辉钼矿湿法冶金新工艺及其机理研究
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摘要
针对现有辉钼矿焙烧工艺污染严重,且难以实现钼及伴生金属高效经济回收的弊端,进行了辉钼矿常温常压湿法浸出和浸出液中钼、铼的分离富集及其机理的研究,为发展环境友好的钼湿法冶金新工艺提供理论基础。
基于含氧氯酸盐在溶液中具有高的电极电位,表现出很强氧化能力的特性,对氯酸钠和次氯酸钠氧化分解辉钼矿、黄铜矿等进行热力学计算。结果表明,氯酸钠、次氯酸钠氧化分解辉钼矿、黄铜矿等过程的△G°值均为负值,且绝对值较大,说明氯酸纳、次氯酸钠都能应用于对辉钼矿的氧化分解。
研究了氯酸钠氧化分解辉钼矿的行为与规律,发明了一种使用电解氯化钠盐水生成的氯酸钠电解液为氧化剂、在酸性水溶液条件下高效氧化分解辉钼矿精矿的新方法。以氯酸钠为氧化剂,在C(NaCl)=0.71 mol/L,C(H2SO4)=73.3 g/L,NaClO3／MoS2=3.6(mol/mol),L／S=30,T=70℃,搅拌速率为400 r/min, t=6 h的条件下,钼的浸出率为98.62%,铼的浸出率为99.52%,铜的浸出率为99.77%,铁的浸出率为99.28%,实现了多种硫化矿物的全分解。在对氯酸钠氧化分解辉钼矿研究的基础上,将电化学合成的氯酸钠电解液,不经净化、结晶等操作,调整酸度后直接用来浸出辉钼矿。优化确定采用的氯酸钠电解合成工艺条件为：槽电压3.0V,电流密度653.2 A／m2,搅拌速度100 r/min,Na2Cr2O7浓度3 g/L, pH=6.7,温度75℃,电解时间35h,此时电解液含氯酸钠290.90g/L,含次氯酸钠7.08g/L,钼、铼浸出率分别为98.51%和99.56%。采用氯酸钠电解液湿法氧化分解辉钼矿的方法具有氧化剂生产成本低、使用安全等优点。浸出动力学研究表明浸出过程符合收缩核模型,表观活化能Ea=104.6 kJ/mol,过程受化学反应控制。
研究并揭示了辉钼矿的电氧化分解过程与机理,发明了一种应用碳酸盐-酸式碳酸盐缓冲体系的辉钼矿选择性电氧化浸出方法,实现了辉钼矿与黄铜矿等金属硫化矿物的选择性浸出,显著提高了电流效率。电氧化分解辉钼矿研究表明,pH=9时,辉钼矿分解速率最快,将碳酸盐-酸式碳酸盐缓冲体系应用到辉钼矿电氧化浸出过程中,在槽电压3.5V电流密度653.2 A／m2,液固比25：1,搅拌速度400 r/min,氯化钠浓度4mol/L,碳酸钠6g/L，碳酸氢铵5g/L,缓冲体系pH=8.5-9.5,室温的条件下,电解240min时钼浸出率可达99.35%,铼浸出率为99.79%,电流效率为62.75%,而非缓冲体系电氧化工艺达到同样钼浸出率时电流效率仅为40.44%。通过缓冲体系的应用,使钼精矿中钼、铼选择性高效氧化浸出,而黄铜矿基本不浸出,浸出渣中铜含量为10.84%,回收率达到97.93%。缓冲体系中电氧化工艺是一种更为高效环保的辉钼矿选择性电氧化分解工艺。该工艺特别适用于德兴铜矿高铜钼精矿的的湿法分解。
通过对电氧化过程阳极氧化、氧化剂变化规律的研究,建立了电氧化浸出过程的动力学模型,揭示了缓冲体系中电氧化分解辉钼矿的浸出机理。结果表明,辉钼矿在阳极上不能直接被氧化,NaCl介质中析氯反应起了主导作用,辉钼矿的氧化浸出主要是借助于阳极析出的氯气转化生成的氧化剂,氧化剂主要成分是NaClO。浸出动力学研究表明,pH=9的碳酸盐-酸式碳酸盐缓冲体系中,辉钼矿的浸出表观活化能Ea=8.56 kJ/mol,过程受固膜扩散控制。
采用溶剂萃取技术对辉钼矿浸出液中的Mo(Ⅵ)、Re(Ⅶ)进行了萃取与反萃研究,建立了萃取反应模型方程,揭示了萃取剂N235对钼、铼金属离子的萃取机理。在优化的萃取条件下,氯酸钠电解液浸出辉钼矿溶液经3级萃取后,Mo(Ⅵ)、Re(Ⅶ)的萃取率分别达到99.84%和95.19%；采用17%的氨水为反萃剂,在优化的反萃条件下,3级反萃后Mo(Ⅵ)、Re(Ⅶ)的反萃率分别达到99.90%和99.73%。在优化的萃取条件下,电氧化分解辉钼矿浸出液经3级萃取后,Mo(Ⅵ)、Re(Ⅶ)的萃取率分别达到99.77%和94.73%；采用17%的氨水为反萃剂,在优化的反萃条件下,3级反萃后Mo(Ⅵ)、Re(Ⅶ)的反萃率分别达到99.89%和99.54%。等摩尔法实验和红外光谱分析表明,N235萃取Mo(Ⅵ)、Re(Ⅶ)离子时,叔胺基与溶液中的离子发生了配位反应,分别以(R3NH2)[(MoO2)2(SO4)3]和R3NH.ReO4形式进行萃取。
