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。     

低品位氧化铜矿氨-硫酸铵体系过硫酸铵氧化浸出

以过硫酸铵为氧化剂,研究低品位氧化铜矿在氨-硫酸铵体系氧化浸出工艺。讨论氨/铵离子摩尔比、总氨浓度,氨、硫酸铵和过硫酸铵的浓度,反应温度,液固比,反应时间和搅拌速度等操作条件对铜浸出的影响。结果表明:在92.8%的矿样粒径小于0.045 mm,氨、硫酸铵和过硫酸铵浓度分别为2.4、1.8和0.100 mol/L,浸出时间为90 min,温度为30℃,液固比(mL/g)为5:1,搅拌速度为500 r/min时的优化条件下,低品位铜矿的铜浸出率达87.7%。     

Kinetics of simultaneous leaching of Ag and Pb from hydrometallurgical zinc residues by chloride

The leaching kinetics of silver and lead simultaneously from zinc residue by chloride was investigated.The effects of stirring speed,temperature,sodium chloride concentration,particle size and liquid/solid ratio on Ag and Pb dissolution in sodium chloride were studied.It was determined that the dissolution rates increased with increasing sodium chloride concentration,temperature and decreasing particle size.The dissolution kinetics followed a shrinking core model,with inter-diffusion through gangue layer as the rate determining step.This finding is in accordance with the apparent activation energy(Ea)of 26.8 kJ.mol-1(Ag)and 26.5 kJ.mo1-1(Pb),and a linear relationship between the rate constant and the reciprocal of squared particle size.The orders of reaction with respect to sodium chloride concentration,temperature and particle size were also achieved.The rate of reaction based on diffusion-controlled process can be expressed by semi-empirical equations.     

焙烧氟碳铈矿硫酸浸出稀土的动力学(英文)

研究了硫酸浸出德昌稀土与天青石共伴生矿的焙烧矿过程。考查粒度、搅拌速度、硫酸浓度和温度对稀土浸出率的影响,并对稀土的浸出动力学进行分析。在选定的浸出条件下:粒径0.074～0.100mm、硫酸浓度1.5mol/L、液固比8:1、搅拌速度500r/min,稀土浸出反应受内扩散控制,表观活化能为9.977kJ/mol。     

Optimization of operating parameters and kinetics for chloride leaching of lead from melting furnace slag

The feasibility and kinetics of lead recovery from the slag of traditional lead melting furnace using chloride leaching were investigated. The effects of operating parameters such as leaching time, NaCl concentration, FeCl₃ concentration, liquid/solid ratio, stirring rate, temperature, and particle size on recovery of lead were studied and the optimization was done through the response surface methodology (RSM) based on central composite design (CCD) model. The optimum conditions were achieved as follows: leaching time 60 min, 80 °C, stirring rate 800 r/min, NaCl concentration 200 g/L, FeCl₃ concentration 80 g/L, liquid/solid ratio 16, and particle size less than 106 μm. More than 96% of lead was effectively recovered in optimum condition. Based on analysis of variance, the reaction temperature, liquid/solid ratio, and NaCl concentration were determined as the most effective parameters on leaching process, respectively. Kinetics study revealed that chloride leaching of galena is a first-order reaction and the diffusion through solid reaction product and chemical reaction control the mechanism. The activation energy of chloride leaching of galena was determined using Arrhenius model as 27.9 kJ/mol.     

Kinetic process of oxidative leaching of chalcopyrite under low oxygen pressure and low temperature

Kinetic process of oxidative leaching of chalcopyrite in chloride acid hydroxide medium under oxygen pressure and low temperature was investigated. The effect on leaching rate of chalcopyrite caused by these factors such as ore granularity, vitriol concentration, sodium chloride concentration, oxygen pressure and temperature was discussed. The results show that the leaching rate of chalcopyrite increases with decreasing the ore granularity. At the early stage of oxidative reaction, the copper leaching rate increases with increasing the oxygen pressure and dosage of vitriol concentration, while oxygen pressure affects leaching less at the later stage. At low temperature, the earlier oxidative leaching process of chalcopyrite is controlled by chemical reactions while the later one by diffusion. The chalcopyrite oxidative leaching rate has close relation with ion concentration in the leaching solution. The higher ion concentration is propitious for chalcopyrite leaching.     

