金精矿碘化浸出过程动力学

采用碘-碘化物体系对金精矿进行浸出,通过实际矿石的浸出试验考察搅拌速度、浸出温度、碘浓度及碘离子浓度对金精矿碘化浸出反应动力学的影响。结果表明:提高搅拌速度对金的浸出有不利影响;通过对影响该体系动力学参数的考察,发现其符合核收缩模型,反应过程总体由界面化学反应控制,其表观活化能为31.674kJ/mol,碘浓度和碘离子浓度的反应级数分别为1和0.5,建立了金精矿碘化浸出过程的反应速率方程。     

湿法炼锌废渣中硫脲浸出银的动力学

探讨了从湿法炼锌废渣中用硫脲浸出回收银的浸出反应动力学,从这种难浸的含银炼锌废渣中用硫脲浸出回收银,浸出反应是一种典型的氧化还原反应并可充分进行,同时,通过动力学推导得出,从湿法炼锌废渣中用硫脲浸出银的反应动力学模型为收缩核动力学模型,同时计算出浸出活化能为13．26kJ/mol。该模型反映了浸出过程中控制整个反应速率的决定步骤是固膜扩散速率,并较好地说明了浸出机理。     

氨-硫酸铵体系过硫酸盐氧化含钙镁碳酸盐低品位铜矿的浸出动力学(英文)

研究在氨?硫酸铵体系中用过硫酸盐氧化低品位铜矿浸出动力学,确定搅拌速度、浸出温度、矿物粒度及氨、硫酸铵和过硫酸钠的浓度对浸出的影响。结果表明,搅拌速度在300r/min以上时对浸出速度无影响,浸出速度随反应温度及氨、硫酸铵和过硫酸钠浓度的增大而增加。对浸出渣的EDS和物相定量分析表明斑铜矿被过硫酸盐氧化而溶解于氨?硫酸铵溶液。用产物层的界面传质和扩散控制的收缩核模型分析铜矿的溶解动力学,其表观活化能为22.91kJ/mol,同时获得了描述浸出过程的半经验动力学方程,其对氨、硫酸铵和过硫酸钠的浓度的表观反应级数分别为0.5、1.2和0.5。     

银离子在氧化浸出黄铜矿中的动力学研究

研究了银离子在氧化剂A氧化浸出黄铜矿中的动力学,讨论了Ag^ 添加、Ag^ 浓度、氧化剂A浓度、黄铜矿粒度等因素对浸出过程的影响。研究结果表明,扩散控制模型可以很好地描述氧化剂A氧化浸出黄铜矿液-固反应过程;在实验范围内,确定了反应的比速率常数、反应级数及表观活化能,并根据实验数据,导出动力学方程为：1-2/3a-(1-a)^2/3=0.0685r。^-1[Ag^ ]^0.27exp(-25096/RT)t     

复杂高硅白合金硫酸氧化浸出工艺及提铜的动力学研究

研究了高硅白合金硫酸氧化体系下铜的浸出工艺与动力学.首先采用可控制变量法,通过单因素实验,系统研究了添加剂、硫酸浓度、反应温度、反应时间和液固比对铜钴浸出率的影响,其次通过X射线衍射(XRD)、电感耦合等离子体(ICP)和扫描电镜?能谱(SEM-EDS)对高硅白合金和浸出渣的物相及化学成分进行了分析对比.结果表明:在次...     

高钛渣氧化过程的动力学

用热重法研究了高钛渣在空气中的等温和非等温氧化动力学。实验结果表明，高钛渣氧化动力学可用缩核模型来描述。利用等温动力学模型公式确定等温氧化初期为化学反应控制，后期为扩散控制，计算得到相应的表观活化能分别为19．62，30．05kJ/mol通过非等温动力学模型公式确定高钛渣的非等温氧化阶段Ⅰ为界面化学反应控制，阶段Ⅱ为界面化学反应和扩散综合控制，阶段Ⅲ为扩散控制，并得到相应的表观活化能。     

微波作用下铝酸钙炉渣非等温浸出动力学

研究了微波作用下铝酸钙炉渣非等温浸出动力学，考察了不同微波辐射功率对炉渣中氧化铝浸出率的影响以及相应微波功率下反应体系温度随时间的变化规律。结果表明：该浸出动力学过程的控制步骤为界面化学反应控制，表观反应活化能为40kJ／mol左右；微波作用可改变反应的频率因子，对反应活化能影响不明显；随着微波辐射功率的增大，频率因子增加，宏观表现出微波功率的增加对反应有促进作用。     

