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。     

低品位铂钯精矿的富氧压浸出

对高镁低品位复杂铂钯精矿进行工艺矿物学分析,提出采用硫酸氧压浸出工艺对该精矿中的贱金属铜、镍、铁选择性浸出分离并富集铂钯的处理工艺。考察磨矿粒度、反应温度、时间、初始硫酸浓度、氧压、搅拌速度、木质素磺酸钙用量、液固比对铜、镍、铁浸出率及渣率的影响,确定最佳工艺参数。实验结果表明:当精矿粒度小于43μm占有率为93%、时间3 h、浸出温度150℃、初始硫酸浓度2 mol/L、氧分压0.7 MPa、搅拌速度400 r/min、添加剂木质素磺酸钙用量0.6 g、液固比5:1的最佳工艺条件下,铜浸出率达99.27%、镍浸出率达98.04%、渣率为37%左右,铂钯几乎不被浸出,铂和钯在浸出渣中富集近3倍。     

某难选金矿加温化学预氧化浸出技术

以某实际含铜金矿为研究对象,在氯盐酸性加温体系下,分析浸出温度、时间、矿物粒度、NaCl浓度、H2SO4浓度、氧气流量等因素对化学预氧化浸出除铜和浸出渣氰化浸金的影响过程.结果表明:在90%矿样粒度小于37 μm、浸出温度95 ℃、初始H2SO4浓度0.75 mol/L、起始NaCl浓度0.7 mol/L、液固比5-1、浸出时间24 h、搅拌速度750 r/min的条件下,可使铜的浸出去除率达到80%以上,预氧化渣金的氰化浸出率达98.23%.     

加压浸出低品位锰矿的工艺

对低品位锰矿硫酸加压浸出工艺进行正交实验和单因素实验,通过正交实验得出:加压浸出低品位锰矿工艺中,影响锰浸出率的主要因素的较佳组合如下:初始酸浓度160 g/L、硫铁矿量50 g、液固比5:1(5 mL/g)、压力1 MPa、温度180℃、时间80 min。为分析低品位锰矿中锰、铁及铝的浸出行为,实现这3种金属元素的高效分离,参考正交实验结果适当地调整工艺参数,进行单因素实验研究,详细考察始酸浓度、反应温度、硫铁矿量、液固比、浸出时间和浸出压力对锰、铁及铝浸出率的影响,得到优化浸出工艺条件如下:低品位锰矿粉100 g,初始硫酸浓度120 g/L,浸出反应温度120℃,硫铁矿量50 g,液固比5:1(5 mL/g),浸出时间100 min,浸出压力0.7 MPa,搅拌转速500 r/min。本工艺具有良好的稳定性,在优化浸出条件下,锰的浸出率为96%,而铝和铁的浸出率分别为38.7%和7.12%,实现锰选择性高效溶出,锰和铝、铁等杂质的分离效果良好,为最终实现低品位锰矿中各种有价元素的清洁高效回收奠定了基础。     

采用碱浸预脱硅-氰化工艺从氰化尾渣中回收金

采用高浓度碱浸对氰化尾渣进行预脱硅处理,考察搅拌速度、固液比、Na OH浓度及温度对硅浸出速率的影响,研究脱硅过程的反应动力学,得到相应的动力学方程。结果表明:当搅拌速度为400 r/min、固液比为1:5、Na OH浓度为80%、反应温度为280℃时,二氧化硅的浸出率为91.8%;碱浸过程受产物层内扩散控制,表观反应活化能为37.375 k J/mol。通过正交实验对氰化浸金的条件进行了优化,在Si O2浸出率为91.8%,Na CN浓度为1.5 g/L,固液比为1:3,浸出时间为48 h的条件下,金的浸出率为87.83%。     

低品位氧化锌矿在MACA体系中的循环浸出

研究Me(Ⅱ)-NH4Cl-NH3-H2O(MACA)体系处理兰坪低品位氧化锌矿的浸出过程,提出用循环浸出方法富集浸出液中锌浓度的工艺技术方案.结果表明:循环浸出的优化条件为液固比4:1、常温、搅拌速度300 r/min、浸出时间3 h;通过循环次数分别为14、15及10的3个阶段的循环浸出试验,获得渣计锌浸出率≥69...     