采用D201树脂对反萃液中的Mo(Ⅵ)、Re(Ⅶ)吸附分离性能进行了考察,建立了树脂吸附动力学和热力学模型,揭示了阴离子交换树脂对Mo(Ⅵ)、Re(Ⅶ)的吸附机理。静态吸附实验表明,D201树脂对Mo(Ⅵ)、Re(Ⅶ)的吸附容量都随吸附时间和金属离子初始浓度的增加而增大。静态分离实验表明,树脂对溶液中的Re(Ⅶ)具有良好的吸附选择性,溶液中Mo(Ⅵ)初始浓度越大分离因子越高。对氯酸钠浸出体系反萃液,在吸附温度30℃,pH=8,吸附时间1h条件下,Re(Ⅶ)、Mo(Ⅵ)吸附率分别为92.41%、3.19%,分离因子为190.49；对电氧化浸出体系反萃液,在吸附温度30℃,pH=8,吸附时间1h条件下,Re(Ⅶ)、Mo(Ⅵ)吸附率分别为92.18%、3.46%,分离因子为169.56。实验数据符合Boyd液膜扩散方程,树脂对Mo(Ⅵ)、Re(Ⅶ)的吸附过程属于液膜扩散控制过程。吸附过程为单分子层吸附,符合Langmuir和Freundlich等温吸附模型,根据拟合结果计算得到的D201树脂对Mo(Ⅵ)、Re(Ⅶ)的饱和吸附容量分别为4.2633 mmol/g、4.2355 mmol/g。
全流程技术经济分析表明,与传统焙烧工艺相比,缓冲体系电氧化浸出工艺是一种高效环保的辉钼矿湿法分解工艺。对于德兴铜矿的辉钼精矿,钼回收率达到99.04%,铼回收率达到85.84%,铜回收率超过97%,年净增经济效益2000多万元,减排SO22681t/年,具有明显的经济效益和环境效益。
The roasting method gradually shows its shortcomings. During roasting, many valuable metals are lost due to volatilization and the SO2 generated is a source of environmental pollution. The novel cold atmospheric leaching of molybdenite, the process of extraction and separation of molybdenum and rhenium in the extract, and their mechanism have been investigated, providing theoretical basis for environment-friendly hydrometallurgy of molybdenite.
Chlorate and hypochlorite in solution have high electrode potential and strong oxidative property. Thermodynamic calculations of the oxygenolysis process of molybdenite by sodium chlorate and sodium hypochlorite were carried out. The results showed that values of△G°were negative and the absolute values of them were high if sodium chlorate and sodium hypochlorite were used as oxidant in the molybdenite leaching process. In the view of thermodynamic point, MoS2, CuFeS2, FeS2 and ReS2 etc. could be leached by acidic chlorate system and alkaline sodium hypochlorite system theoretically.
