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。     

从铜合金渣中选择性回收锡及其热处理合成SnO2（英文）

采用选择性溶解方法对铜合金废渣中的锡进行富集和分离。使用HNO3溶液溶解铜合金废渣,渣中的锡不会溶解从而使其从中得到分离。通过离心分离生成的H2SnO3沉淀相,然后经过干燥加热生成SnO2粉末。对SnO2粉末的相组成、表面形貌和化学成分进行分析。     

Na2CO3-Na2SO4复盐熔炼法从碱溶渣中高效分离钨的工艺研究

钨合金废料资源化利用过程中容易产生含钨碱溶渣,且该部分含钨碱溶渣中钨品位较高。本文开发了一种复盐(Na₂CO₃-Na₂SO₄)熔炼工艺,以对含钨碱溶渣中的钨进行高效分离;探究Na₂CO₃和Na₂SO₄的添加量、复盐熔炼温度、熔炼时间以及水浸液固比、水浸温度对钨回收率的影响。结果表明：从含钨碱溶渣中高效分离钨的最优条件为n(W)∶n(Na₂CO₃)∶n(Na₂SO₄)=1∶1.25∶0.54,复盐熔炼温度为800℃,熔炼时间为3 h,水浸液固比为2.5,水浸温度为75℃,在该最优条件下可将含钨碱溶渣中99.93%的钨分离出来。同时,本文通过XRD分析以及热力学分析对复盐熔炼的反应机理进行了探讨,复盐体系的构建有助于降低体系共熔点,降低能耗,同时有助于促进碱溶渣与反应介质的充分接触,提高反应效率。     

微波场中FeCl3溶液浸出闪锌矿动力学

该文研究微波场中 FeCl_3溶液常压浸出闪锌矿动力学。考查了微波场中温度、FeCl_3浓度及粒度对 Zn 浸出率的影响,得到了非恒温动力学方程,并且证明微波辐照加热方式较传统加热方式的Zn 浸出速率快。     

Na2S选择性浸出铜转炉烟灰中的砷及Fe2(SO4)3稳定化砷(英文)

研究砷(As₂O₃)的Na₂S碱性浸出,以及用Fe₂(SO₄)₃沉淀砷。采用基于中心组合设计的响应面法对相关因素的影响进行定量和定性分析,提出用于参数优化的统计模型。结果表明,在Na₂S浓度100 g/L,固液比0.163 g/mL和温度80℃的最优预测条件下,89%的砷可从烟灰中去除。研究发现,固液比和Na₂S浓度是影响浸出过程的显著因素。在沉淀过程中,当pH为4.8、Fe³⁺与砷的摩尔比和H₂O₂与砷的摩尔比分别为5:1和4:1时,浸出液中99.93%以上的砷以无定形砷酸铁的形式被除去。Fe³⁺与砷的摩尔比和pH值为最显著因素,而且他们之间的相互作用也是显著的。     

铟的还原浸出动力学及分步沉淀选择性分离（英文）

开展硫化锌精矿还原浸出高铁锌浸出渣高效浸铟及浸出液中铟选择性分离的研究。结果表明：在固体物料粒度74～105μm、反应温度90℃、浸出时间300 min、硫酸浓度1.4 mol/L的条件下,铟的浸出率达95%以上。采用收缩核模型对还原浸出动力学进行分析,不同条件下的浸出实验结果表明反应受穿过固体产物层的扩散控制,活化能为17.96 k J/mol,相对于硫酸浓度的反应级数为2.41。铁粉置换沉铜过程铜和砷的沉淀率均达99%以上。98%以上的铟从含高亚铁离子浓度的硫酸锌溶液中选择性分离,获得铟含量约为2.4%的富铟渣,经酸浸-萃取-电积工艺流程进一步处理后可得到纯铟。     

氧化硼对保护渣中TiO2溶解动力学的影响

含钛钢连铸过程中,部分TiO₂夹杂物进入保护渣,导致渣性能恶化,进而影响铸坯质量。采用旋转法,研究了TiO₂在保护渣中的溶解速率与旋转速度、温度、B₂O₃含量等因素的关系。结果表明,TiO₂在保护渣中溶解速率与棒旋转的角速度的平方根成正比;且在温度升高时,TiO₂在保护渣中的溶解速率有明显的上升;在w(B₂O₃)=0～9%时,随着B₂O₃含量上升,TiO₂溶解速率升高,溶解活化能由162.99降至123.95 J/mol,但B₂O₃含量对溶解速率的影响要明显小于温度的影响。以上研究结果较好阐明了TiO₂在保护渣中的溶解机理,为含钛钢连铸过程提供一定的参考。     

从氧化锌矿浸出渣中回收锌

低品位氧化锌矿酸浸后,浸出渣中夹带3％以上的锌,采用水洗-P204萃取可回收酸浸渣中的水溶性锌,得到的反萃液经过净化后可电积沉锌。该工艺可与湿法炼锌工艺相结合处理低品位氧化锌矿。     

从氧化锌矿浸出渣中回收锌

低品位氧化锌矿酸浸后,浸出渣中夹带3%以上的锌,采用水洗-P204萃取可回收酸浸渣中的水溶性锌,得到的反萃液经过净化后可电积沉锌.该工艺可与湿法炼锌工艺相结合处理低品位氧化锌矿.     

