基于选冶结合的镍钼矿提钼新工艺研究
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摘要
镍钼矿是我国特有的复杂多金属矿物资源,矿中有价金属镍和钼储量巨大,具有很高的经济价值。由于矿中镍和钼的品位相对较低,造成现有冶炼工艺生产成本偏高。
中南大学着眼于选冶结合开展工作。在选矿方面,已成功地实现了镍钼矿的浮选富集。经选矿处理后,粗精矿中有价元素钼、镍等品位约为原矿的三倍,原矿中约80%的钼主要以硫化钼形态富集在精矿中,其余部分钼则主要以氧化物形态进入尾矿中。本课题作为冶金研究部分,以浮选后的富钼粗精矿为对象,开发了一条制备钼酸铵产品的全湿法新工艺。
主要研究内容包括富钼精矿联合浸出镍和钼、溶剂萃取法从浸出液中提钼、铵镁盐沉淀法从钼酸铵溶液除砷、离子交换法从钼酸铵溶液除镁、酸沉结晶钼酸铵产品。最终钼酸铵产品达到了GB/T3460-2007产品标准。与现有镍钼矿的处理工艺相比,该工艺具有有价金属回收率高、环境友好以及生产成本低的特点。主要研究结果如下：
研究了富钼精矿空气氧化浸出过程的动力学。结果表明：钼浸出过程分为两个阶段,反应初期20min内,钼浸出率迅速达到20%左右,反应速度受温度及氢氧化钠浓度的影响较小；第二阶段钼浸出速度随反应温度及氢氧化钠浓度的增大而加快,且在不同的温度范围内体现出不同的动力学特征：反应温度低于60℃时,浸出过程受化学反应控制,表观活化能为62.2kJ/mol,表观反应级数为0.73；当温度大于60℃时,浸出过程受扩散和化学反应混合控制,反应表观活化能为22.3kJ/mol,表观反应级数为0.64。
在此基础上进行了浸出工艺研究。浸出过程中空气高度弥散加强了氧的溶解以及扩散传质,强化了钼的浸出,精矿中绝大部分钼进入浸出液中。空气氧化浸出过程中,增大氢氧化钠用量、温度、浸出时间、空气流量及液固比,有利于钼和硒的浸出。针对精矿中部分晶化程度高的硫化钼难以被空气氧化浸出(渣中钼含量仍为1%左右)的问题,将这部分残留钼的提取和后续提镍过程有机结合,在氯酸钠氧化时实现同时浸出。最佳工艺条件下,镍和钼的总浸出率分别为96.7%和98.3%,硒的浸出率约为50%。
富钼精矿浸出液钼浓度较低([Mo]=5～10g/L),且含有大量硫的低价氧化物(SO32-, S2O32-)及砷、磷等杂质离子。采用强碱性树脂或萃取剂从碱性浸出液中吸附钼时,杂质离子与钼酸根竞争吸附,树脂工作交换容量很低。根据钼的溶液化学特点,将浸出液调至弱酸性,以使钼聚合成与有机相亲和力较大的多酸离子。但研究发现钼会进一步被硫的低价氧化物还原成为钼蓝,相应采用了有利于钼蓝扩散传质的萃取工艺处理,使钼蓝优先被叔胺N-235萃取,而其他离子则留在萃余液中,实现了钼的选择性分离。当萃取剂浓度为15v%、母液pH值为3、相比O/A为1/4、母液中钼浓度为5.72gm,钼萃取率高达99.43%。负载有机相用15m%的氨水反萃,得到含钼125.8gm的钼酸铵溶液。研究还发现,浸出液酸化时,溶液中的Se032-易被硫的低价氧化物还原为单质硒而析出。当溶液中硒浓度为0.15g/L、pH值为2.5、室温下反应30分钟,硒还原率为79.6%。
溶剂萃取时,砷与钼形成杂多蓝一起进入有机相中,导致氨水反萃得到的钼酸铵溶液中砷等杂质含量较高。针对钼的氨性溶液的特点,采用铵镁盐沉淀法从溶液中除砷。当溶液中钼浓度为96.8gm、砷浓度为8.75gm,氯化镁用量为理论量的1.2倍、室温下反应30min,溶液中残留的砷和镁的浓度分别为0.046gm和0.52g/L,钼损失率为0.34%。
针对钼酸铵溶液除砷后残留镁,容易导致钼酸铵产品中镁含量超标的问题,采用724型阳离子树脂吸附除镁。在接触时间为30min时,1倍体积的树脂可以处理13.2倍体积的料液,树脂的工作交换容量为6.58mg/mli：湿树脂)。负镁树脂可用5倍树脂体积的2mol/L的盐酸完全解析镁。
除杂后的钼酸铵溶液通过酸沉结晶法制备钼酸铵产品,四钼酸铵结晶率为95.6%。钼酸铵产品达到了GB/T3460-2007-MSA-2标准。酸沉结晶母液钼浓度约3g/L,可返回溶剂萃取过程回收钼。
Ni-Mo ore is a polymetallic mineral resource which is uniquely in China. Ni-Mo ore contains huge amounts of molybdenum and nickel with high economic value. Since the grades of Ni and Mo in the ore are relatively low, it causes higher costs of present metallurgical processes.