In this paper, the technology of molybdenum extraction from molybdenite concentrate by using sodium chlorate has been investigated, and a novel hydrometallurgical extraction technology of molybdenite by using sodium chlorate electrolyte was introduced. The optimized leaching conditions are as followed:C(NaCl)=0.71 mol/L, C(H2SO4)=73.3 g/L, NaClO3/MoS2=3.6 (mol/mol), L/S=30, T=70℃, stirring rate is 400 r/min, t=6 h. Under these conditions, leaching yield of molybdenum is 98.62%, leaching yield of copper is 99.77%, leaching yield of iron is 99.28%, leaching yield of rhenium,99.52%; while leaching of molybdenite is completed, Cu, Fe, Re were almost completely leached into the liquid. The acidic sodium chlorate electrolyte by electrochemical synthesis, without purification, crystallization, but just adjust the acidity and was directly used for leaching of molybdenite. The optimized electrosynthesis conditions of sodium chlorate are as followed:cell voltage of 3.0 V, current density of 653.2 A/m2, stirring rate of 100 r/min, concentration of potassium dichromate of 3 g/L, pH=6.7,75℃,35 h. Under this electrosynthesis conditions, sodium chlorate is 290.90 g/L, sodium hypochlorite is 7.08 g/L, leaching yield of molybdenum is 98.51%, and leaching yield of rhenium is 99.56%. It was an efficient and economic wet leaching process that sodium chlorate electrolyte was used for leaching molybdenite, and it can also reduce cost largely. The kinetic study showed that the process of leaching molybdenite using acid sodium chlorate system is represented by shrinking core model, the apparent activation energy (Ea) for the dissolution reaction was calculated as 104.6.6 kJ/mol showed that the chemical reaction is the control step of leaching process.
According to the characteristic of electric-oxidation leaching molybdenite, the electro-oxidation technology in CO32--HCO3-buffer system was introduced, molybdenite can be selectively oxidized, and current efficiency increases obviously. Experimental studies showed that when pH=9, the oxidation rate of molybdenite is the fastest. The optimized leaching conditions are as followed:cell voltage of 3.5 V, current density of 653.2 A/m2, liquid-solid ratio of 25:1, stirring rate of 400 r/min, concentration of sodium chloride of 4 mol/L, sodium carbonate of 6 g/L, ammonium acid carbonate of 5 g/L, pH value of buffer system of 8.5-9.5, at room temperature. Under this process conditions, the leaching yield of Mo reaches 99.35%, leaching yield of rhenium is 99.79%after leached 240 min. The current efficiency of buffer system is of 62.75%, but that of unbuffered system is only of 40.44%. In CO32--HCO3- buffer system, it restrained leaching reaction for chalcopyrite, chalcopyrite would not be leached which reduced its electricity consumption, copper content of 10.84% in residue and recovery rate of 97.93%.The application of the buffer system can significantly reduce the energy consumption for electro-oxidation leaching of molybdenite.
Anode oxidation and oxidant have been investigated, and dynamic model of leaching process was established. The results showed that molybdenite couldn't be electro-oxidated at the anode directly in the buffer system. The chlorine evolution reaction has played a dominant role in sodium chloride medium. Sodium hypochlorite is the main oxidant of oxidation leaching, which transformed from chloration liberated at the anode. The kinetic study showed that a shrinking core model is presented to describe the dissolution and to analyze the data. It was established that the leaching process is mainly controlled by diffusion through a porous product layer, the apparent activation energy of this dissolution process was found to be 8.56 kJ/mol.
