally, the ammonation ammonium carnallite was calcined at 750 ℃ into anhydrous magnesium chloride containing only 0.1% （mass fraction） of MgO. On the other hand, dehydrated ammonium carnallite was mixed with the solid ammonium chloride at mass ratio 1 ： 4 at high temperature and with the differential pressure of NH3 above 30.5 kPa. The dehydrated ammonium carnallite of mixture was dehydrated at 410 ℃, and then calcined at 700 ℃ into anhydrous magnesium chloride with only 0. 087% （mass fraction） of MgO. X-ray diffraction and electron microscopy analysis results prove that anhydrous magnesium chloride obtained by both methods hasn＇t mixed phases,the particle is large and even has good dispersion, which is suitable for preparation of metal magnesium in the electrolysis.     

NH_3-(NH_4)_2SO_4体系浸析PCB污泥中铜的研究

利用NH3-(NH4)2SO4体系,对印刷电路板(printed circuit boards,PCB)生产过程中产生大量的含铜污泥中的铜进行浸析。对PCB污泥中重金属的质量分数进行了测定,其中铜在污泥中的质量分数为33.500%,其余金属质量分数较小。重点探讨了氨-硫酸铵浓度及pH、浸析时间、液固比、温度等条件对浸析率的影响。NH3-(NH4)2SO4体系对PCB污泥中铜浸析的最优条件为氨、硫酸铵浓度分别为3.0mol/L和1.5mol/L,液固比为20mL/g,浸析时间为180min,浸析温度为25℃。在最优条件下进行了浸析应用试验,结果表明铜的浸析率可达到97.5%,此方法的相对标准偏差(relative standard deviation,RSD)为0.63%。     

Zinc leaching from electric arc furnace dust in alkaline medium

Physical and chemical properties of electric arc furnace (EAF) dust from Tianjin seamless Pipe Company were measured and analyzed. The zinc leaching tests in alkaline medium were carried out under variation of leaching agent concentration, leaching temperature, leaching cumulative time and solid-to-liquid ratio. The thermodynamics and kinetics of the zinc leaching process were also analyzed. The results show that the EAF dust contains 10% (mass fraction) zinc and the median particle size is 0.69 μm. The zinc recovery of 73.4% is obtained under the condition of 90 °C, 6 mol/L NaOH, and 60 min leaching time. With the increase of concentration of NaOH and the cumulative time, zinc leaching will be significantly increased. The kinetics study demonstrates that the leaching reaction is chemically controlled and the reaction activation energy is 15.73 kJ/mol.     

Electric-oxidation extraction of molybdenite concentrate in alkaline NaCl electrolyte

The dissolution of molybdenite concentrate in NaCl electrolyte was investigated. The results show that the dissolution rate increases with the increase in liquid-to-solid ratio, stirring speed, NaCl concentration and temperature. When the liquid-to-solid ratio is 30:1, stirring speed is 400 r/min, concentration of NaCl is 4 mol/L at pH=9 and room temperature, the leaching efficiency of molybdenite concentrate will reach 99.5% in 240 min. Molybdenite concentrate cannot be electro-oxidized directly on the anode. The kinetic studies show that the dissolution of molybdenite concentrate is represented by shrinking core model with diffusion through a porous product layer of element sulfur, and the apparent activation energy for the dissolution reaction is 8.56 kJ/mol.     

Electric-oxidation extraction of molybdenite concentrate in alkaline NaCl electrolyte

The dissolution of molybdenite concentrate in NaCl electrolyte was investigated. The results show that the dissolution rate increases with the increase in liquid-to-solid ratio, stirring speed, NaCl concentration and temperature. When the liquid-to-solid ratio is 30:1, stirring speed is 400 r/min, concentration of NaCl is 4 mol/L at pH=9 and room temperature, the leaching efficiency of molybdenite concentrate will reach 99.5% in 240 min. Molybdenite concentrate cannot be electro-oxidized directly on the anode. The kinetic studies show that the dissolution of molybdenite concentrate is represented by shrinking core model with diffusion through a porous product layer of element sulfur, and the apparent activation energy for the dissolution reaction is 8.56 kJ/mol.     

