赞比亚某氧化铜矿石的硫酸酸浸研究

赞比亚某低品位高结合率难处理氧化铜矿石铜品位为1.56%,主要铜矿物为赤铜矿、黄铜矿、铜蓝、水胆矾;主要脉石矿物为石英、云母、铁白云石等。铜氧化率高达82.85%,以结合氧化铜为主;硫化铜仅占17.15%,主要为原生硫化铜。为确定该矿石的合理开发利用工艺,进行了系统的硫酸酸浸试验。结果表明:①提高浸出试样细度,延长浸出时间,提高浸出温度,增大液固质量比和搅拌速度均有利于改善氧化铜矿石的浸出效果。②矿石在磨矿细度为-200目占60%、硫酸浓度为50 g/L、液固质量比为3、浸出温度为65℃、搅拌速度为300 r/min,浸出时间为120 min情况下,铜的浸出率达78.64%。③硫酸浸出该矿石的浸出动力学受化学反应模型控制,反应的表观活化能为37.83 k J/mol。

基于还原焙烧-酸浸法的电解锰阳极泥中锰铅分离研究

采用硫磺还原焙烧-硫酸浸出处理电解锰阳极泥(EMAS),实现其中锰、铅分离。研究了焙烧过程中焙烧温度、硫磺与锰物质的量比、焙烧时间、EMAS粒径以及浸出过程中搅拌速度、浸出温度、浸出时间、固液比及硫酸浓度对锰、铅分离的影响。结果表明,焙烧优化条件为:焙烧温度550℃、硫磺与锰物质的量比1∶1、焙烧时间50 min、EMAS粒径-0.177 mm,酸浸优化条件为:搅拌转速350 r/min、浸出温度40℃、浸出时间20 min、固液比1∶20、硫酸浓度2.5 mol/L,此条件下阳极泥中锰浸出率达97.4%(滤液中锰离子浓度为38.84 g/L),浸出渣中Pb含量达45.26%,有效实现了锰与铅的分离。

江西弋阳蛇纹石硫酸浸出过程工艺条件的优化研究

文章在常压、较低温度(<100℃)下,开展了江西弋阳蛇纹石硫酸浸出工艺条件的研究。以蛇纹石中氧化镁的浸出率为主要考察指标,通过单因素条件实验和正交实验,分别考察了蛇纹石酸浸反应温度、酸浸反应时间、硫酸溶液质量浓度、液固比等因素的变化对蛇纹石中氧化镁浸出率的影响,所得较佳的工艺条件为:酸浸反应温度95℃、酸浸反应时间5 h、硫酸溶液质量浓度40%、液固比4.5︰1。此条件下的重复实验结果:氧化镁的平均浸出率为99.18%。

Acidithiobacillus ferrooxidans-硫酸酸溶联合浸出金川铜镍硫化矿尾矿砂实验研究

金川铜镍硫化矿尾矿砂富含Cu、Ni和Co等金属,为对其进行回收,对新尾矿库不同深度的尾矿砂分别进行了直接硫酸酸溶浸出与A.f.-硫酸酸溶联合浸出实验,利用ICP-AES分析了不同阶段浸出液中Cu2+、Ni2+、Co2+含量。实验结果表明,A.f.-硫酸酸溶联合浸出工艺对尾矿砂中有价金属Cu、Ni和Co的浸出效果优于单独酸溶浸出工艺。其中,A.f.-硫酸酸溶联合浸出法对1m深度的尾矿砂中Cu、Ni、Co的浸出率最高,分别达到95.78%、98.01%和75.13%,而单独硫酸酸溶作用的离子浸出率分别为75.69%、74.33%和42.33%。

电镀污泥中铜和镍的回收工艺研究——污泥的酸浸出工艺

研究了电镀污泥中铜和镍浸出的方法,对比选取硫酸作为浸出剂,考察了酸的用量对浸出效果的影响,得到最佳浸出条件为:污泥颗粒d=0.15mm,每2g污泥加10%硫酸10mL,常温下振荡0.5h。该条件下电镀污泥中铜、镍的浸出率均较高,达95%以上,为后续电解回收工艺奠定了较好的基础。