氯化浸取钢渣中钒的动力学研究

为了提高湿法浸出低钒钢渣中钒的浸出率,并为湿法浸出低钒钢渣中钒提供理论依据,从动力学角度分析整个浸出过程,并考察温度、液/固比、浸出时间和搅拌速度对浸出过程的影响。结果表明,在90℃,液/固比为10:1以及4.0mol/L盐酸,过氧化氢8.0mL,浸取90min条件下,低钒钢渣中钒的浸出率可达到98.8%。通过正交实验和动力学推导,得到描述浸出过程的经验方程。低钒钢渣湿法浸出钒的动力学模型为未反应收缩核模型,浸出过程的表观活化能为7.21kJ/mol。该模型表明浸出过程中的控制步骤取决于边界层的扩散速度。提高温度、液/固比和浸出时间,均可增加钒的浸出速度,提高钒的浸出率。     

低钒转炉钢渣提钒湿法工艺的动力学研究

为了提高湿法浸出低钒钢渣中钒的浸出效率,并对湿法浸出低钒钢渣中钒提供理论依据,从动力学角度分析整个浸出过程。考察温度、液固比、硫酸质量分数和搅拌速率对浸出过程的影响。研究结果表明:在90℃、液固比为10?1以及硫酸浓度6.0mol/L时,浸取9h,低钒钢渣中钒的浸出率可达到95.3%。通过正交实验和动力学推导,得到描述浸出过程的经验方程,低钒钢渣湿法浸出钒的动力学模型为收缩核动力学模型,浸出过程的表观活化能为12.794kJ/mol,该模型表明浸出过程中的控制步骤取决于固膜扩散速率。提高温度、液固比和硫酸质量分数,均可加速钒的浸出速度,提高钒的浸出率。     

包头稀土精矿的配合浸出及动力学

采用氯化铝盐酸体系配合浸出包头混合稀土精矿,并对浸出过程动力学进行研究,浸出过程主要考察盐酸和氯化铝的浓度、液固比、搅拌速度、温度及反应时间对精矿浸出的影响。结果表明,随着盐酸和氯化铝的浓度和液固比的增大、反应时间的延长和反应温度的升高,精矿的浸出率逐渐增大,得到的优化浸出工艺条件如下:HCl和AlCl3浓度分别为4.0 mol/L和1.5 mol/L,液固比为20 mL/g,搅拌速度为300 r/min,温度为85℃,时间为90 min。SEM-EDS及动力学分析结果表明,精矿浸出过程符合一种受固体颗粒表面的界面交换和固膜扩散混合控制的新缩小核模型,表观活化能为35.3 kJ/mol,阿伦尼乌斯常数k0=419.95,反应级数a,b和c分别为1.265,1.208和1.22,通过计算推导出反应动力学方程。     

机械活化对从复杂硫化精矿中硫代硫酸盐浸取金的影响

采用机械活化提高复杂硫化矿CuPbZn中金的回收。研究研磨时间、球尺寸、球料比和球磨转速对金的硫代硫酸盐浸取的影响。在最优条件下（球磨时间1h,球尺寸20mm,球料比15/1,球磨速度600r/min）,矿石的非晶化度达到78%,颗粒尺寸从30μm下降到8um,比表面积从1.3m2/g增加到4.6m2/g,金的回收率从7.4%提高到73.26%。     

Microbial oxidation of refractory gold sulfide concentrate by a native consortium

A defined mesophilic consortium including an iron oxidizing bacterium and a sulfur oxidizing bacterium was constructed to evaluate its ability for bioleaching a flotation concentrate from Andacollo mine in Neuquén, Argentina. Experiments were performed in shake flasks with a pulp density of 10% (w/v), using a basal salt medium containing ferrous iron at pH 1.8. The leaching solutions were analyzed for pH, redox potential (using specifics electrodes), ferrous iron (by UV-Vis spectrophotometry) and metal concentrations (by atomic absorption spectroscopy). The results showed that the consortium was able to reduce the refractory behavior of the concentrate, allowing 91.6% of gold recovery; at the same time, high dissolution of copper and zinc was reached. These dissolutions followed a shrinking core kinetic model. According to this model, the copper solubilization was controlled by diffusion through a product layer (mainly jarosite), while zinc dissolution did not show a defined control step. This designed consortium, composed of bacterial strains with specific physiological abilities, could be useful not only to optimize gold recovery but also to decrease the leachates metallic charge, which would be an environmental advantage.     