从铜合金渣中选择性回收锡及其热处理合成SnO_2(英文)

采用选择性溶解方法对铜合金废渣中的锡进行富集和分离。使用HNO3溶液溶解铜合金废渣,渣中的锡不会溶解从而使其从中得到分离。通过离心分离生成的H2SnO3沉淀相,然后经过干燥加热生成SnO2粉末。对SnO2粉末的相组成、表面形貌和化学成分进行分析。     

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。     

Leaching kinetics of selenium from copper anode slimes by nitric acid-sulfuric acid mixture

The leaching kinetics of selenium from copper anode slimes was studied in a nitric acid-sulfuric acid mixture. The effects of main parameters on selenium leaching showed that the leaching rate of selenium was practically independent of stirring speed, while dependent on temperature and the concentrations of HNO₃ and H₂SO₄. The leaching of selenium includes two stages. The activation energy in the first stage is 103.5 kJ/mol, and the chemical reaction is the rate controlling step. It was almost independent of H₂SO₄ concentration and dependent on HNO₃ concentration since the empirical reaction order with respect to HNO₃ concentration is 0.5613. In the second stage, the activation energy is 30.6 kJ/mol, and the process is controlled by a mixture of diffusion and chemical reaction. The leaching of selenium was almost independent of HNO₃ concentration.     

Recovery of valuable metals from zinc leaching residue by sulfate roasting and water leaching

Zinc leaching residue (ZLR), produced from traditional zinc hydrometallurgy process, is not only a hazardous waste but also a potential valuable solid. The combination of sulfate roasting and water leaching was employed to recover the valuable metals from ZLR. The ZLR was initially roasted with ferric sulfate at 640 °C for 1 h with ferric sulfate/zinc ferrite mole ratio of 1.2. In this process, the valuable metals were efficiently transformed into water soluble sulfate, while iron remains as ferric oxide. Thereafter, water leaching was conducted to extract the valuable metals sulfate for recovery. The recovery rates of zinc, manganese, copper, cadmium and iron were 92.4%, 93.3%, 99.3%, 91.4% and 1.1%, respectively. A leaching toxicity test for ZLR was performed after water leaching. The results indicated that the final residue was effectively detoxified and all of the heavy metal leaching concentrations were under the allowable limit.     

在H2SO4-HCl-H2O复合体系中氧化浸出镍钼矿冶炼烟尘中的硒

对在H2SO4-HCl-H2O复合体系中氧化浸出镍钼矿冶炼烟尘中的硒进行热力学分析,确定从镍钼矿冶炼烟尘中浸出硒的新工艺及其最优技术参数.采用XRD对镍钼矿冶炼烟尘及其浸出渣进行表征.结果表明:在最佳技术条件下,硒浸出率达到98%,浸出渣含硒0.16%(质量分数);冶炼烟尘中硒以单质形式存在,未见硒及其化合物出现,表明烟尘中的硒浸出较完全;浸出渣主要由SiO2、CaSO4、A12SiO5、As2O3和KAlSi3O8组成.     

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。     

Na2SO3浸出法提纯粗硒工艺研究

采用Na2SO3浸出法提纯粗硒,研究浸硒过程Na2SO3溶液浓度、浸出温度、浸出时间、搅拌速度和净化过程Na2S溶液浓度对提纯粗硒的影响,采用XRD、SEM对所得样品进行表征.结果表明,Na2SO3溶液浓度为300 g/L、浸出温度为98℃、浸出时间为2 h,搅拌速度为150 r/min、Na2S溶液浓度为2 g/L时...     

Recycling of waste carbon residue from spent lithium-ion batteries via constant-pressure acid leaching

Waste carbon residue (WCR) was efficiently detoxicated and regenerated to high-purity graphite (PGC) used in lithium-ion batteries through the constant-pressure acid leaching technique. The leaching conditions were optimized by the combination of orthogonal and single-factor experiments. Results show that PGC with 99.5% purity is regenerated at temperature of 60 °C, initial acid concentration of 12%, leaching time of 180 min, and liquid-to-solid ratio of 25:1, satisfying the requirements of commercial graphite. Meanwhile, the sodium hydroxide precipitation process was designed to recover valuable components from leachate efficiently. Ni, Co, Mn, and Al recoveries reach 96.92%, 87.5%, 97.83%, and 92.17%, respectively, at pH=11. Moreover, the co-product NaF can be recovered with purity over 99% via evaporative crystallization. The loss rate of fluorine is less than 0.5%, thereby eliminating the pollution risk of fluorine to the environment. The proposed process shows considerable environmental and economic benefits.