河北某难选钼矿石酸浸浸钼试验

河北某难选钼矿石氧化程度较高,氧化钼占总钼的52.54%,主要以钼华、钼钙矿和钼铅矿等形式存在;硫化钼矿物主要为非晶质或隐晶质胶硫钼矿,多以胶状集合体产出,单体解离困难,可浮性较差。为高效开发利用该钼矿资源,进行了酸浸工艺条件研究。结果表明,100 g粒度为-2 mm的试样中加入30 g浓硫酸,在液固比为1.5m L/g、浸出温度为90℃、搅拌器转速为250 r/min情况下浸出3 h,钼浸出率达86.19%。因此,硫酸酸浸工艺是该资源的高效开发利用工艺。

氧化铝赤泥酸浸富集回收镓的新工艺研究

研究了拜耳法赤泥硫酸浸出镓的过程。硫酸浸出镓的溶液,用氢氧化钠反应生成沉淀,再用盐酸溶解沉淀提取镓的过程。采用正交试验考察浸出过程中温度、时间、液固比、浓度对镓浸出率的影响的主次顺序以及最佳条件。结果表明,该方法比单纯用盐酸浸出镓的浸出率高。镓的高浸出率是保证高的回收率的基础。

硫酸盐溶解性的考查剖析

高中教材出现了4种常见难溶的硫酸盐,分别是BaSO_(4)、CaSO_(4)、PbSO_(4)与Ag_(2)SO_(4),溶解性见表1.关于溶解性的考查特别容易被忽视,本文归纳出5个考查角度便于师生的教与学.1工艺流程中的分离提纯在工艺流程中,矿石中的Ca、Pb元素在硫酸酸浸时形成CaSO_(4)、PbSO_(4)滤渣被除去.

富铼渣浸出工艺改进试验研究

针对某铜业公司铼酸铵回收系统铼浸出率不高及产品铼酸铵品位较低的现实情况,将原有的高酸氧化浸出改为单一双氧水体系氧化浸出,采取实验室小型试验研究方式,进行富铼渣浸出工艺改进试验研究。试验结果表明:在双氧水用量400 m L/100 g渣、时间120 min、浸出温度70℃条件下,浸出液中铼的浓度可达1 988 mg/L,相较原有生产工艺,铼浸出率得到大幅提高,平均达到90%以上;同时杂质铋、砷的浓度分别从110~150 mg/L和260~400 mg/L下降到33.62 mg/L和14.35 mg/L,杂质浸出明显降低。

国外某氧化铜矿的浸出研究

以国外某氧化铜矿床为研究对象,本文采用硫酸酸浸工艺,考察浸出时间、反应温度、液固比、硫酸消耗、磨矿粒度等因素对铜浸出率的影响,得到最佳浸出条件。研究表明,提高浸出温度和酸矿比,混合矿Cu浸出率有所提高,但是增幅较小,同时杂质Fe浸出率增大。浸出温度升高,酸耗相应增大。因此,推荐室温下浸出,酸矿比130 kg/t,浸出时间2 h,此时Cu浸出率约为92%。此外,矿石粒度对Cu的浸出率无明显影响。

复杂高含铟物料铟浸出工艺研究

某铅冶炼厂的复杂高含铟物料为冶炼的中间产物,此物料采用常规酸浸工艺铟的浸出率较低。针对此问题,本文采用盐酸-硫酸混酸浸出、加压酸浸法、硫酸化焙烧-低酸浸出等3种工艺对此复杂高含铟物料进行了试验,在各试验的较佳工艺条件下,试验结果为:采用盐酸-硫酸混酸浸出工艺,铟的浸出率可以达到75.87%,浸出渣铟含量可以降低至0.32%;采用加压酸性浸出工艺,铟的浸出率可以达到82.54%,浸出渣铟含量可降低至0.21%;采用硫酸化焙烧-低酸浸出工艺,铟的浸出率可以达到93.28%,浸出渣铟含量可降低至0.09%。综合分析对比3种铟的浸出工艺,"硫酸化焙烧-低酸浸出"工艺比较适合此复杂含铟物料提铟。

