从含钼石煤中提取钼的研究

采用XRD和光谱分析确定了含钼石煤中钼的主要物相。将石煤粉碎过0.074mm筛,添加碳酸钠造球,氧化焙烧,然后用水作为浸出剂浸出钼,考查了焙烧温度、焙烧时间和球团粒径对钼浸出率的影响。结果表明,最佳焙烧温度为650℃、焙烧时间为3h、球团粒径大小为5～15mm,在该条件下,钼浸出率可达96.32%。     

泉山金矿氰化尾矿焙烧—超声波强化硫脲提金试验研究

对泉山金矿氰化尾矿进行了焙烧预处理—超声波强化硫脲浸金试验研究。其结果表明：该尾矿经焙烧后硫脲浸出,金的最高浸出率比未焙烧时提高了45.12%;尾矿焙烧后再经超声波强化硫脲浸出,金的最高浸出率进一步提高了9.6%,达到77.5%,且大大缩短了浸出时间,提高了浸金效率。     

氯离子作用下N235对铋的萃取性能

采用溶剂萃取法脱除铜电解液中的杂质,在水相料液中添加助萃剂Cl-,研究Cl-作用下萃取剂N235对铋的萃取性能.考察有机相N235体积分数、水相助萃剂Cl-浓度、有机相与水相的体积比 (相比)、萃取时间、萃取温度等因素对铋萃取率的影响.研究结果表明:有机相N235体积分数、水相助萃剂Cl-浓度和相比是影响铋萃取率的主要因素;较为适宜萃取铋的条件为:水相硫酸浓度为3.0 mol/L,氯离子浓度为0.1 mol/L,萃取剂N235体积分数为20 %,相比为1:1,萃取时间为5 min.在此实验条件下,铋的一级萃取率达到97.3 %(质量分数).      

从锌浸出渣中湿法回收锌

研究了采用硫酸焙烧—水浸出工艺从锌浸出渣中回收锌。考察了硫酸用量、焙烧时间、焙烧温度、液固体积质量比及浸出时间对锌浸出率的影响。试验结果表明:锌浸出渣与70%浓硫酸混合,在250℃下焙烧2.5h,然后以水为浸出剂,按液固体积质量比4∶1、常温下浸出1h,锌浸出率达85%以上。     

溶剂萃取法从加压浸出液中提取钼

硫化钼精矿的处理方法通常采用氧化焙烧法,但在中国,该方法存在的最大问题是低浓度SO2污染。近年来,一些可行的工艺被研究,如加压浸出(POX)-萃取(SX)工艺。北京矿冶研究总院(BGRIMM)采用该工艺进行了一系列的试验研究,研究结果表明,采用该工艺钼总回收率大于97%。加压浸出过程中有15%～20%的钼进入溶液,该部分钼可采用溶剂萃取的方法进行提取,在有机相组成N235+煤油,O/A=3/1,混合时间3min的条件下,钼的萃取率可达到98%以上,负载有机相采用25%的氨水反萃,O/A=1/3,在两级逆流萃取的条件下钼反萃率超过98%。今后在中国,POX-SX工艺在钼提取方面是一个非常具有竞争优势的工艺。     

浸出条件对稀土和铁浸出率的影响试验

以稀土精矿浓硫酸焙烧工艺中焙烧矿水浸过程为对象,研究了焙烧矿浸出温度、浸出时间、焙烧矿粒度等条件对稀土、铁浸出率的影响,并对水浸渣中稀土赋存状态进行了研究。研究表明,浸出温度和焙烧矿粒度对稀土、铁的浸出速率有较大影响,但对其浸出率没有影响,延长浸出时间,焙烧矿中的可溶性稀土、铁均可被浸出。水浸渣中的稀土主要以磷酸盐和氟氧化稀土形式存在,铁主要以磷酸铁形式存在,并含有少量硫化铁。     

HBL101从镍钼矿焙烧料高酸浸出液中直接萃取钼的研究

采用新型萃取剂HBL101从镍钼矿焙烧料高酸浸出液中直接萃取钼。考察有机相组成、料液酸度、相比、振荡频率、平衡时间、温度对钼萃取过程的影响,并绘制了HBL101萃钼等温曲线。结果表明,在优化的工艺条件下,钼萃取率达96.8%以上,有机相饱和容量为12.09g/L;负载有机相用纯水洗涤后经3级逆流氨水反萃,钼反萃率达99.9%以上,实现了钼镍分离及钼的富集转型。     

The leaching of copper from sulphur activated chalcopyrite with cupric sulphate in nitrile-water mixtures

If chalcopyrite is roasted with sulphur at 400–450°C pyrite and idaite or bornite are produced. Bornite plus pyrite are also prepared by roasting a 1:1 mixture of chalcopyrite and covellite. These copper-iron sulphides were leached with acidified aqueous cupric sulphate solutions containing acetonitrile or hydracrylonitrile and the results are compared with leaching with acidified cupric chloride in brine. The nitrile route has the advantage of a less corrosive sulphate medium for subsequent copper recovery processes.Bornite appears to be the most attractive product from the roasting of sulphur and chalcopyrite because much of its copper can be readily leached. Iron reports to the solution only in the latter stages of extraction. Up to 80% of the copper in this bornite is leached with CuSO₄/RCN/H₂O at 60°C. Copper is recovered from the resulting cuprous sulphate solution by electrowinning with inert anode. The products are copper cathodes and cupric sulphate, which is recycled. The leach residue may be used to reactivate further chalcopyrite or is leached of its copper by established routes.     

