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%。     

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。     

废盐酸浸出菱锰矿制备四水氯化锰

以某低品位菱锰矿为原料,采用废盐酸浸出,研究浸出温度、浸出时间、液固比、反应时酸过量系数对矿石中锰的浸出效果的影响,讨论锰浸出的最佳工艺条件.结果表明:浸出温度80 ℃、浸出时间60 min、液固比2.5-1、酸过量系数1.3为最佳工艺条件.该工艺为低品位锰矿的开发利用及钛厂废盐酸的综合利用开辟了一条新途径.     

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。     

葡萄糖在氧化锰矿浸出过程中的分解动力学

研究在硫酸水溶液体系葡萄糖还原浸出氧化锰矿过程中的分解动力学,通过测定溶液中COD含量的变化,考察葡萄糖用量、硫酸浓度、氧化锰矿用量和反应温度对葡萄糖氧化分解速率的影响。结果表明:提高硫酸浓度、氧化锰矿用量和反应温度可以加快浸出反应过程葡萄糖的氧化分解速率。葡萄糖的氧化分解过程可以用指数经验模型来描述,属于扩散-化学反应混合控制。通过拟合动力学数据,获得葡萄糖氧化分解反应的活化能为41.80 k J/mol,COD、硫酸和氧化锰矿的表观反应级数分别为2.0,0.927和0.976。     

焦化废水还原浸出软锰矿动力学

以焦化废水为还原剂,在硫酸介质中利用收缩芯模型研究焦化废水加压直接还原浸出软锰矿的动力学,考察搅拌速度、反应温度、软锰矿粒径、硫酸浓度和焦化废水CODcr浓度对锰浸出速率的影响。结果表明:锰浸出率随反应温度、硫酸浓度、焦化废水CODcr浓度的增加和软锰矿粒径的减小而增大。在393 K~423 K间,焦化废水还原浸出软锰矿受固膜扩散控制,表观活化能为15.2 kJ/mol,硫酸和焦化废水CODcr浓度的表观响应级数分别为1.21和0.98。机理分析表明:软锰矿氧化焦化废水中大分子有机物分解成小分子而被还原溶出,或被小分子直接还原溶出;反应过程中,软锰矿表面残留的Fe、Si、Al形成孔洞薄壁而影响锰的溶出过程。     

铵盐焙烧-酸浸法从石油重整废催化剂中富集回收铂的研究

采用铵盐焙烧-酸浸法进行了从石油重整废催化剂中富集回收铂的研究。结果表明,适宜的铵盐焙烧条件为:硫酸铵与废催化剂质量比为7.5,焙烧温度350℃,焙烧时间5 h;稀酸浸出条件:液固比12.5,浸出温度80℃,硫酸浓度为0.5 mol/L,浸出时间3 h。在上述条件下,铂的富集倍数达到274倍以上。通过XRD分析,焙烧产物的物相主要以硫酸铝铵形式存在,焙烧产物通过稀酸浸出,实现铂的富集回收。     

燕麦秸秆还原浸出低品位软锰矿及其动力学

以燕麦秸秆为还原剂,在硫酸介质中直接浸出某低品位软锰矿,研究浸出过程的工艺条件和动力学。结果表明:在硫酸浓度为150 g/L、液固比为9 mL/g、麦秆与软锰矿质量比为0.3、浸出温度90℃、浸出60 min的条件下,锰浸出率可达96.1%,此时,铁浸出率为46.8%,浸出液化学需氧量(COD)值为6.2 g/L。动力学研究表明:锰的浸出过程在333~363 K内符合未反应核缩减模型,浸出过程主要受界面化学反应步骤控制,其浸出表观活化能为69.4 kJ/mol。     

硫酸肼还原酸浸出废旧锂离子电池中的有价金属（英文）

采用硫酸肼作为锂、镍、钴和锰从废锂离子电池中浸出时的还原剂,结合条件实验对浸出机理和浸出动力学进行研究。在最优条件:硫酸2.0 mol/L、硫酸肼30 g/L、固液比50 g/L、温度80℃和浸出时间60 min下,97%的Li、96%的Ni、95%的Co以及86%的Mn被浸出。通过浸出动力学分析得出Li、Ni以及Co的浸出活化能分别为44.32、59.37和55.62 k J/mol,表明浸出过程受化学反应控制。XRD和SEM-EDS分析结果表明,浸出渣的主要组成为MnO_2。上述研究结果表明,硫酸肼可作为废锂离子电池中有价金属浸出的有效还原剂。     

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 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.     

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。     

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.     

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%。     

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。     

Stepwise recovery of magnesium from low-grade ludwigite ore based on innovative and clean technological route

A novel and clean technological route for the comprehensive utilization of low-grade ludwigite ore was proposed, in which magnesium was extracted by metallizing reduction-magnetic separation, sulfuric acid leaching and ethanol precipitation operation. Meanwhile, iron-rich product, silicon-rich product and boron-rich product were obtained, respectively. In the process of metallizing reduction-magnetic separation, 94.6% of magnesium was enriched in the non-magnetic substance from the ore reduced at 1250 °C for 60 min with the ore size of 0.50–2.00 mm and coal size of 0.50–1.50 mm. When the non-magnetic substance was leached at 90 °C for 15 min with the liquid-to-solid ratio of 7:1, 87.4% of magnesium was leached into the liquor separated from silicon gathering in leaching residue. The ethanol precipitation was conducted for 30 min with the ethanol-to-original liquid volume ratio of 1.5:1 at room temperature. 97.2% of magnesium was precipitated out with the initial concentration of 0.8 mol/L in the form of MgSO₄·7H₂O.     

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.     

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.