Reductive acid leaching of cadmium from zinc neutral leaching residue using hydrazine sulfate

Zinc neutral leaching residue（ZNLR） from hydrometallurgical zinc smelting processing can be determined as hazardous intermediate containing considerable amounts of Cd and Zn which have great threats to the environment. The ZNLR contained approximately 35.99% Zn, 15.93% Fe and 0.26% Cd, and Cd mainly existed as ferrites in the ZNLR in this research. Reductive acid leaching of ZNLR was investigated. The effects of hydrazine sulfate concentration, initial sulfuric acid concentration, temperature, duration and liquid-to-solid ratio on the extraction of Cd, Zn and Fe were examined. The extraction efficiencies of Cd, Zn and Fe reached 90.81%, 95.83% and 94.19%, respectively when the leaching parameters were fixed as follows： hydrazine sulfate concentration, 33.3 g/L; sulfuric acid concentration, 80 g/L; temperature, 95 °C; duration of leaching, 120 min; liquid-to-solid ratio, 10 m L/g and agitation, 400 r/min. XRD and SEM-EDS analyses of the leaching residue confirmed that lead sulfate（Pb SO4） and hydrazinium zinc sulfate（（N2H5）2Zn（SO4）2） were the main phases remaining in the reductive leaching residue.

复杂稀有金属伴生矿富集渣提取稀土和铌的工艺研究

研究了从复杂稀有金属伴生矿富集渣中提取稀土和铌的工艺。结果表明,采用硫酸酸化-分段浸出工艺可实现富集渣中稀土和铌的高效浸出。在酸矿质量比1.8、酸化温度350℃、酸化时间120min、一段浸出液固比1∶1、浸出温度80℃、浸出时间90min、二段浸出液固比8∶1、浸出温度25℃、浸出时间90min的条件下,浸出渣中REO含量为0.96%,Nb2O5含量为0.75%,稀土浸出率为85.03%,铌浸出率为80.88%。其中铌一段浸出率为80.26%,稀土二段浸出率为83.85%,可通过分别处理一段浸出液和二段浸出液实现铌和稀土的回收。

以硫铁矿烧渣制备絮凝剂(PFS)的最佳路线研究

研究了在常压下从硫铁矿烧渣中浸取铁的最佳实验条件,并进行絮凝剂的制备及其性能的测定。考虑了温度、硫酸的质量分数、酸渣比对浸取效率的影响,并在相同条件下探讨了不同助溶剂对浸取率的影响。实验表明当温度为110℃、硫酸质量分数为40%、酸渣比为100:2时,铁的浸取率高达56%。

尼日利亚某氧化铜矿石选矿试验

尼日利亚某铜矿石属于铜品位高、氧化程度深、含泥量大、铜矿物组成复杂且嵌布粒度粗细不均的难选氧化铜矿石。为确定矿石的合理开发利用工艺,分别进行了硫化浮选工艺和硫酸酸浸工艺研究。结果表明,矿石在磨矿细度为-0.074 mm占90%的情况下,采用1粗4精3扫浮选流程处理,可获得铜品位为20.23%、铜回收率为74.35%的铜精矿;矿石在磨矿细度为-0.074 mm占55%、硫酸浓度为74 g/L、矿浆浓度为33%、浸出时间为2 h、搅拌速度为300 r/min的情况下,铜浸出率可达77.22%。从节能、增效角度考虑,酸浸工艺相对更适合该矿石的处理,在磨矿细度-0.074 mm含量从90%降至55%的情况下,铜浸出率高出浮选工艺铜回收率2.87个百分点。

含铜含镍污泥典型资源化利用技术研究

含铜含镍污泥资源化利用具有良好的生态效益和经济效益。富氧侧吹浸没熔池熔炼法、酸浸+碳酸钠沉淀法目前已有一定的工程应用。本文详细介绍了这两种典型工艺的流程与原理,并对其处置规模、节能环保、成品纯度、投资强度、运营成本等多角度进行对比,推荐较为适合的资源化利用技术。

Extraction of molybdenum and nickel from roasted Ni-Mo ore by hydrochloric acid leaching, sulphation roasting and water leaching