In Central South University, scholars focused on the combined process of beneficiation and hydrometallurgy of Ni-Mo ore and realized the flotation of the ore successfully. After flotation, the grade of Mo and Ni was enriched and reached3times raw Ni-Mo ore. Besides, molybdenum oxide and molybdenum sulfide can be separated smoothly. In the Ni-Mo ore, about80%molybdenum entered into the molybdenum concentrate in the form molybdenum sulfide, while the rest of molybdenum went into talling being as molybdenum oxide. In this study, with the molybdenum concentrate as a raw material, a novel hydrometallurgical process was developed to prepare ammonium molybdate tetrahydrate.
The research mainly focused on joint leaching of molybdenum concentrate, molybdenum extraction from the leaching solution with solvent extraction, arsenic removal from ammonium molybdate solution by chemical precipitation with magnesium salt, magnesium removal from the ammonium molybdate solution by ion exchange and crystallization of ammonium tetramolybdate by acidification. The final product meets the standard of GB/T3460-2007(ammonium polymolybdate). Compared with current treatment processes of Ni-Mo ore, this technology has lots of advantages such as higher valuable metal recovery, environmental-friendliness and lower production cost. Main results are as follows:
Kinetics of atmospheric leaching molybdenum from the concentrate by air oxidation was studied. The results show that the leaching process consists of the following two stages. Within the initial20minutes, the leaching ratio of molybdenum reached to20%quickly. The leaching rate was not influenced obviously by temperature and the concentration of sodium hydroxide. At the second stage, the leaching rate of Molybdenum increased gradually with the increase of temperature and sodium hydroxide concentration. Moreover, different kinetic charcateristics of the leaching process was presented at different temperature ranges. When the temperature was below60℃, the leaching process was controlled by chemical reaction with the apparent activation energy of62.22kJ/mol and apparent reaction order of0.73. While the temperature was above60℃, the leaching process belongs to hybrid control with apparent activation energy of62.22kJ/mol and apparent reaction order of0.73.
Based on the kinetic of leaching process, different condition experiments were carried out. The results show that the leaching rate of molybdenum was accelerated by means of high distribution of air to intensify dissolution of oxygen and diffusion of the leaching system, and most of molybdenum in the ore entered into the leaching solution. The results show that the increase of alkali dosage, temperature, time, gas flow rate and liquid-solid ratio is beneficial to the leaching of molybdenum and selenium in the process. As the molybdenum concentrate contains parts of high crystallized molybdenum sulfide, which is difficult to be oxidized by air, the content of molybdenum in leaching residue will exceed1%even though excessive dosage of sodium hydroxide was provided. In order to improve molybdenum recovery, the leaching residue was leached in the sulphuric solution by adding sodium chlorate to extract molybdenum and nickel simultaneously. Under optimum experimental conditions, the total leaching ratio of Mo, Ni and Se were98.3%,96.7%and50%, respectively.
