从高温合金废料浸出渣中浸出钌的实验研究

针对高温合金浸出渣中的钌,采用碱焙烧-水浸工艺进行实验研究,考察了碱料比、焙烧温度、焙烧时间、水浸温度、水浸时间等因素对钌浸出率的影响。结果表明,碱料比、焙烧温度对钌的浸出率影响较大。在最佳的实验条件：碱料比8：1、焙烧温度600℃、焙烧时间4 h、水浸温度95℃、水浸时间3 h下,钌的浸出率为98.3%。     

铀钼矿硫酸焙烧提取铀钼

采用硫酸焙烧-水浸法强化过程高效提取铀钼矿中的铀钼,运用单因素试验考察焙烧过程参数对铀钼浸出率的影响。结果表明:硫酸焙烧过程推荐参数为酸矿质量比0.5:1、硫酸浓度82%、焙烧温度270℃、焙烧时间120 min。优化条件下验证实验所得焙烧熟料水浸后铀钼浸出率可达92%～93%和81%～84%,较现有低温直接酸浸过程铀浸出率(85%～90%)有一定提高,钼浸出率(45%～50%)有大幅提高。对原矿、焙烧熟料、浸出渣物相进行XRD分析后发现,铀钼矿经过硫酸焙烧和水浸后主要组分已由原矿中铝硅酸盐和SiO_2转变为浸出渣中SiO_2。     

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%。     

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。     

锌浸渣硫酸铵焙烧-选择性浸出回收锌

针对锌浸渣中锌难于选择性浸出回收的难题,提出硫酸铵焙烧-选择性浸出回收锌的新工艺。该工艺通过硫酸铵焙烧改变锌浸渣中锌铁物相,在浸出过程对锌进行选择性浸出回收。研究硫酸铵加入量、焙烧温度、焙烧时间等工艺参数对铁酸锌分解和锌铁浸出的影响,并获得最佳的工艺参数,即硫酸铵和铁酸锌质量比为4、一段焙烧温度和时间分别为450℃和90 min,二段焙烧温度和时间分别为650℃和60 min。在该条件下,锌浸出率可以达到92.63%,而铁的浸出率仅为2.04%,实现了锌浸渣中锌的选择性浸出。     

以废茶叶为还原剂在硫酸溶液中还原浸出氧化锰矿（英文）

采用废茶叶在硫酸溶液中还原浸出加蓬和湘西氧化锰矿石,探索废茶叶用量、硫酸浓度、固液比、浸出温度和反应时间对浸出过程的影响。对加蓬氧化锰矿,优化的浸出条件为:氧化锰矿与废茶叶的质量比10:4、硫酸浓度2.5 mol/L、固液比7.5:1、浸出温度368 K、浸出时间8 h;在此条件下,加蓬氧化锰矿的浸出率几乎达100%。对于湘西氧化锰矿,优化浸出条件为:氧化锰矿与废茶叶的质量比10:1、硫酸浓度1.7 mol/L、液固比7.5:1、温度368 K、浸出时间8 h;在此条件下,锰的浸出率达到99.8%。氧化锰矿的还原浸出过程符合内扩散控制模型,加蓬和湘西氧化锰矿石的还原浸出反应表观活化能分别为38.2 kJ/mol和20.4 kJ/mol。采用X-射线衍射(XRD)和扫描电子显微镜(SEM)对浸出前、后的锰渣进行表征。     

再生铝过程铝灰渣提取冶金级氧化铝研究

本文以再生铝冶炼过程产生的二次铝灰渣为原料,经加碱焙烧、水解浸出、液固分离等工艺,综合回收利用铝灰渣。通过单因素试验考察了碱量、焙烧温度、焙烧时间、液固比、浸出温度等因素对铝灰渣中铝浸出率的影响。结果表明,铝灰渣与碳酸钠质量比为1∶1.4,焙烧温度1000℃,焙烧时间2 h,焙烧熟料在液固比为4∶1,浸出温度60℃,浸出时间60 min条件下,铝灰渣中铝的浸出率达到85.54%,加碱焙烧后铝灰渣产出的浸出渣为第Ⅱ类一般固体废弃物,实现了铝灰渣的资源化和无害化利用。     

响应曲面法优化电解锰阳极渣还原浸出工艺

对国内某电解锰厂含铅量高的阳极渣进行了回收锰的实验研究。实验采用葡萄糖作还原剂在硫酸体系中还原浸出电解锰阳极渣。通过基于中心复合设计的响应曲面法对浸出温度、硫酸用量和葡萄糖用量的工艺参数进行研究并优化。研究表明:温度对锰浸出率的影响最显著,葡萄糖的次之,硫酸的最小;硫酸对铅浸出率影响最显著,温度的次之,而葡萄糖则几乎没有影响。在浸出温度80℃,葡萄糖与锰阳极渣质量比为0.175:1、酸渣质量比为0.8:1的条件下,锰的浸出率可达93.22%,铅的浸出率仅为0.39%,锰、铅分离效果明显,锰阳极渣浸出前后的物相通过X射线衍射仪进行表征。实验证明:在硫酸体系中利用葡萄糖还原浸出电解锰阳极渣的方法是可行的。     