To extract molybdenum and nickel from the roasted Ni-Mo ore, a process of hydrochloric acid leaching, sulphation roasting and water leaching was investigated. The results showed that this process could get a high leaching rate of Mo and Ni. Under the optimum conditions of hydrochloric acid leaching （roasted Ni-Mo ore leached with 0.219 mL/g hydrochloric acid addition at 65 ℃ for 30 min with a L/S ratio of 3 mL/g）, sulphation roasting （51.9% sulfiaric acid addition, roasting temperature 240 ℃ for 1 h）, followed by leaching with the first stage hydrochloric acid leaching solution at 95 ℃ for 2 h, the leaching rates of Mo and Ni reached 95.8% and 91.3%, respectively.

Reductive acid leaching of valuable metals from spent lithium-ion batteries using hydrazine sulfate as reductant

Hydrazine sulfate was used as a reducing agent for the leaching of Li,Ni,Co and Mn from spent lithium-ion batteries.The effects of the reaction conditions on the leaching mechanism and kinetics were characterized and examined.97%of the available Li,96%of the available Ni,95%of the available Co,and 86%of the available Mn are extracted under the following optimized conditions:sulfuric acid concentration of 2.0 mol/L,hydrazine sulfate dosage of 30 g/L,solid-to-liquid ratio of 50 g/L,temperature of 80℃,and leaching time of 60 min.The activation energies of the leaching are determined to be 44.32,59.37 and 55.62 k J/mol for Li,Ni and Co,respectively.By performing X-ray diffraction and scanning electron microscopy in conjunction with energy dispersive X-ray spectroscopy,it is confirmed that the main phase in the leaching residue is MnO2.The results show that hydrazine sulfate is an effective reducing agent in the acid leaching process for spent lithium-ion batteries.

Extraction of aluminium as aluminium sulphate from thermal power plant fly ashes

Valuable metal extraction technology from thermal power plant fly ash is limited.In the present study,aluminium is extracted from fly ash as highly pure aluminium sulphate(>99.0%)by leaching with sulphuric acid,followed by pre-concentration and successive crystallization.Two types of fly ashes from different sources,i.e.,Talcher Thermal Power Station(TTPS)and Vedanta Aluminium Company Limited(VAL)were chosen for comparative study on the extraction of aluminium as aluminium sulphate.The product is characterized by powder X-ray diffraction(XRD),Fourier transform infrared spectroscopy(FTIR)and thermogravimetric analysis(TGA).Purity of aluminium sulphate was also investigated by inductively coupled plasma?optical emission spectrometry(ICP?OES).The extraction efficiency of aluminium depends on the varied solid-to-liquid ratio(fly ash:18mol/L H2SO4,g/mL)and particle size of fly ashes.Physico-chemical analysis indicates that the obtained product is Al2(SO4)3·18H2O,having low iron content(0.08%).

Enhancement of CaMoO_(4) calcine decomposition and recovery of calcium resource by HCl cycle leaching

The thermodynamic equilibrium diagrams of CaMoO_(4)-CaSO_(4)-H_(2)SO_(4)-H_(2)O,CaMoO_(4)-HCl-H_(2)O and CaSO_(4)-CaCl_(2)-HCl-H_(2)O systems at 298 K were established.The calculation results demonstrated that HCl displays a much higher solubility of CaSO_(4)than H_(2)SO_(4).The leaching mechanism of Mo from CaMoO_(4)calcine was systematically investigated from the perspective of the micro particle properties variation.HCl exhibits an excellent leaching performance for Mo from CaMoO_(4)calcine due to the elimination of surface coating and the dissolution of a mass of Mo embedded in CaSO_(4)matrix.Excellent Mo leaching efficiency of 99.7%was achieved under the optimal conditions of decomposing CaMoO_(4)calcine by 2.4 mol/L HCl with a liquid/solid ratio of 10:1 at 50°C for 60 min.Based on the experimental results,a highly efficient and green cycle leaching process of molybdenum from molybdenite was proposed,which eliminated surface coating and physical entraining,and converted most of the calcium from CaMoO_(4)calcine into high purity gypsum by-product.