The studies on extraction of Mo(Ⅵ), Re(Ⅶ) in the solution and its mechanism by using N235 extractant have been investigated. For extract solution of sodium chlorate, the percentage extraction rates of Mo(Ⅵ) and Re(Ⅶ) under the optimized extraction conditions were about 99.84%and 95.19%; Stripping of molybdenum to aqueous phase was efficient when 17% ammonia liquor were applied, the Mo(Ⅵ) and Re(Ⅶ) stripping efficiencies under the optimized stripping conditions were about 99.90% and 99.73%. For electric-oxidation extract solution, the percentage extraction rates of Mo(Ⅵ) and Re(Ⅶ) under the optimized extraction conditions were about 99.77%and 94.73%; Stripping of molybdenum to aqueous phase was efficient when 17% ammonia liquor were applied, the Mo(Ⅵ) and Re(Ⅶ) stripping efficiencies under the optimized stripping conditions were about 99.89% and 99.54%. The results of equimolar seriers method and infrared spectral analysis indicates that the coordinate reactions in aqueous solution happened while N235 extracted Mo(Ⅵ) and Re(Ⅶ) ions, (R3NH2)[(MoO2)2(SO4)3] and R3NH·Re04 come into being, respectively。
The adsorption and separation capability for Mo(Ⅵ) and Re(Ⅶ) ions of D201 resin have been investigated, adsorption kinetics and thermodynamics were revealed by analyzing adsorption data. The results showed that the adsorption capacity of D201 for Mo(Ⅵ) and Re(Ⅶ) ions increases with increase in adsorbing time and the initial ions concentration. For dualistic mixed solution of Mo(Ⅵ) and Re(Ⅶ), the separation capability of D201 for Re(Ⅶ) and Mo(Ⅵ) has been investigated, it indicates that D201 has excellent adsorption selection for Re(Ⅶ) ions, and satisfy the separation demands. For stripping solution of sodium chlorate system, the adsorption rate of Re(Ⅶ) and Mo(Ⅵ) are 92.41%,3.19%,respectively, and separating factor is 190.49,when adsorption condition is 30℃, pH=8, and absorbing time 1 h; For stripping solution of electric-oxidation system, the adsorption rate of Re(Ⅶ) and Mo(Ⅵ) are 92.18%,3.46%, respectively, and separating factor is 169.56,when adsorption condition is 30℃, pH=8, and absorbing time 1 h.The experimental data fit Boyd's diffuse equation of liquid film, indicating that the adsorption is controlled by liquid film diffuse; the isothermal adsorption obeys Langmuir and Freundlich equation, especially the former equation. The saturated adsorption capacity of D201 for Mo(Ⅵ) and Re(Ⅶ) calculated base on simulated results were 4.2633 mmol/g,4.2355 mmol/g, respectively.
The results of rough cost and economic analysis for molybdenite of Dexing Copper Mine showed that, compared to conventional roasting process, the overall yield of molybdenum is 99.04%, the overall yield of rhenium is 85.84%, increase the recovery for rhenium by 25%, and more than 97% copper can be recovered, The new-increased income is over 2000 million yuan per annum, the reduced quality of SO2 is 2681 t per annum.
基于含氧氯酸盐在溶液中具有高的电极电位,表现出很强氧化能力的特性,对氯酸钠和次氯酸钠氧化分解辉钼矿、黄铜矿等进行热力学计算。结果表明,氯酸钠、次氯酸钠氧化分解辉钼矿、黄铜矿等过程的△G°值均为负值,且绝对值较大,说明氯酸纳、次氯酸钠都能应用于对辉钼矿的氧化分解。