高铁锌焙砂还原焙烧-碱浸工艺

利用还原焙烧-碱性浸出工艺处理高铁锌焙砂以解决现有炼锌工艺锌铁分离的难题,通过还原焙烧将高铁锌焙砂中铁酸锌分解为氧化锌和铁氧化物,氧化锌在碱性体系被选择性浸出,铁氧化物赋存于浸出渣中实现锌铁分离。以锌、铁浸出率为评价指标,考察还原焙烧及碱性浸出条件对锌铁分离效果的影响,并对焙烧产物及浸出渣进行XRD、SEM-EDS分析。结果表明:最佳还原焙烧条件如下,焙烧时间45 min,焙烧温度800℃,CO浓度4%(体积分数);最佳浸出条件如下,NaOH浓度240 g/L,液固比12:1,浸出温度80℃,浸出时间60 min。在最佳条件下总锌浸出率约为90%,总铁浸出率约为0.25%,SEM分析显示:浸出渣中锌铁氧化物镶嵌现象严重,这是锌浸出率不能进一步提高的原因。     

电解锰阳极渣还原浸出锰

针对电解锰阳极渣难处理、铅含量高的缺点,提出利用桔子皮作还原剂在硫酸体系中还原浸出电解锰阳极渣工艺。以国内某电解锰厂阳极渣为原料,对桔子皮加入量、浸出时间、浸出温度以及硫酸加入量等工艺参数进行探讨和优化。结果表明:在浸出温度为80℃,时间为2 h,固液比为1:4,桔子皮/锰阳极渣质量比为1:5,酸渣质量比为1.2:1的条件下,锰的浸出率可达96%,铅的浸出率仅为0.2%,有效地实现了铅锰分离。实验证明,在硫酸体系中利用桔子皮作还原剂浸出电解锰阳极渣的方法可行。     

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.     

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.     

Process optimization and modeling of recycling Mo (VI) from spent Mo-Fe2O3/Al2O3 catalyst by roasting with sodium carbonate using response surface methodology (RSM)

The oil-containing spent Mo-Fe₂O₃/Al₂O₃ catalyst can be deemed as an environmental threat and an attractive source of minerals that can reduce the consumption of natural resources. Herein, recovery of Mo from spent Mo-Fe₂O₃/Al₂O₃ catalyst was conducted by the Na₂CO₃ roast-leach process, response surface methodology (RSM) coupled with central composite design (CCD) was employed to optimize the roasting process and a polynomial equation was derived to predict the response. The three roasting independent variables the roasting temperature, the Na₂CO₃/sample weight ratio, and the roasting duration were investigated in the Na₂CO₃ roasting process while water-leaching parameters were identical. The predictions of model showed that the roasting temperature had a major effect on the response with respect to other parameters. According to analysis of variance (ANOVA), the proprosed model equation had shown satisfactory agreement with the experimental data with a correlation coefficient (R²) of 0.9811. The optimum conditions for Mo recovery were predicted to be as the roasting temperature of 771.2 °C, the Na₂CO₃/sample weight ratio of 2.09 and the roasting duration of 93.56 min. Under the optimum conditions, maximal value of Mo recovery rate was reached as 92.58%.     

A Sustainable and Selective Roasting and Water-Leaching Process to Simultaneously Extract Valuable Metals from Low-Grade Ni-Cu Matte

Due to stringent environmental requirements and the complex occurrence of valuable metals, traditional pyrometallurgical methods are unsuitable for treating low-grade nickel-copper matte. A clean and sustainable two-stage sulfating roasting and water-leaching process was used to simultaneously extract valuable metals from low-grade nickel-copper matte. Ammonium and sodium sulfate were used as sulfating agents. The first roasting temperature, mass ratio of ammonium sulfate to matte, roasting time, dosage of sodium sulfate, second roasting temperature and leaching temperature were studied. Under optimal conditions, 98.89% of Ni, 97.48% of Cu and 95.82% of Co, but only 1.34% of Fe, were extracted. X-ray diffraction (XRD) and scanning electron microscopy (SEM) were used to reveal the sulfating mechanism during the roasting process.     