铜冶炼烟尘的控电位氧化浸出（英文）

采用氧化浸出和电位控制技术从铜冶炼烟尘中浸出金属,研究H2O2用量、H2O2加入速度、初始盐酸浓度、浸出温度、初始液固比和浸出时间对金属浸出率的影响。最终得到最优浸出条件为:H2O2用量0.8mL/g(氧化还原电位为429 mV)、H2O2加入速度1.0 mL/min、初始硫酸浓度1.0 mol/L、初始盐酸浓度1.0 mol/L、浸出温度80°C、初始液固比5:1 mL/g以及浸出时间1.5 h。在此最优条件下,铜冶炼烟尘中的铜和砷能被有效地浸出,剩下的浸出渣可作为一种合适的铅冶炼资源。此时,铜、砷和铁的平均浸出率分别为95.27%、96.82%和46.65%。     

低品位软锰矿碱浸预脱硅-流化焙烧制备锰酸钠（英文）

采用低浓度碱浸对低品位软锰矿进行预脱硅处理,考察NaOH浓度、液固比、浸出温度、浸出时间及搅拌速率对硅浸出率的影响,研究碱浸过程动力学。结果表明:在NaOH起始浓度为20%、液固比为4:1、浸出温度为180°C、浸出时间为4h、搅拌速率为300r/min的条件下,硅浸出率达到91.2%。缩核模型表明,碱浸过程受化学表面反应控制,其表观反应活化能为53.31kJ/mol。通过正交试验对脱硅渣流化焙烧制备锰酸钠的条件进行优化,在硅浸出率为91.2%、NaOH/MnO_2质量比为3:1、焙烧温度为500°C、焙烧时间为4h的条件下,锰酸钠的转化率为89.7%,且锰酸钠转化率随硅浸出率的升高而增加。     

从熔炉渣中氯盐浸出铅的参数优化和动力学（英文）

研究采用氯盐浸出法从传统熔铅炉渣中回收铅的可行性和动力学。考察工艺参数如浸出时间、NaCl浓度、FeCl_3浓度、液/固比、搅拌速度、浸出温度、颗粒尺寸对铅回收率的影响。基于中心复合设计模型,利用响应面法对上述参数进行优化,得到优化工艺条件如下:浸出时间60 min、浸出温度80℃、搅拌速度800 r/min、NaCl浓度200 g/L、FeCl_3浓度80 g/L、液固比16、颗粒尺寸小于106μm。在此优化条件下96%Pb被回收。基于方差分析法,确定浸出温度、液固比和NaCl浓度为影响浸取过程的最有效参数。动力学研究结果表明,方铅石的氯盐浸出过程为一级反应过程。反应机理为固态反应产物的扩散和化学反应。采用Arrhenius模型计算得到从方解石中氯盐浸出铅的激活能为27.9 kJ/mol。     

以硫化钙为还原剂还原-酸浸提取低品位氧化锰矿中的锰元素（英文）

以CaSO4制备得到的CaS为还原剂,研究氧化锰矿的还原-酸浸过程,考察硫化钙与矿石的质量比、还原温度、还原时间、液固比、搅拌速率、浸出温度、浸出时间和H2SO4浓度对氧化锰矿中锰及铁浸出率的影响。结果表明:优化的还原工艺条件为硫化钙与矿石质量比1:6.7、液固比5:1、搅拌速率300 r/min、还原温度95°C、还原时间2.0 h;酸浸工艺条件为搅拌速率200 r/min、H2SO4浓度1.5 mol/L、浸出温度80°C、浸出时间5 min。在此优化条件下,锰的浸出率达到96.47%,而铁的浸出率仅为19.24%。该工艺可以应用于不同类型氧化锰矿中锰的提取,且锰的浸出率均高于95%。     