研究了氯酸钠氧化分解辉钼矿的行为与规律,发明了一种使用电解氯化钠盐水生成的氯酸钠电解液为氧化剂、在酸性水溶液条件下高效氧化分解辉钼矿精矿的新方法。以氯酸钠为氧化剂,在C(NaCl)=0.71 mol/L,C(H2SO4)=73.3 g/L,NaClO3／MoS2=3.6(mol/mol),L／S=30,T=70℃,搅拌速率为400 r/min, t=6 h的条件下,钼的浸出率为98.62%,铼的浸出率为99.52%,铜的浸出率为99.77%,铁的浸出率为99.28%,实现了多种硫化矿物的全分解。在对氯酸钠氧化分解辉钼矿研究的基础上,将电化学合成的氯酸钠电解液,不经净化、结晶等操作,调整酸度后直接用来浸出辉钼矿。优化确定采用的氯酸钠电解合成工艺条件为：槽电压3.0V,电流密度653.2 A／m2,搅拌速度100 r/min,Na2Cr2O7浓度3 g/L, pH=6.7,温度75℃,电解时间35h,此时电解液含氯酸钠290.90g/L,含次氯酸钠7.08g/L,钼、铼浸出率分别为98.51%和99.56%。采用氯酸钠电解液湿法氧化分解辉钼矿的方法具有氧化剂生产成本低、使用安全等优点。浸出动力学研究表明浸出过程符合收缩核模型,表观活化能Ea=104.6 kJ/mol,过程受化学反应控制。
研究并揭示了辉钼矿的电氧化分解过程与机理,发明了一种应用碳酸盐-酸式碳酸盐缓冲体系的辉钼矿选择性电氧化浸出方法,实现了辉钼矿与黄铜矿等金属硫化矿物的选择性浸出,显著提高了电流效率。电氧化分解辉钼矿研究表明,pH=9时,辉钼矿分解速率最快,将碳酸盐-酸式碳酸盐缓冲体系应用到辉钼矿电氧化浸出过程中,在槽电压3.5V电流密度653.2 A／m2,液固比25：1,搅拌速度400 r/min,氯化钠浓度4mol/L,碳酸钠6g/L，碳酸氢铵5g/L,缓冲体系pH=8.5-9.5,室温的条件下,电解240min时钼浸出率可达99.35%,铼浸出率为99.79%,电流效率为62.75%,而非缓冲体系电氧化工艺达到同样钼浸出率时电流效率仅为40.44%。通过缓冲体系的应用,使钼精矿中钼、铼选择性高效氧化浸出,而黄铜矿基本不浸出,浸出渣中铜含量为10.84%,回收率达到97.93%。缓冲体系中电氧化工艺是一种更为高效环保的辉钼矿选择性电氧化分解工艺。该工艺特别适用于德兴铜矿高铜钼精矿的的湿法分解。
通过对电氧化过程阳极氧化、氧化剂变化规律的研究,建立了电氧化浸出过程的动力学模型,揭示了缓冲体系中电氧化分解辉钼矿的浸出机理。结果表明,辉钼矿在阳极上不能直接被氧化,NaCl介质中析氯反应起了主导作用,辉钼矿的氧化浸出主要是借助于阳极析出的氯气转化生成的氧化剂,氧化剂主要成分是NaClO。浸出动力学研究表明,pH=9的碳酸盐-酸式碳酸盐缓冲体系中,辉钼矿的浸出表观活化能Ea=8.56 kJ/mol,过程受固膜扩散控制。
采用溶剂萃取技术对辉钼矿浸出液中的Mo(Ⅵ)、Re(Ⅶ)进行了萃取与反萃研究,建立了萃取反应模型方程,揭示了萃取剂N235对钼、铼金属离子的萃取机理。在优化的萃取条件下,氯酸钠电解液浸出辉钼矿溶液经3级萃取后,Mo(Ⅵ)、Re(Ⅶ)的萃取率分别达到99.84%和95.19%；采用17%的氨水为反萃剂,在优化的反萃条件下,3级反萃后Mo(Ⅵ)、Re(Ⅶ)的反萃率分别达到99.90%和99.73%。在优化的萃取条件下,电氧化分解辉钼矿浸出液经3级萃取后,Mo(Ⅵ)、Re(Ⅶ)的萃取率分别达到99.77%和94.73%；采用17%的氨水为反萃剂,在优化的反萃条件下,3级反萃后Mo(Ⅵ)、Re(Ⅶ)的反萃率分别达到99.89%和99.54%。等摩尔法实验和红外光谱分析表明,N235萃取Mo(Ⅵ)、Re(Ⅶ)离子时,叔胺基与溶液中的离子发生了配位反应,分别以(R3NH2)[(MoO2)2(SO4)3]和R3NH.ReO4形式进行萃取。
采用D201树脂对反萃液中的Mo(Ⅵ)、Re(Ⅶ)吸附分离性能进行了考察,建立了树脂吸附动力学和热力学模型,揭示了阴离子交换树脂对Mo(Ⅵ)、Re(Ⅶ)的吸附机理。静态吸附实验表明,D201树脂对Mo(Ⅵ)、Re(Ⅶ)的吸附容量都随吸附时间和金属离子初始浓度的增加而增大。静态分离实验表明,树脂对溶液中的Re(Ⅶ)具有良好的吸附选择性,溶液中Mo(Ⅵ)初始浓度越大分离因子越高。对氯酸钠浸出体系反萃液,在吸附温度30℃,pH=8,吸附时间1h条件下,Re(Ⅶ)、Mo(Ⅵ)吸附率分别为92.41%、3.19%,分离因子为190.49；对电氧化浸出体系反萃液,在吸附温度30℃,pH=8,吸附时间1h条件下,Re(Ⅶ)、Mo(Ⅵ)吸附率分别为92.18%、3.46%,分离因子为169.56。实验数据符合Boyd液膜扩散方程,树脂对Mo(Ⅵ)、Re(Ⅶ)的吸附过程属于液膜扩散控制过程。吸附过程为单分子层吸附,符合Langmuir和Freundlich等温吸附模型,根据拟合结果计算得到的D201树脂对Mo(Ⅵ)、Re(Ⅶ)的饱和吸附容量分别为4.2633 mmol/g、4.2355 mmol/g。
全流程技术经济分析表明,与传统焙烧工艺相比,缓冲体系电氧化浸出工艺是一种高效环保的辉钼矿湿法分解工艺。对于德兴铜矿的辉钼精矿,钼回收率达到99.04%,铼回收率达到85.84%,铜回收率超过97%,年净增经济效益2000多万元,减排SO22681t/年,具有明显的经济效益和环境效益。
The roasting method gradually shows its shortcomings. During roasting, many valuable metals are lost due to volatilization and the SO2 generated is a source of environmental pollution. The novel cold atmospheric leaching of molybdenite, the process of extraction and separation of molybdenum and rhenium in the extract, and their mechanism have been investigated, providing theoretical basis for environment-friendly hydrometallurgy of molybdenite.
Chlorate and hypochlorite in solution have high electrode potential and strong oxidative property. Thermodynamic calculations of the oxygenolysis process of molybdenite by sodium chlorate and sodium hypochlorite were carried out. The results showed that values of△G°were negative and the absolute values of them were high if sodium chlorate and sodium hypochlorite were used as oxidant in the molybdenite leaching process. In the view of thermodynamic point, MoS2, CuFeS2, FeS2 and ReS2 etc. could be leached by acidic chlorate system and alkaline sodium hypochlorite system theoretically.