铝电解质钙化焙烧多元素回收利用研究

针对我国铝电解工业每年约产生60万t过剩铝电解质的全元素资源化利用进行了研究。在分析铝电解质化学成分和物相组成的基础上,研究了有效氧化钙添加量对铝电解质钙化焙烧反应的影响,揭示了铝电解质钙化焙烧产物在碱性溶液中的浸出及反应产物。试验结果表明,有效氧化钙添加量为40%~50%的条件下进行钙化焙烧反应并在碱液中浸出,铝电解质中的铝、钠、氟元素回收率分别为93.35%、90.21%、90.53%,实现铝电解质中的氟、钠、铝、钙等多组分有价元素的资源化利用。     

综合回收钼酸母液中镍、铋、镁的新工艺

采用酸浸法浸出废催化剂中的钼、镍、铋、镁，将浸出液沉钼酸后，对钼酸母液中的镍、铋、镁经过多级沉淀处理后加以综合回收。实现了工业化生产，达到了资源综合利用的目的。     

微波焙烧对石煤钒矿浸出的影响

研究了微波焙烧预处理石煤钒矿,考察了微波功率、微波焙烧时间等因素对钒浸出率的影响。结果表明,微波焙烧预处理钒矿石12min,然后在搅拌条件下用硫酸浸出2h,钒浸出率达69.04%,与相同浸出时间内未焙烧矿样相比,提高了近20%。微波焙烧预处理后浸出效率明显提高,浸出时间缩短50%以上,而且能耗和污染也大幅度降低。     

Kinetics of nitric acid leaching of cerium from oxidation roasted Baotou mixed rare earth concentrate

The kinetics of nitric acid leaching of cerium was investigated for the oxidation roasted Baotou mixed rare earth concentrate. The effects of leaching temperature, HNO₃ concentration, liquid–solid ratio (L/S) and stirring rate on rare earth extraction were studied. The XRD and SEM mapping analysis of the samples before and after acid leaching shows that the roasted bastnaesite is completely leached. Besides, the decomposition process of oxidizing roasting was also obtained by TG–MS and XRD. Different kinetics models were applied in this leaching process. The results of dynamic fitting show that the leaching process can be described by a new variant of the shrinking-core model. And the leaching rate is controlled by both the interfacial transfer and diffusion through the product layer. The apparent activation energy is calculated as 76.78 kJ/mol and the reaction orders with respect to HNO₃ concentrations and liquid–solid ratio are determined to be 7.609 and 2.516, respectively. Besides, an empirical rate equation is obtained to describe the process.     

Leaching behavior of metals from a limonitic nickel laterite using a sulfation–roasting–leaching process

The leaching behavior of metals from a limonitic laterite was investigated using a sulfation–roasting–leaching process for the recovery of nickel and cobalt. The ore was mixed with water and concentrated sulfuric acid followed by roasting and finally leaching with water. Various parameters were studied including the amount of acid added, roasting temperature and time, sample particle size, addition of Na₂SO₄ and solid/liquid ratio in leaching process. More than 88% Ni, 93% Co and < 4% Fe are extracted under the determined conditions. Simultaneously, about 90% Mn and Cu, 70% Mg, 45% Al, 25% Zn, 4% Cr and Ca are extracted respectively. The pH of the leach solution is about 2. The leaching efficiency is independent of sample particle size due to decomposition of ferric sulfate formed during roasting. The roasted mass was characterized by various physico-chemical techniques such as DSC/TGA, XRD and SEM. This process provides a simple and effective way for the extraction of nickel and cobalt from laterite ore.     

Recovery of germanium and other valuable metals from zinc plant residues

The main purpose of this study was to characterize and to extract germanium from the copper cake of Çinkur Zinc Plant. The physical, chemical and mineralogical characterization of the ground copper cake sample obtained from Çinkur showed that it was 84% below 147 μm containing 700 ppm germanium. The copper cake also contained 15.33% Cu, 15.63% Zn, 1.66% Cd, 1.33% Ni, 0.64% Co, 0.35% Fe, 2.62% Pb, 12.6% As, 0.18% Sb and 3.42% SiO₂. The mineralogical analysis indicated the complex nature of the copper cake which was mainly composed of metallic and oxidized phases containing copper, arsenic, zinc, cadmium, etc. The sulfuric acid leaching experiments were performed under the laboratory conditions. The optimum collective extraction of germanium and other valuable metals was obtained at a temperature range 60 to 85 °C for a leaching duration of 1 h with sulfuric acid concentration of 150 gpl and using a solid–liquid ratio 1/8 g/cc. Under these conditions, the recovery of germanium was 92.7% while the other metals were leached almost completely. The optimum selective leaching conditions of germanium was determined as half an hour leach duration, 1/8 g/cc solid–liquid ratio, 100 gpl sulfuric acid concentration and a temperature range 40 to 60 °C. Under these conditions the leach recovery of germanium was 78%. The dissolution's of other metals like cobalt, nickel, iron, copper, cadmium and arsenic were almost low. So, germanium would be separated more selectively at the following precipitation by tannin stage.     