Preparation of sodium manganate from low-grade pyrolusite by alkaline predesilication–fluidized roasting technique

Low concentration alkaline leaching was used for predesilication treatment of low-grade pyrolusite. The effects of initial NaOH concentration, liquid-to-solid ratio, leaching temperature, leaching time and stirring speed on silica leaching rate were investigated and the kinetics of alkaline leaching process was studied. The results show that silica leaching rate reached 91.2% under the conditions of initial NaOH concentration of 20%, liquid-to-solid ratio of 4:1, leaching temperature of 180 °C, leaching time of 4 h and stirring speed of 300 r/min. Shrinking-core model showed that the leaching process was controlled by the chemical surface reaction with activation energy E_a of 53.31 kJ/mol. The fluidized roasting conditions for preparation of sodium manganate were optimized by the orthogonal experiments using the desiliconized residue. The conversion rate of sodium manganate was obtained to be 89.7% under the conditions of silica leaching rate of 91.2%, NaOH/MnO₂ mass ratio of 3:1, roasting temperature of 500 °C and roasting time of 4 h, and it increased with the increase of silicon leaching rate.     

Dissolution kinetics of low grade complex copper ore in ammonia-ammonium chloride solution

The leaching kinetics of Tang-dan refractory low grade complex copper ore was investigated in ammonia-ammonium chloride solution. The concentration of ammonia and ammonium chloride, the ore particle size, the solid-to-liquid ratio and the temperature were chosen as parameters in the experiments. The results show that temperature, concentration of ammonia and ammonium chloride have favorable influence on the leaching rate of copper oxide ores. But, leaching rate decreases with increasing particle size and solid-to-liquid ratio. The leaching process is controlled by the diffusion of the lixiviant and the activation energy is determined to be 23.279 kJ/mol. An equation was also proposed to describe the leaching kinetics.     

Oxidation leaching of copper smelting dust by controlling potential

A study was conducted for metal extraction from copper smelting dust using the oxidation leaching and control of potential technology. The effects of H₂O₂ dosage, H₂O₂ feeding speed, initial HCl concentration, leaching temperature, liquid-to-solid ratio and leaching time on metals leaching efficiencies were investigated. The following optimized leaching conditions were obtained: H₂O₂ dosage of 0.8 mL/g (redox potential of 429 mV), H₂O₂ feeding speed of 1.0 mL/min, initial H₂SO₄ concentration of 1.0 mol/L, initial HCl concentration of 1.0 mol/L, leaching temperature of 80 °C, initial liquid-to-solid ratio of 5:1 mL/g and leaching time of 1.5 h. Under the optimized conditions, copper and arsenic can be effectively leached from copper smelting dust, leaving residue as a suitable lead resource. The average leaching efficiencies of copper, arsenic and iron are 95.27%, 96.82% and 46.65%, respectively.     

Leaching performance of Al-bearing spent LiFePO4 cathode powder in H2SO4 aqueous solution

The leaching performance and leaching kinetics of LiFePO₄ (LFP) and Al in Al-bearing spent LFP cathode powder were systematically studied. The effects of temperature (273?368 K), stirring speed (200?950 r/min), reaction time (0?240 min), acid-to-material ratio (0.1:1?1:1 mL/g) and liquid-to-solid ratio (3:1?9:1 mL/g) on the leaching process were investigated. The results show that the concentration of reactants and the temperature have a greater impact on the leaching of Al. Under the optimal conditions, leaching efficiencies of LFP and Al are 91.53% and 15.98%, respectively. The kinetic study shows that the leaching of LFP is kinetically controlled by mixed surface reaction and diffusion, with an activation energy of 22.990 kJ/mol; whereas the leaching of Al is only controlled by surface chemical reaction, with an activation energy of 46.581 kJ/mol. A low leaching temperature can effectively suppress the dissolving of Al during the acid leaching of the spent LFP cathode material.     