In this paper, the technology of molybdenum extraction from molybdenite concentrate by using sodium chlorate has been investigated, and a novel hydrometallurgical extraction technology of molybdenite by using sodium chlorate electrolyte was introduced. The optimized leaching conditions are as followed:C(NaCl)=0.71 mol/L, C(H2SO4)=73.3 g/L, NaClO3/MoS2=3.6 (mol/mol), L/S=30, T=70℃, stirring rate is 400 r/min, t=6 h. Under these conditions, leaching yield of molybdenum is 98.62%, leaching yield of copper is 99.77%, leaching yield of iron is 99.28%, leaching yield of rhenium,99.52%; while leaching of molybdenite is completed, Cu, Fe, Re were almost completely leached into the liquid. The acidic sodium chlorate electrolyte by electrochemical synthesis, without purification, crystallization, but just adjust the acidity and was directly used for leaching of molybdenite. The optimized electrosynthesis conditions of sodium chlorate are as followed:cell voltage of 3.0 V, current density of 653.2 A/m2, stirring rate of 100 r/min, concentration of potassium dichromate of 3 g/L, pH=6.7,75℃,35 h. Under this electrosynthesis conditions, sodium chlorate is 290.90 g/L, sodium hypochlorite is 7.08 g/L, leaching yield of molybdenum is 98.51%, and leaching yield of rhenium is 99.56%. It was an efficient and economic wet leaching process that sodium chlorate electrolyte was used for leaching molybdenite, and it can also reduce cost largely. The kinetic study showed that the process of leaching molybdenite using acid sodium chlorate system is represented by shrinking core model, the apparent activation energy (Ea) for the dissolution reaction was calculated as 104.6.6 kJ/mol showed that the chemical reaction is the control step of leaching process.
According to the characteristic of electric-oxidation leaching molybdenite, the electro-oxidation technology in CO32--HCO3-buffer system was introduced, molybdenite can be selectively oxidized, and current efficiency increases obviously. Experimental studies showed that when pH=9, the oxidation rate of molybdenite is the fastest. The optimized leaching conditions are as followed:cell voltage of 3.5 V, current density of 653.2 A/m2, liquid-solid ratio of 25:1, stirring rate of 400 r/min, concentration of sodium chloride of 4 mol/L, sodium carbonate of 6 g/L, ammonium acid carbonate of 5 g/L, pH value of buffer system of 8.5-9.5, at room temperature. Under this process conditions, the leaching yield of Mo reaches 99.35%, leaching yield of rhenium is 99.79%after leached 240 min. The current efficiency of buffer system is of 62.75%, but that of unbuffered system is only of 40.44%. In CO32--HCO3- buffer system, it restrained leaching reaction for chalcopyrite, chalcopyrite would not be leached which reduced its electricity consumption, copper content of 10.84% in residue and recovery rate of 97.93%.The application of the buffer system can significantly reduce the energy consumption for electro-oxidation leaching of molybdenite.