Optimization of Fluidized Roasting Reduction of Low-Grade Pyrolusite Using Biogas Residual as Reductant

Research on the novel technology of fluidized roasting reduction of samples of low-grade pyrolusite using biogas residual as reductant has been conducted. According to the response surface design and the analysis of results, orthogonal experiments have been conducted on the major factors, and the effects on the manganese reduction efficiency have been studied. The maximum manganese reduction efficiency could be optimized to nearly 100%, when the mass ratio of biogas residual to pyrolusite was 0.16:1, the dosage of sulfuric acid was 1.6 times that of the stoichiometric amount, the roasting temperature was 680°C, and the roasting time was 70?min. The results in terms of manganese reduction efficiency of the actual experiments were close to those anticipated by modeling the experiments, indicating that the optimum conditions had a high reliability. Other low-grade pyrolusites such as Guangxi pyrolusite (China), Hunan pyrolusite (China), and Guizhou pyrolusite (China) were tested and all these materials responded well, giving nearly 100% manganese reduction efficiency.     

Recovering indium with sulfating roasting from copper-smelting ash

A technology for recovering indium from Jinchuan copper-smelting ash was developed. Indium in the ash was first enriched to the leaching-slag in leaching process,and then recovered by sulfating roasting. The method included mixing the leaching-slag with sulfuric acid,making them into particles,roasting the mixture,and then leaching the calcine with hot water. Above 90% of indium in calcine could be dissolved into the leaching solution. The optimized conditions were determined as follows： the mass ratio of sulfuric acid to leaching slag was 0.1,the roasting time was about 1 to 1.5 h in the temperature range of 200-250℃,and the calcine was leached for 1 h with 5：1 of liquid/solid ratio at 60℃. Over 99% of indium in leaching solution was finally enriched by Zn substitution or sulfide precipitation.     

火试金重量法测定金碳基催化剂中金含量

采用火试金重量法测量含金碳基催化剂中金含量,通过焙烧温度、称样量、物料配比、金银质量比以及分金酸度等确定最优实验条件。在最优实验条件下通过加标实验,标准物质回收率99.0%~100.6%。方法准确度良好,可运用于实际生产分析。     

氯化焙烧-水浸法从锂云母矿提锂(英文)

采用氯化焙烧-水浸法处理锂云母矿,并对氯化处理温度、时间、氯化剂的类型及用量进行研究。条件优化实验表明,在锂云母、氯化钠、氯化钠的质量比为 1:0.6:0.4,氯化处理温度为 880 °C,氯化处理时间为 30 min时,锂的提取率可达 92.86%,钾、铷、铯的提取率分别为 88.49%、93.60%和 93.01%,。采用 XRD 对锂云母原矿及焙烧后物料的物相进行分析。XRD 结果分析表明,当将锂云母和混合氯化剂一起焙烧(氯化钙及氯化钠)时,所得物相为 SiO2、CaF2、KCl、CaSiO3、CaAl2Si2O8、NaCl 和 NaAlSi3O8。     

废弃SCR脱硝催化剂与废NaCl盐焙烧回收催化剂中的钛、钒、钨（英文）

通过废弃选择性催化还原(SCR)脱硝催化剂与废NaCl盐焙烧,可以将催化剂中的钨和钒与钛分离。在最佳浸出条件下(焙烧温度900℃,焙烧时间3 h,废盐与废催化剂的质量比为0.5,浸出温度80℃,反应时间60 min),钨和钒的浸出率分别达到84.63%和66.42%,同时钛的损失率仅为1.3%。废NaCl盐和焙烧温度可以促进锐钛矿型TiO₂转化为金红石型TiO₂,反应后得到了金红石型TiO₂。金红石型TiO₂中的钛的价态为四价,晶格氧和化学吸附氧分别占57.26%和42.74%。该方法可以同时解决2种废弃物的处置问题。     

从镍钼矿中提取镍钼的工艺

针对现行镍钼矿处理工艺存在的钼镍需要分别提取的缺陷,提出镍钼矿加钙氧化焙烧-低温硫酸化焙烧-水浸提取镍钼的新工艺。以贵州遵义镍钼矿为原料,对CaO加入量、氧化焙烧温度、氧化焙烧时间、硫酸加入量、硫酸化焙烧温度、硫酸化焙烧时间以及焙砂水浸工艺参数对镍钼浸出率的影响进行研究。结果表明:在最佳工艺条件下,钼的浸出率为97.33%,镍的浸出率为93.16%,且最佳工艺参数为100 g镍钼矿加入35 g CaO,700℃氧化焙烧2 h,得到的焙砂加入70 mL浓硫酸,再经250℃硫酸化焙烧2 h;硫酸化焙烧得到的焙砂按液固比2:1加水搅拌,经98℃浸出2 h。加入CaO不仅能有效减少镍钼矿氧化焙烧烟气对环境造成的污染,而且能显著提高镍的浸出率。