硝酸溶液中氧化铜矿浸出工艺参数优化及动力学模拟

研究Tunceli孔雀石矿物在硝酸溶液中的溶出行为,以评估各种实验参数的影响.研究为分两个阶段.在第一步中,确定浸出过程的最佳条件,而在第二步中,对该过程进行动力学评估.在优化实验中,以硝酸浓度、温度、搅拌速度和固液比为自变量,采用中心组合设计法(CCD)获得实验数据.确定硝酸浓度、温度、固液比和搅拌速度的最佳值分别为...     

Na_2SO_3从硒碲富集物中浸出硒动力学(英文)

研究用Na2SO3溶液从硒碲富集物中浸出硒的动力学。该硒碲富集物的微观形貌主要为球状体和柱状体,其粒径范围为17.77～56.58μm,且其主要成分为41.73%Se和40.96%Te。研究Na2SO3浓度(126～315 g/L)、搅拌速度(100～400 r/min)、反应温度(23～95°C)、液固比(7:1～14:1)及硒碲富集物平均粒径(17.77～56.58μm)对Se浸出率的影响。结果表明:增加Na2SO3浓度、提高搅拌速度、升高反应温度和增加液固比均可以提高Se浸出率,而增大硒碲富集物的粒径会导致Se的浸出率降低;反应温度对Se浸出率影响较大,当反应温度从23°C升高至95°C时,Se浸出率从21%增至67%;该浸出过程符合Avrami模型,其模型特征参数和表观活化能分别为0.235和20.847kJ/mol。     

氟碳铈矿盐酸浸出过程的动力学

考察盐酸浓度、浸出温度、液固比和矿物粒度对浸出速度的影响.结果表明:在盐酸浓度6 mol/L、浸出温度90C、液固比15:1、矿物粒度25 μm的条件下浸出90 min后,氟碳铈矿中稀土碳酸盐的浸出率达到89.6％,而稀土氟化物的浸出率仪为1.5％.氟碳铈矿浸出过程符合产物层扩散模型,稀土碳酸盐和氟化物浸出过程的表观活...     

石煤钒矿硫酸活化常压浸出提钒工艺

研究石煤钒矿的硫酸活化提钒方法。分别考察矿石粒度、硫酸浓度、活化剂用量、催化剂用量、反应温度、反应时间和浸出液固比等因素对钒浸出率的影响。结果表明:石煤提钒的优化条件为矿石粒度小于74μm的占80%、硫酸浓度150 g/L、活化剂CaF2用量(相对于矿石)60 kg/t、催化剂R用量20 g/L、反应温度90℃、反应时间6 h、液固比(体积/质量,mL/g)2:1,在此优化条件下,钒浸出率可达94%以上;在优化条件下,采用两段逆流浸出,可有效减少活化剂CaF2以及浸出剂硫酸的消耗量;经过两段逆流浸出萃取反萃氧化水解工艺,全流程钒资源总回收率可达86.9%;V2O5产品纯度高于99.5%。     

Leaching kinetics of antimony-bearing complex sulfides ore in hydrochloric acid solution with ozone

The leaching kinetics of Sb and Fe from antimony-bearing complex sulfides ore was investigated in HCl solution by oxidation–leaching with ozone. The effects of temperature, HCl concentration, stirring speed and particle size on the process were explored. It is found that the recoveries of Sb and Fe reach 86.1% and 28.8%, respectively, when the reaction conditions are 4.0 mol/L HCl, 900 r/min stirring speed at 85 °C with <0.074 mm particle size after 50 min leaching. XRD analysis indicates that no new solid product forms in the leaching residue and the leaching process can be described by shrinking core model. The leaching of Sb corresponds to diffusion-controlled model at low temperature (15–45 °C) and mixed-controlled model at high temperature (45–85 °C), and the apparent activation energies are 6.91 and 17.93 kJ/mol, respectively. The leaching of Fe corresponds to diffusion-controlled model, and the apparent activation energy is 1.99 kJ/mol. Three semi-empirical rate equations are obtained to describe the leaching process.     