In the leaching solution of molybdenum concentrate, the content of Mo is low while it contains much low valence sulfur oxides (SO32-, S2O32-) as well as As and P. When extract molybdenum from the leaching solution with basic resin or extractant, competitive adsorption of MoO42-and impurities result in low operating exchange capacity. According to the chemical characteristic of molybdenum, by acidizing the leaching solution,Mo042-can transform into heptamolybdateand which has high affinity to organic phaseand can bereduced to molybdenum by low valence sulfur oxides. When tertiary amine N-235was used the extractant of molybdenum, molybdenum blue was extracted preferentially and other impurities remained in the raffinate, so selective separation of molybdenum was achieved. In the process of solvent extraction, the extraction efficiency of Mo was99.43%on the condition of molybdenum concentration of5.72g/L, pH3, time2min and O/A ratio1:4with15v%N-235. The loaded organic phase was stripped with15m%ammonia solution. The concentration of Mo in the strip liquor was125.82g/L. In addition, SeO32-was apt to be reduced by low valence sulphur oxide to element selenium in the acidification process of leaching solution. Under the optimum conditions:pH2.5, temperature297K, time30min, the reduction ratio of selenium reached79.6%.
In solvent extraction process, hteropoly-molybdenum blue formed by arsenic and molybdenum entered into the organic phase, which was stripped with ammonia solution. Thus, arsenic concentration in the strip liquor was high. It was apt to remove arsenic from the ammonium molybdate solution by chemical precipitation with magnesium due to high ammonium concentration. Experimental results show that when molybdenum concentration and arsenic concentration in the ammonium molybdate solution were96.8g/L and8.75g/L, respectively, the dosage of magnesium chloride was1.2times of stoichiometric quantity, the reaction temperature was297K and time was30min, the residual arsenic concentration and magnesium concentration were0.046g/L and0.52g/L respectively, and the lost ratio of Mo was0.34%.
Since the residual magnesium in the ammonium molybdate solution can result in magnesium content over standard in the ammonium tetramolybdate product, the excess magnesium was removed by ion exchange with cationic resin724. Experimental results show that the operating adsorption capacity of the resin was6.58mg/ml (wet resin) until the volume ratio of effluent/resin was around13.2when magnesium concentration was0.5g/L, pH value was9.2and contact time was30min. The loaded resin can be stripped with2mol/L HC1with5times the volume of loaded resin.
Ammonium molybdate tetrahydrate was obtained by acidic precipitation and crystallization after remove impurities from ammonium molybdate solution. Its crystallization ratio is95.6%. The final product conforms to the standard of GB/T3460-2007-MSA-2. The Mo concentration in the mother liquor after acidic precipitation and crystallization is around3g/L, and the mother liquor can be return to the process of solvent extraction for Mo recovery.
中南大学着眼于选冶结合开展工作。在选矿方面,已成功地实现了镍钼矿的浮选富集。经选矿处理后,粗精矿中有价元素钼、镍等品位约为原矿的三倍,原矿中约80%的钼主要以硫化钼形态富集在精矿中,其余部分钼则主要以氧化物形态进入尾矿中。本课题作为冶金研究部分,以浮选后的富钼粗精矿为对象,开发了一条制备钼酸铵产品的全湿法新工艺。
主要研究内容包括富钼精矿联合浸出镍和钼、溶剂萃取法从浸出液中提钼、铵镁盐沉淀法从钼酸铵溶液除砷、离子交换法从钼酸铵溶液除镁、酸沉结晶钼酸铵产品。最终钼酸铵产品达到了GB/T3460-2007产品标准。与现有镍钼矿的处理工艺相比,该工艺具有有价金属回收率高、环境友好以及生产成本低的特点。主要研究结果如下：
研究了富钼精矿空气氧化浸出过程的动力学。结果表明：钼浸出过程分为两个阶段,反应初期20min内,钼浸出率迅速达到20%左右,反应速度受温度及氢氧化钠浓度的影响较小；第二阶段钼浸出速度随反应温度及氢氧化钠浓度的增大而加快,且在不同的温度范围内体现出不同的动力学特征：反应温度低于60℃时,浸出过程受化学反应控制,表观活化能为62.2kJ/mol,表观反应级数为0.73；当温度大于60℃时,浸出过程受扩散和化学反应混合控制,反应表观活化能为22.3kJ/mol,表观反应级数为0.64。
