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.     

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.     

工业废弃二氧化硅制水玻璃的正交试验研究

在单因素考察的基础上采用正交试验设计,对正交试验的结果进行了极差分析和方差分析,系统地研究了NaOH的质量分数、物料比、碱浸时间和碱浸温度对水玻璃模数的影响。然后用油浴代替水浴,提高制备水玻璃的温度,并对优化后的制备条件进行了单因素和正交试验的研究。结果表明,最优方案为NaOH溶液质量分数为10%,物料比为1∶3,碱浸时间为4 h,碱浸温度为150℃。最优方案下制得的水玻璃模数为2.4128。     

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     

Preparation of SiO2 Nanoparticles by Ion Exchanging and Study on Their Dispersion Stability

By adopting sodium silicate as a major material , SiO2 nanopartieles ( size in 8-15 rm ) waterdispersiod was prepared by ion exchanging. The effects of sodiwm silicate concentration, surface-rnodifying time,temperature and technologieal conditions on their diameter, size distribution and dispersion stability were also studied. The result show that, the hydrophilic litophlie and hydrophilic-lipophilie SiO2 nanopartieles water-dispersoid can be prepared through differenl kinds of surace-modifiers and the optimum reaction conditions have been determined as follows:sodiium silicate sohaion concentration :8w%; silicone dosage:3% of the total mass of nano SiO2 water-distprsoid； adding way and time of surface-moditfer:continual dropping for 2h;surface-modifying tempera-ture : 60- 70℃ .     

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.     

Electrical conductivity and dielectric property of fly ash geopolymer pastes

The electrical conductivity and dielectric property of fly ash geopolymer pastes in a frequency range of 100 Hz-10 MHz were studied.The effects of the liquid alkali solution to ash ratios(L/A) were analyzed.The mineralogical compositions and microstructures of fly ash geopolymer materials were also investigated using X-ray diffraction(XRD) and scanning electron microscopy(SEM).The 10 mol sodium hydroxide solution and sodium silicate solution at a sodium silicate-to-sodium hydroxide ratio of 1.0 were used in...     

Recovery of molybdenum from residues by simultaneous ultrafine milling and alkali leaching

A new technology of treating molybdenum residues by simultaneous ultmfme milling and alkali leaching was put forward to recover molybdenum from metallurgical residues. The effects of residue size, milling time, solid content, n(Na2CO3)/n(Mo) and slurry pH value on molybdenum leaching rate were investigated. The results indicate that a simpler process, lower slurry temperature, 50% shorter treating time, 60% decrease of Na2CO3 content and 15% increase of molybdenum leaching rate can be obtained by the new technology compared with the traditional process. The leaching kinetic equation was determined, and calculation of active energy (E = 56.2 kJ/mol) shows that the leaching process of molybdenum residues by simultaneous ultmfine milling and alkali leaching is controlled by chemical reaction. Potential exists for the new process to form the basis for an economically viable, environmentally friendly process to recover valuable elements from residues.     

Modification technology for separation of oily sludge

Based on the analysis of the properties of oily sludge samples, the effect of modification parameters, such as liquid to solid (L/S) ratio, agitation temperature, agitation intensity, agitation time and pH on the modification of oily sludge was investigated with the content of oil remnants in dry sludge as a reference index. Remixing experiments were carried out according to a simplex-lattice design, where Sx4056 was used as the demulsifier, petroleum sulfonate as the surfactant and sodium silicate (Na₂SiO₃) as the dispersant. The surface modification reagent formulation was optimized by a regression equation on the modified effect and based on the amounts of surface modification reagents. The results show that the content of the oil remaining in dry oily sludge is 0.28% of 10.15% oily sludge, when the reagent concentration rises to 3.5 g/L under the optimum experimental conditions.     

Preparation of spherical microfine silicon powder

Under certain conditions of proper temperature and pH value, sodium silicate was hydrolyzed in liquid ammonia chloride, and spherical microfine SiO2 powder in micrometer-size was prepared. In this experiment, the relationship between needed time and proper pH value, temperature, density of the solution, and its current capacity were found .The optimum conditions are pH 8.5, 70 ~C, and the concentration of sodium silicate is 0.6 mol/L for the density of the solution. The structure of spherical microfine silicon was characterized by SEM and XRD.     

Extracting gold from pyrite roster cinder by ultra-fine-grinding and resin-in-pulp

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氢氧化钠改性粉煤灰制吸附材料研究

以粉煤灰为原料制取水处理吸附材料。研究氢氧化钠浓度、回流时间、液固比、热处理时间和热处理温度对Cd2＋吸收率的影响。结果表明：氢氧化钠处理能显著提高粉煤灰的吸收能力;与未处理的粉煤灰相比,吸收率可由40．1％提高到80％以上：通过实验和分析得出氢氧化钠改性粉煤灰的较佳工艺条件：氢氧化钠浓度为3mol／L,回流时间为10h,氢氧化钠溶液与粉煤灰的液固比为6,热处理时间6h,热处理温度为90℃。     