从彩钼铅粗精矿碱性浸出液中提取钼的新工艺

研究了一种从彩钼铅粗精矿碱性浸出液中回收钼的新工艺。该工艺涉及镁盐除硅、N235萃取钼、氨水溶液反萃取钼、盐酸沉淀钼等工序。试验结果表明:在溶液中ρ(Mo)=9.2g/L、ρ(SiO2)=1.01g/L,除硅温度75℃,pH=8.5,反应1h,氯化镁加入量为理论量4倍条件下,除硅率达87.31%;以15%N235-10%仲辛醇-75%煤油溶液作为萃取剂、在Va∶Vo=2.5∶1、pH为1.7～2.0条件下,混合萃取3min,钼的3级逆流萃取率为99.55%;经反萃取和沉淀钼,最终获得钼质量分数64%以上的氧化钼产品。该工艺钼回收率高,除硅效果较好。     

高铅钼钒物料中分离钼钒的工艺研究

试验用碳酸钠与氧化后的物料进行钠化焙烧,然后水浸出其中的钼、钒。在浸出过程中,钼钒进入溶液,而其它元素Ph、Ag等则留在浸出渣中,并获得了浸出的最佳条件。同时研究了在钼钒溶液中分离钼、钒的工艺,其中用氯化铵沉淀钒,用盐酸与氯化铵沉淀钼。钼的沉淀率大于95％,钒的沉淀率大于92％。     

微波强化焙烧一水硬铝石矿浸出效果的研究

将一水硬铝石矿和过饱和NaOH溶液及Ca(OH)2混合均匀后在微波炉中焙烧;熟料在稀碱性溶液、常压条件下进行浸出,研究配料比、微波焙烧和浸出条件对熟料浸出的影响。结果表明,升高微波焙烧温度、增加NaOH添加量或延长保温时间都可以增加熟料中铝的浸出率。增加Ca(OH)2添加量可以减少溶液中硅的溶解率,但过量添加Ca(OH)2反而导致铝的浸出效果下降。适当的浸出温度和浸出时间有助于减少二次反应发生,提高铝硅分离效果。在最优试验条件下,熟料中铝、钠、硅的浸出率分别达到96.7%、98.2%和7.05%,赤泥颗粒较大,利于过滤过程。     

Decomposition of mixed rare earth concentrate by NaOH roasting and kinetics of hydrochloric acid leaching process

A new clean extraction technology for the decomposition of Bayan Obo mixed rare earth concentrate by NaOH roasting is proposed.The process mainly includes NaOH roasting to decompose rare earth concentrate and HCl leaching roasted ore.The effects of roasting temperature,roasting time,NaOH addition amount on the extraction of rare earth and factors such as HCl concentration,liquid-solid ratio,leaching temperature and leaching time on the dissolution kinetics of roasted ore were studied.The experimental results show that when the roasting temperature is 550℃ and the roasting time is 60 min,the mass ratio of NaOH:rare earth concentrate is 0.60:1,the concentration of HCl is 6.0 mol/L,the ratio of liquid to solid(L/S) 6.0:1.0,and the leaching temperature 90℃,leaching time 45 min,stirring speed 200 r/min,and the extraction of rare earth can reach 92.5%.The relevant experimental data show that the process of HCl leaching roasted ore conforms to the shrinking core model,but the control mechanism of the che mical reaction process is different when the leaching temperature is different.When the leaching temperature is between 40 and 70℃,the chemical reaction process is controlled by the diffusion of the product through the residual layer of the inert material.The average surface activation energy of the rare earth element is E_a=9.96 kJ/mol.When the leaching temperature is 75-90℃,the chemical reaction process is controlled by the interface transfer across the product layer(product layer interface mass transfer) and diffusion.The average surface activation energy of rare earth elements is E_a=41.65 kJ/mol.The results of this study have certain significance for the green extraction of mixed rare earth ore.     

含钒矿石提取钒的绿色工艺研究

以钒矿石为原料,加入自制添加剂A,经焙烧、一段水浸、二段过氧化氢浸出提取钒。结果表明,该钒矿样混入8％添加剂A于810℃焙烧4h、常温下水浸8h后再用1％过氧化氢浸24h,钒浸出率达70％以上。该方法得到较高浸出率的同时也大大减少了对环境的污染,应用前景广阔。     

采用N235从镍钼矿盐酸浸出液中萃取钼的研究

采用N235对镍钼矿盐酸浸出液进行了萃取钼的研究。试验结果表明,在最佳工艺条件下,5级逆流钼萃取率可达98%以上,镍损失率小于1%,负载有机相经稀酸洗涤除铁后采用氨水反萃,1级反萃率达97%以上,反萃液钼浓度为50 g/L左右,达到了钼镍分离及钼富集转型的目的。