用碱性Na_2EDTA溶液从次氧化锌烟灰中选择性回收铅(英文)

采用碱性Na2EDTA溶液从次氧化锌烟灰中回收铅。探讨温度、浸出时间、Na2EDTA浓度和起始NaOH浓度对铅、锌浸出率的影响。得到最优实验条件如下:液固比5:1 mL/g、搅拌速度650 r/min、Na2EDTA浓度0.12mol/L、NaOH初始浓度0.5 mol/L、温度70°C、浸出时间120 min。在最优实验条件下,铅、锌、氟和氯的平均浸出率分别为89.92%、0.94%、62.84%和90.02%。浸出液用于电沉积铅粉。在温度为60°C、电流密度为200A/m2、H3PO4浓度为1.5 g/L、铅离子浓度不低于5 g/L时,电沉积铅粉平均电流效率大约为93%,阴极铅纯度高于98%。电沉积1 kg铅粉大约消耗0.218 kg Na2EDTA和0.958 kW·h电能。     

Reduction treatment of low-grade manganese ore by biomass roasting

The reduction of manganese dioxide in low-grade manganese ore by biomass roasting process was investigated for extracting manganese from poor manganese ore more effectively. In this study,the cinder of ore fines and sawdust was further leached by sulphuric acid to obtain MnSO4. Over 97% manganese in ores can be converted into MnSO4. Effects of the mass ratio of manganese ore to sawdust, roasting temperature and time, leaching temperature and time, leaching agent concentration and liquid-solid ratio were studied. The manganese recovery achieved 97.71% under the conditions: the mass ratio of manganese ore to sawdust of 5:1, roasting temperature 500℃ for 40min, leaching temperature 60℃ for 40min, sulphuric acid concentration of 1mol/L and liquid-solid ratio of 10:1. This technology can be suitable for extraction of Mn in low-grade manganese ore.     

Leaching of chromite ore in concentrated KOH by catalytic oxidation using CuO as catalyst

CuO was used as a catalyst in the concentrated KOH solution to enhance the leaching of chromium from the chromite ore. The impacts of temperature, KOH-to-chromite ore mass ratio, CuO-to-chromite ore mass ratio, and gas flow rate on the chromium leaching rate were investigated. The results indicated that CuO played an important role in improving the chromium leaching rate. The leaching rate reached 98% after leaching for 6 h when CuO was applied, whereas it was only 60.8% without CuO under the same reaction conditions: temperature 230 °C, KOH-to-ore mass ratio 6:1, stirring speed 700 r/min, gas flow rate 1 L/min. According to the kinetics study, the catalytic oxidation was controlled by surface chemical reaction and the activation energy was calculated to be 15.79 kJ/mol when the temperature was above 230 °C. In contrast, without CuO, the rate-determining step was external diffusion and the apparent activation energy was 38.01 kJ/mol.     

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...     

Pressure leaching of converter vanadium slag with waste titanium dioxide

The process of pressure leaching the converter vanadium slag with waste titanium dioxide without roasting was studied. Mineralogy analysis indicates that the converter vanadium slag contains mainly three mineral phases: magnetite, titanium magnetite, and silicate phase. Vanadium is in combination with iron, titanium, manganese, aluminum, and silicon. The impacts of leaching temperature, leaching time, stirring speed, liquid-to-solid ratio, and initial leaching agent concentration were investigated on the waste titanium dioxide leaching process. The results indicate that under the optimal conditions, the vanadium and the iron leaching rates are 96.85 % and 93.50 %, respectively, and the content of titanium is 12.6 % in the residue. The main mineral phases for the residues under the optimal operation conditions are quartz, ilmenite, anatase, and silicate phase, and the residues can be reused as the extraction of titanium raw materials for titanium dioxide production technology by the sulfate method.