Anode oxidation and oxidant have been investigated, and dynamic model of leaching process was established. The results showed that molybdenite couldn't be electro-oxidated at the anode directly in the buffer system. The chlorine evolution reaction has played a dominant role in sodium chloride medium. Sodium hypochlorite is the main oxidant of oxidation leaching, which transformed from chloration liberated at the anode. The kinetic study showed that a shrinking core model is presented to describe the dissolution and to analyze the data. It was established that the leaching process is mainly controlled by diffusion through a porous product layer, the apparent activation energy of this dissolution process was found to be 8.56 kJ/mol.
The studies on extraction of Mo(Ⅵ), Re(Ⅶ) in the solution and its mechanism by using N235 extractant have been investigated. For extract solution of sodium chlorate, the percentage extraction rates of Mo(Ⅵ) and Re(Ⅶ) under the optimized extraction conditions were about 99.84%and 95.19%; Stripping of molybdenum to aqueous phase was efficient when 17% ammonia liquor were applied, the Mo(Ⅵ) and Re(Ⅶ) stripping efficiencies under the optimized stripping conditions were about 99.90% and 99.73%. For electric-oxidation extract solution, the percentage extraction rates of Mo(Ⅵ) and Re(Ⅶ) under the optimized extraction conditions were about 99.77%and 94.73%; Stripping of molybdenum to aqueous phase was efficient when 17% ammonia liquor were applied, the Mo(Ⅵ) and Re(Ⅶ) stripping efficiencies under the optimized stripping conditions were about 99.89% and 99.54%. The results of equimolar seriers method and infrared spectral analysis indicates that the coordinate reactions in aqueous solution happened while N235 extracted Mo(Ⅵ) and Re(Ⅶ) ions, (R3NH2)[(MoO2)2(SO4)3] and R3NH·Re04 come into being, respectively。
The adsorption and separation capability for Mo(Ⅵ) and Re(Ⅶ) ions of D201 resin have been investigated, adsorption kinetics and thermodynamics were revealed by analyzing adsorption data. The results showed that the adsorption capacity of D201 for Mo(Ⅵ) and Re(Ⅶ) ions increases with increase in adsorbing time and the initial ions concentration. For dualistic mixed solution of Mo(Ⅵ) and Re(Ⅶ), the separation capability of D201 for Re(Ⅶ) and Mo(Ⅵ) has been investigated, it indicates that D201 has excellent adsorption selection for Re(Ⅶ) ions, and satisfy the separation demands. For stripping solution of sodium chlorate system, the adsorption rate of Re(Ⅶ) and Mo(Ⅵ) are 92.41%,3.19%,respectively, and separating factor is 190.49,when adsorption condition is 30℃, pH=8, and absorbing time 1 h; For stripping solution of electric-oxidation system, the adsorption rate of Re(Ⅶ) and Mo(Ⅵ) are 92.18%,3.46%, respectively, and separating factor is 169.56,when adsorption condition is 30℃, pH=8, and absorbing time 1 h.The experimental data fit Boyd's diffuse equation of liquid film, indicating that the adsorption is controlled by liquid film diffuse; the isothermal adsorption obeys Langmuir and Freundlich equation, especially the former equation. The saturated adsorption capacity of D201 for Mo(Ⅵ) and Re(Ⅶ) calculated base on simulated results were 4.2633 mmol/g,4.2355 mmol/g, respectively.
The results of rough cost and economic analysis for molybdenite of Dexing Copper Mine showed that, compared to conventional roasting process, the overall yield of molybdenum is 99.04%, the overall yield of rhenium is 85.84%, increase the recovery for rhenium by 25%, and more than 97% copper can be recovered, The new-increased income is over 2000 million yuan per annum, the reduced quality of SO2 is 2681 t per annum.
引文
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