液相球磨法由氧化铋制备次碳酸铋的动力学

为解决传统次碳酸铋生产过程中成本高和环境污染问题,提出采用液相球磨法由氧化铋制备次碳酸铋的新工艺,研究用碳酸氢氨和氧化铋制备次碳酸铋反应过程的动力学,考察反应温度、氧化铋粒度、液固比以及碳酸氢氨浓度对氧化铋转化率的影响。结果表明,在9-30°C的范围内,升高反应温度、减小氧化铋粉末的粒度、扩大液固比以及提高碳酸氢氨浓度均有利于氧化铋转化率的提高。对反应产物的表征分析表明,产品的纯度较高、杂质少;SEM结果显示产品次碳酸铋主要呈针棒状形态。反应过程受产物层的扩散控制,可用未反应收缩核模型描述,反应的表观活化能为9.783 kJ/mol,同时获得了描述反应过程的半经验动力学方程。     

Kinetics of indium dissolution from marmatite with high indium content in pressure acid leaching

The kinetics of oxygen pressure acid leaching marmatite with high indium content was studied. The effects of particle size, agitation rate, temperature, H_2SO_4 concentration, and oxygen partial pressure on leaching rate of indium were investigated. The results show that when the agitation rate is above 600 r-min~(-1), its influence on indium leaching rate is insignificant. It is determined that the leaching rates increase with the increase in sulfuric acid concentration, temperature, partial oxygen pressure, and the decrease in particle size. Moreover, the results demonstrate that the process of indium leaching is controlled by interface chemical reaction, with apparent activation energy of 65.7 k J-mol~(-1). The apparent reaction orders of sulfuric acid and oxygen partial pressure are determined to be 0.749 and 1.260, respectively. The leaching reaction process follows shrinking unreacted core model. And finally, the kinetics model equation is established for indium.     

Co-extraction of valuable metals and kinetics analysis in chlorination process of low-grade nickel-copper sulfide ore

To efficiently co-extract Ni and Cu from low-grade nickel-copper sulfide ore, chlorination roasting with NH₄Cl followed by a water leaching process was investigated. The results show that 98.4% Ni and 98.5% Cu can be synchronously extracted when the ore particle size is 75-80 μm, the roasting time is 2 h, the mass ratio of NH₄Cl to ore is 1.6:1 and the roasting temperature is 550 °C. The evolution behavior of various minerals was elucidated using X-ray diffraction (XRD) coupled with scanning electron microscopy (SEM). The kinetics of the chlorination process based on the differential thermal and thermogravimetric analysis (DTA-TG) data was analyzed by Kissinger method and Flynn-Wall-Ozawa (FWO) method. The chlorination process of low-grade nickel-copper sulfide ore mainly contains two stages: the decomposition of NH₄Cl and the chlorination of ore. The maximum apparent activation energies (E_a) at two stages are determined to be 114.8 and 144.6 kJ/mol, respectively. The condensed product of exhaust gas is determined to be ammonium chloride, which can be recycled as the reactant again, making the process economic and clean.     

Leaching of cuprite through NH4OH in basic systems

Cuprite is a difficult oxide to leach under acidic conditions (for the maximum extraction of 50%). In this research, the feasibility of leaching cuprite in an ammoniacal medium was studied. The working conditions addressed here were the liquid/solid ratio (120:1–400:1 mL/g), stirring speed (0–950 r/min), temperature (10–45 °C) and NH₄OH concentration (0.05–0.15 mol/L). In addition, different ammoniacal reagents (NH₄F and (NH₄)₂SO₄) were analyzed. The experiments were performed in a 2 L reactor with a heating mantle and a condenser. The most important results were that the maximum leaching rate was obtained at pH 10.5, 0.10 mol/L NH₄OH, 45 °C, 4 h, 850 r/min and a liquid/solid ratio of 400:1, reaching a copper extraction rate of 82%. This result was related to the non-precipitation of copper in solution by the formation of copper tetra-amine. The liquid/solid ratio and stirring speed were essential for increasing the cuprite leaching. The maximum leaching rate was achieved at higher temperatures; however, significant copper leaching rate occurred at temperatures near the freezing point of water (17.9% over 4 h). Increasing NH₄OH concentration and decreasing particle size increased the cuprite leaching rate. The two ammoniacal reagents (NH₄F and (NH₄)₂SO₄) had low extraction rate of copper compared with NH₄OH. The kinetic model representing cuprite leaching was a chemical reaction on the surface. The order of the reaction with respect to the NH₄OH concentration was 1.8, and it was inversely proportional to the radius of the ore particles. The calculated activation energy was 44.36 kJ/mol in the temperature range of 10–45 °C.     