在此基础上进行了浸出工艺研究。浸出过程中空气高度弥散加强了氧的溶解以及扩散传质,强化了钼的浸出,精矿中绝大部分钼进入浸出液中。空气氧化浸出过程中,增大氢氧化钠用量、温度、浸出时间、空气流量及液固比,有利于钼和硒的浸出。针对精矿中部分晶化程度高的硫化钼难以被空气氧化浸出(渣中钼含量仍为1%左右)的问题,将这部分残留钼的提取和后续提镍过程有机结合,在氯酸钠氧化时实现同时浸出。最佳工艺条件下,镍和钼的总浸出率分别为96.7%和98.3%,硒的浸出率约为50%。
富钼精矿浸出液钼浓度较低([Mo]=5～10g/L),且含有大量硫的低价氧化物(SO32-, S2O32-)及砷、磷等杂质离子。采用强碱性树脂或萃取剂从碱性浸出液中吸附钼时,杂质离子与钼酸根竞争吸附,树脂工作交换容量很低。根据钼的溶液化学特点,将浸出液调至弱酸性,以使钼聚合成与有机相亲和力较大的多酸离子。但研究发现钼会进一步被硫的低价氧化物还原成为钼蓝,相应采用了有利于钼蓝扩散传质的萃取工艺处理,使钼蓝优先被叔胺N-235萃取,而其他离子则留在萃余液中,实现了钼的选择性分离。当萃取剂浓度为15v%、母液pH值为3、相比O/A为1/4、母液中钼浓度为5.72gm,钼萃取率高达99.43%。负载有机相用15m%的氨水反萃,得到含钼125.8gm的钼酸铵溶液。研究还发现,浸出液酸化时,溶液中的Se032-易被硫的低价氧化物还原为单质硒而析出。当溶液中硒浓度为0.15g/L、pH值为2.5、室温下反应30分钟,硒还原率为79.6%。
溶剂萃取时,砷与钼形成杂多蓝一起进入有机相中,导致氨水反萃得到的钼酸铵溶液中砷等杂质含量较高。针对钼的氨性溶液的特点,采用铵镁盐沉淀法从溶液中除砷。当溶液中钼浓度为96.8gm、砷浓度为8.75gm,氯化镁用量为理论量的1.2倍、室温下反应30min,溶液中残留的砷和镁的浓度分别为0.046gm和0.52g/L,钼损失率为0.34%。
针对钼酸铵溶液除砷后残留镁,容易导致钼酸铵产品中镁含量超标的问题,采用724型阳离子树脂吸附除镁。在接触时间为30min时,1倍体积的树脂可以处理13.2倍体积的料液,树脂的工作交换容量为6.58mg/mli：湿树脂)。负镁树脂可用5倍树脂体积的2mol/L的盐酸完全解析镁。
除杂后的钼酸铵溶液通过酸沉结晶法制备钼酸铵产品,四钼酸铵结晶率为95.6%。钼酸铵产品达到了GB/T3460-2007-MSA-2标准。酸沉结晶母液钼浓度约3g/L,可返回溶剂萃取过程回收钼。
Ni-Mo ore is a polymetallic mineral resource which is uniquely in China. Ni-Mo ore contains huge amounts of molybdenum and nickel with high economic value. Since the grades of Ni and Mo in the ore are relatively low, it causes higher costs of present metallurgical processes.
In Central South University, scholars focused on the combined process of beneficiation and hydrometallurgy of Ni-Mo ore and realized the flotation of the ore successfully. After flotation, the grade of Mo and Ni was enriched and reached3times raw Ni-Mo ore. Besides, molybdenum oxide and molybdenum sulfide can be separated smoothly. In the Ni-Mo ore, about80%molybdenum entered into the molybdenum concentrate in the form molybdenum sulfide, while the rest of molybdenum went into talling being as molybdenum oxide. In this study, with the molybdenum concentrate as a raw material, a novel hydrometallurgical process was developed to prepare ammonium molybdate tetrahydrate.
The research mainly focused on joint leaching of molybdenum concentrate, molybdenum extraction from the leaching solution with solvent extraction, arsenic removal from ammonium molybdate solution by chemical precipitation with magnesium salt, magnesium removal from the ammonium molybdate solution by ion exchange and crystallization of ammonium tetramolybdate by acidification. The final product meets the standard of GB/T3460-2007(ammonium polymolybdate). Compared with current treatment processes of Ni-Mo ore, this technology has lots of advantages such as higher valuable metal recovery, environmental-friendliness and lower production cost. Main results are as follows:
Kinetics of atmospheric leaching molybdenum from the concentrate by air oxidation was studied. The results show that the leaching process consists of the following two stages. Within the initial20minutes, the leaching ratio of molybdenum reached to20%quickly. The leaching rate was not influenced obviously by temperature and the concentration of sodium hydroxide. At the second stage, the leaching rate of Molybdenum increased gradually with the increase of temperature and sodium hydroxide concentration. Moreover, different kinetic charcateristics of the leaching process was presented at different temperature ranges. When the temperature was below60℃, the leaching process was controlled by chemical reaction with the apparent activation energy of62.22kJ/mol and apparent reaction order of0.73. While the temperature was above60℃, the leaching process belongs to hybrid control with apparent activation energy of62.22kJ/mol and apparent reaction order of0.73.