Molybdenum removal from copper ore concentrate by sodium hypochlorite leaching

The removal of molybdenum from a copper ore concentrate by sodium hypochlorite leaching was investigated. The results show that leaching time, liquid to solid ratio, leaching ternperature, agitation speed, and sodium hypochlorite and sodium hydroxide concentrations all have a significant effect on the removal of molybdenum. The optimum process operating parameters were found to be: time, 4 h: sodium hydroxide concentration, 10%; sodium hypochlorite concentration, 8%; liquid to solid ratio, 10:1; temperature, 50℃; and,agitation speed, 500 r/min. Under these conditions the extraction of molybdenum is greater than 99.9% and the extraction of copper is less than 0.01%. A shrinking particle model could be used to describe the leaching process. The apparent activation energy of the dissolution reaction was found to be approximately 8.8 kJ/mol.     

Removal and utilization of calcite existed in scheelite by preparation of calcium sulfate whiskers

In order to eliminate the effect of calcite associated with scheelite on the scheelite flotation, hydrochloric acid was used to dissolve the calcite, and the soaking solution was used to prepare CaSO_4 whiskers by hydrothermal reaction with sulfuric acid at ambient pressure. First, the condition experiments of preparing CaSO_4 whiskers by using CaCl_2 and H_2SO_4 were carried out to optimize reaction parameters of the crystallization process. The optimal conditions were: at 102 ℃ reaction temperature, 0,5 mol/L reactant concentration and 60 min reaction time. Then based on the condition experiments and considering keeping acid concentration stable for achieving HCl recycling, Calcium sulfate whiskers with the average diameter of 1.41 μm and the average aspect ratio of 109 were prepared by the soaking solution after evaporating to half of its volume and 1.0 mol/L H_2SO_4 at 102 ℃ for 60 min: After ion exchange processing,the filtrate could be used as HCl in the process of HCl dissolution.     

Synthesis of γ-Al2O3 nanoparticles by chemical precipitation method

Highly pure active γ-Al₂O₃ nanoparticles were synthesized from aluminum nitrate and ammonium carbonate with a little surfactant by chemical precipitation method. The factors affecting the synthesis process were studied. The properties of γ-Al₂O₃ nanoparticles were characterized by DTA, XRD, BET, TEM, laser granularity analysis and impurity content analysis. The results show that the amorphous precursor Al(OH)₃ sols are produced by using 0.1 mol/L Al(NO₃)₃ · 9H₂O and 0.16 mol/L (NH₄)₂CO₃ · H₂O reaction solutions, according to the volume ratio 1.33, adding 0.024% (volume fraction) surfactant PEG600, and reacting at 40 °C, 1 000 r/min stirring rate for 15 min. Then, after stabilizing for 24 h, the precursors were extracted and filtrated by vacuum, washed thoroughly with deionized water and dehydrated ethanol, dried in vacuum at 80°C for 8 h, final calcined at 800 °C for 1 h in the air, and high purity active γ-Al₂O₃ nanoparticles can be prepared with cubic in crystal system, O _H ⁷ -FD3M in space group, about 9 nm in crystal grain size, about 20 nm in particle size and uniform size distribution, 131. 35 m²/g in BET specific surface area, 7 – 11 nm in pore diameter, and not lower than 99.93% in purity. Foundation item: Project(03JJY3015) supported by the Natural Science Foundation of Hunan Province     

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。     

Simulation calculation of solubility of insoluble compound MmAa in complex system

A simulation calculation model for the solubility of insoluble compound M _m A _a in complex system was established. According to coordination equilibrium principle, relevant dissociation reaction, complexation reaction, self-complexation reaction and protonation reaction during insoluble compound dissolving were considered and then the mass balance equations about solubility calculation were obtained. In the case analysis, the solubility of silver chloride in ammonia system was obtained by simulation calculation, and curved surface charts of thermodynamic equilibrium about the total concentration of silver ions, pH and concentration of ammonia ions were drawn correspondingly. The results show that under the conditions of room temperature and 6 mol/L ammonia concentration, the calculated solubility value of silver chloride (34 g/L) is close to the actual value (31 g/L), demonstrating that this model is suitable for solubility calculation of insoluble compound M _m A _a in the complex system.