四川某低品位微细粒氧化金矿非氰浸出试验研究

对金品位为2.02 g/t的某低品位氧化微细粒金矿开展了全泥浸出提取金的试验研究。优选出非氰浸出剂CC-1,确定了相应工艺参数,在此基础上开展了3个粒级柱浸试验,对柱浸含金溶液进行了活性炭吸附试验,研究表明该矿石适宜于利用非氰浸出剂CC-1堆浸回收金。矿石磨至-200目占80%、矿浆液固比2:1、石灰用量3000 g/t原矿、CC-1浓度0.10%、浸出时间30 h条件下金浸出率92.75%;在石灰用量3000 g/t、CC-1浓度0.10%、浸出时间10 d时-10 mm矿样Au浸出率92.46%,浸出时间15 d时-20 mm及-30 mm矿样Au浸出率分别为91.49%、89.24%。采用CC-1作为浸出剂的含Au溶液活性炭吸附率为95.72%~97.11%。     

Recovery prediction of copper oxide ore column leaching by hybrid neural genetic algorithm

The artificial neural network (ANN) and hybrid of artificial neural network and genetic algorithm (GANN) were applied to predict the optimized conditions of column leaching of copper oxide ore with relations of input and output data. The leaching experiments were performed in three columns with the heights of 2, 4 and 6 m and in particle size of <25.4 and <50.8 mm. The effects of different operating parameters such as column height, particle size, acid flow rate and leaching time were studied to optimize the conditions to achieve the maximum recovery of copper using column leaching in pilot scale. It was found that the recovery increased with increasing the acid flow rate and leaching time and decreasing particle size and column height. The efficiency of GANN and ANN algorithms was compared with each other. The results showed that GANN is more efficient than ANN in predicting copper recovery. The proposed model can be used to predict the Cu recovery with a reasonable error.     

在DMF中用氯化铜氧化浸出钯的研究

在N,N-二甲基甲酰胺(DMF)为溶剂的无水体系中,用氯化铜(CuCl2)氧化浸出钯;以单因素试验及响应面优化法进行了浸出条件优化研究。结果表明,钯在浸出液中以Pd(II)形态存在;钯的浸出率随CuCl2浓度升高而增加,随温度升高而增加,随浸出时间增加而增加,随固液比增大而增加。在CuCl2浓度为1.25 mol/L,固液比为1:700 g/mL,80℃浸出90 min的优化条件下,钯的浸出率可以达到98.5%。浸出液中加入丁二酮肟溶液后,96.3%的钯和5%的铜能够沉淀下来,方法可为电子废料中钯的回收提供参考。     

湿法炼锌净化渣酸浸的动力学

对湿法炼锌净化渣的浸出动力学进行了研究,并探讨了硫酸浓度、反应温度、粒度等对钴、锌浸出率的影响规律。从动力学的角度分析了整个浸出过程,得到优化条件：液固比50：1(mL/g),硫酸浓度100 g/L,反应温度70°C,粒度75~80μm,反应时间20 min。在此优化条件下钴的浸出率为99.8%,锌的浸出率为91.97%。结果表明：在硫酸体系中钴的浸出符合不生成固体产物层的“未反应收缩核”模型。通过 Arrhenius 经验公式求得钴和锌表观反应活化能分别为11.693 kJ/mol和6.6894 kJ/mol,这表明浸出过程受边界层扩散控制。     

提高陕西某黄金冶炼厂金浸出率的研究与实践

为提高某矿山氰化浸金率,降低尾渣金品位,提高矿粉日处理量,考察了球磨时间、液固比、浸出时间、氰化钠浓度对金浸出率的影响。结果表明,以NaOH作保护碱,在NaCN浓度0.25%,液固比1.5：1,浸出时间48 h,矿粉粒度-325目占94.17%的条件下,金的浸出率可达97.80%,尾渣金品位小于1.5 g/t。根据实验结果,对现有生产工艺流程和设备进行改造,通过改变加料方式,调整球磨机转速,增加磨矿次数、分级次数和浸出槽数目,各项生产技术指标得到明显提升。