Coupling leaching of sphalerite concentrate

Coupling process of sphalerite concentrate leaching in H2SO4-HNO3 and tetrachloroethylene extracting of sulfur was investigated. Effects of leaching temperature, leaching time, mass ratio of liquid to solid and tetrachloroethylene addition on zinc leaching processes were examined separately. SEM images of sphalerite concentrate and residues were performed by using JEM-6700F field emission scanning electron microscope. The relationship between the number of recycling and extraction ratio of zinc was studied. The results indicate that 99.6% zinc is obtained after leaching for 3 h at 85℃ and pressure of 0.1MPaO2, with 20g sphalerite concentrate in 200 mL leaching solution containing 2.0mol/L H2SO4 and 0.2mol/L HNO3, in the presence of 10 mL C2Cl4. The leaching time of zinc is 50% shorter than that in the common leaching. The coupling effect is distinct. The recycled C2Cl4 exerts little influence on extraction ratio of zinc.     

Kinetics of zinc sulfide concentrate direct leaching in pilot plant scale and development of semi-empirical model

The direct leaching kinetics of an iron-poor zinc sulfide concentrate in the tubular reactor was examined. All tests were carried out in the pilot plant. To allow the execution of hydrostatic pressure condition, the slurry with ferrous sulfate and sulfuric acid solution was filled into a vertical tube (9 m in height) and air was blown from the bottom of the reactor. The effects of initial acid concentration, temperature, particle size, initial zinc sulfate concentration, pulp density and the concentration of Fe on the leaching kinetics were investigated. Results of the kinetic analysis indicate that direct leaching of zinc sulfide concentrate follows shrinking core model (SCM). This process was controlled by a chemical reaction with the apparent activation energy of 49.7 kJ/mol. Furthermore, a semi-empirical equation is obtained, showing that the order of the iron, sulfuric acid and zinc sulfate concentrations and particle radius are 0.982, 0.189, ?0.097 and ?0.992, respectively. Analysis of the unreacted and reacted sulfide particles by SEM–EDS shows that insensitive agitation in the reactor causes detachment of the sulfur layer from the particles surface in lower than 60% Zn conversion and lixiviant in the face with sphalerite particles.     

硫化锌精矿常压富氧直接浸出行为

借助工艺矿物学分析对常压富氧直接浸出条件下锌精矿中主要硫化物的浸出行为进行研究。结果表明,除黄铁矿外,其他硫化矿均会明显溶解。基于对浸出渣中单质硫与反应残余硫化物之间关系的分析,认为闪锌矿、黄铜矿、铜蓝、方铅矿的溶出可能遵循间接氧化方式,即硫化物首先酸溶,生成的H2S脱离矿物表面并迁移至溶液本体中进而氧化成单质硫。上述硫化矿的浸出过程可能受界面化学反应控制。对于磁黄铁矿的溶出,直接电化学氧化可能起主导作用,其浸出过程可能受产物层单质硫的扩散控制。     

Accelerating column leaching trial on copper sulfide ore

The main measures to accelerate leaching sulfide ore are large spraying intensity,manual oxygen supply,temperature control,and acclimated bacteria.The indoor experiment accelerating sulfide ore leaching detected the temperature during leaching process,dissolvability of oxygen,bacterial concentration,Cu concentration and slag grade.At the same time,this paper also analyzed the effect of four factors,which are bacterial diversity cultivation stage,spraying intensity,air supply,and whether to control temperature,on the leaching efficiency of copper.The results indicate that the oxygen content of leach solution has a close relationship with temperature but it is rarely affected by air supply.The bacterial concentration preserves from 106 to 107 mL^-1,and temperature has a great effect on the bacterial activity under the condition of proper temperature and oxygen supply,and the lack of nutrition prevents the bacterial concentration from rising in the late stage.The relationships of the copper leaching efficiency to temperature,air feed,and spraying intensity are directly proportional.The leaching efficiencies of the cultivated bacteria and acclimation bacteria are 1.2 and 1.4 times as large as that of the original bacteria.     

Refractory concentrate gold leaching: Cyanide vs. bromine

Gold extraction, recovery and economics for two refractory concentrates were investigated using cyanide and bromine reagents. Gold extractions for cyanide leaching (24–48 hours) and bromine leaching (six hours) were the same and ranged from 94 to 96%. Gold recoveries from bromine pregnant solutions using carbon adsorption, ion exchange, solvent extraction, and zinc and aluminum precipitation methods were better than 99.9%. A preliminary economic analysis indicates that chemical costs for cyanidation and bromine process are $11.70 and $11.60 respectively, per tonne of calcine processed.