Based on the kinetic of leaching process, different condition experiments were carried out. The results show that the leaching rate of molybdenum was accelerated by means of high distribution of air to intensify dissolution of oxygen and diffusion of the leaching system, and most of molybdenum in the ore entered into the leaching solution. The results show that the increase of alkali dosage, temperature, time, gas flow rate and liquid-solid ratio is beneficial to the leaching of molybdenum and selenium in the process. As the molybdenum concentrate contains parts of high crystallized molybdenum sulfide, which is difficult to be oxidized by air, the content of molybdenum in leaching residue will exceed1%even though excessive dosage of sodium hydroxide was provided. In order to improve molybdenum recovery, the leaching residue was leached in the sulphuric solution by adding sodium chlorate to extract molybdenum and nickel simultaneously. Under optimum experimental conditions, the total leaching ratio of Mo, Ni and Se were98.3%,96.7%and50%, respectively.
In the leaching solution of molybdenum concentrate, the content of Mo is low while it contains much low valence sulfur oxides (SO32-, S2O32-) as well as As and P. When extract molybdenum from the leaching solution with basic resin or extractant, competitive adsorption of MoO42-and impurities result in low operating exchange capacity. According to the chemical characteristic of molybdenum, by acidizing the leaching solution,Mo042-can transform into heptamolybdateand which has high affinity to organic phaseand can bereduced to molybdenum by low valence sulfur oxides. When tertiary amine N-235was used the extractant of molybdenum, molybdenum blue was extracted preferentially and other impurities remained in the raffinate, so selective separation of molybdenum was achieved. In the process of solvent extraction, the extraction efficiency of Mo was99.43%on the condition of molybdenum concentration of5.72g/L, pH3, time2min and O/A ratio1:4with15v%N-235. The loaded organic phase was stripped with15m%ammonia solution. The concentration of Mo in the strip liquor was125.82g/L. In addition, SeO32-was apt to be reduced by low valence sulphur oxide to element selenium in the acidification process of leaching solution. Under the optimum conditions:pH2.5, temperature297K, time30min, the reduction ratio of selenium reached79.6%.
In solvent extraction process, hteropoly-molybdenum blue formed by arsenic and molybdenum entered into the organic phase, which was stripped with ammonia solution. Thus, arsenic concentration in the strip liquor was high. It was apt to remove arsenic from the ammonium molybdate solution by chemical precipitation with magnesium due to high ammonium concentration. Experimental results show that when molybdenum concentration and arsenic concentration in the ammonium molybdate solution were96.8g/L and8.75g/L, respectively, the dosage of magnesium chloride was1.2times of stoichiometric quantity, the reaction temperature was297K and time was30min, the residual arsenic concentration and magnesium concentration were0.046g/L and0.52g/L respectively, and the lost ratio of Mo was0.34%.
Since the residual magnesium in the ammonium molybdate solution can result in magnesium content over standard in the ammonium tetramolybdate product, the excess magnesium was removed by ion exchange with cationic resin724. Experimental results show that the operating adsorption capacity of the resin was6.58mg/ml (wet resin) until the volume ratio of effluent/resin was around13.2when magnesium concentration was0.5g/L, pH value was9.2and contact time was30min. The loaded resin can be stripped with2mol/L HC1with5times the volume of loaded resin.
Ammonium molybdate tetrahydrate was obtained by acidic precipitation and crystallization after remove impurities from ammonium molybdate solution. Its crystallization ratio is95.6%. The final product conforms to the standard of GB/T3460-2007-MSA-2. The Mo concentration in the mother liquor after acidic precipitation and crystallization is around3g/L, and the mother liquor can be return to the process of solvent extraction for Mo recovery.
引文
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