Conversion of coal gangue into alumina,tobermorite and TiO_2-rich material

A large amount of coal gangue from coal mining and processing is regarded as waste and usually stockpiled directly. In order to recycle the valuable elements from the coal gangue, an integrated process is proposed. The process consists of three steps: 1concentrating alumina from the coal gangue via activation roasting followed by alkali leaching of Si O2 which produces alumina concentrate for alumina extraction by the Bayer process; 2) synthesizing tobermorite whiskers from the filtrated alkali liquo containing silicate via a hydrothermal method and reusing excess caustic liquor; and 3) enriching titanium component from the Baye process residue by sulfuric acid leaching. Alumina concentrate with 69.5% Al_2O_3 and mass ratio of alumina to silica(A/S) of 5.9pure 1.1 nm tobermorite whisker and TiO_2-rich material containing 33% TiO_2 are produced, respectively, with the optimal parameters Besides, the actual alumina digestion ratio of alumina concentrate reaches 80.4% at 270 oC for 40 min in the Bayer process.     

Extraction of nickel from molybdenum leaching residue of metalliferous black shale by segregation roasting and acid leaching

The recovery of nickel from molybdenum leach residue by the process of segregation roasting-sulfuric acid leaching-solvent extraction was investigated. The residue was characterized by microscopic investigations, using X-ray fluorescence spectrometry (XRF) and X-ray diffractometry (XRD) techniques and the residue after segregation roasting was characterized by chemical phase analysis method. A series of experiments were conducted to examine the mass ratio of activated carbon (AC) to the residue, segregation roasting time and temperature, sulfuric acid concentration, liquid-to-solid ratio, leaching time, leaching temperature, addition amount of 30% H₂O₂, stirring speed (a constant) on the leaching efficiency of nickel. A maximum nickel leaching efficiency of 90.5% is achieved with the mass ratio of AC to the residue of 1:2.5, segregation roasting time of 2 h, segregation roasting temperature of 850 °C, sulfuric acid concentration of 4.5 mol/L, liquid-to-solid ratio of 6:1, leaching time of 5 h, leaching temperature of 80 °C, addition of 30% H₂O₂ of 0.6 mL for 1 g dry residue. Under these optimized conditions, the average leaching efficiency of nickel is 89.3%. The nickel extraction efficiency in the examined conditions is about 99.6%, and the nickel stripping efficiency in the examined conditions is about 99.2%.     

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.     

电镀污泥中铜的浸出工艺及其动力学

以硫酸为浸出剂,对某表面处理工业园电镀废水处理污泥中的铜做了浸出试验研究.将污泥干燥、研磨,X射线衍射和X射线能谱仪分析表明污泥中含铜19.03%.采用单因素优化试验探讨了固液比、反应时间、浸出温度、硫酸质量分数、搅拌速度对铜浸出率的影响.结果表明:当硫酸质量分数为20%,固液比为1∶10,搅拌速率为700r/min时,在20℃下反应40min,铜的浸出率可达97%以上;根据未反应核收缩模型,对硫酸浸铜过程的动力学机理进行了研究,结果表明:硫酸浸铜过程的控制步骤为固体膜扩散控制,其反应级数为0.828 2,浸出活化能为11.809kJ/mol.研究为含铜电镀污泥安全处置提供理论依据.     

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.     

克二中区S7^2储层砂静态碱耗实验研究

选择新疆克摘玛依油田轱采试验区Ｔ２Ｋ１克中区Ｔ２Ｋ１的实际储层砂,在３０℃、ｔ０℃下,按１／２（ｇ／ｍｌ）,１／５（ｇ／ｍｌ）,１／１０（ｇ／ｍｌ）三个固／液比分别与１．６％ｗｔ的Ｎａ２ＣＯ３溶液和１％ｗｔ的ＮａＯＨ溶液反应８ｈ,１２０ｈ,３６０ｈ。监测了反应前后厌液浓度的变化,计算了各个条件下的绝对碱耗量。结果表明：随着固／液化减小和温度升高,储层砂的绝对碱耗量增加;１％ｗｔＮａＯＨ的碱耗量大于     

一种改性钛渣钛组分的分离方法

研究了碱浸出过程中的影响因素,考察了浸出温度,NaOH溶液的浓度,以及物料粒度和液固比等因素对于浸出效果和回收率的影响.得到了碱浸阶段适宜的工艺条件为：溶液中NaOH的质量分数为15%、浸出温度100℃、浸出时间6 h、液固比20 mL/g、钛渣粒度小于38μm,搅拌速度500 r/min.采用三段浸出的方法获得了品位为91.42%,回收率为78.47%的人造金红石产品.     

高硅氧化锌矿浸出工艺的研究

针对高硅氧化锌矿的处理进行了研究,提出了二段循环酸浸的湿法冶金工艺.通过条件优化实验确定一、二段浸出条件分别为:始酸浓度75g/L、温度70℃、液固比6∶1、浸出时间1.5 h;始酸浓度18g/L、温度60℃、液固比3∶1、浸出时间0.5 h.本试验采用一段六轮循环浸出及二段三轮循环浸出,运行稳定,锌的总浸出率达到93.4%.