锂云母中有价金属的高效提取研究
摘要
江西省宜春市拥有全国最大的锂云母矿,现已探明Li20储量达110万吨,目前还未能有效地综合开发利用。本文在综合评述锂资源开发现状及锂云母提取锂研究现状的基础上,分别采用氯化焙烧法、硫酸盐法、压煮法,对宜春锂云母矿中的锂及伴生的钾、铷、铯进行综合提取研究,取得了较好的效果。本论文的主要研究内容及结论如下：
在热力学理论分析的基础上,对氯化焙烧-水浸法提取锂云母矿中锂、钾、铷、铯等有价金属的工艺进行了研究。结果表明,在采用复合氯化剂质量比CaCl2:NaCl为0.6：0.4,锂云母矿与氯化剂质量比为1.0,氯化焙烧温度为880℃,氯化焙烧时间为30min时,锂、钾、铷、铯的浸出率分别达到92.86%、88.49%、94.05%、93.06%。由于氯化反应过程中锂云母矿中的铝、硅可与CaCl2、NaCl中的钙和钠反应生成相应的不溶于水的铝硅酸盐,而有价碱金属锂、钾、铷、铯与氯离子结合形成相应的可溶于水的碱金属氯化物,氯化处理后所得熟料于常压下以水作浸出剂进行选择性浸出,可有效的将锂、钾、铷、铯浸出,而铝、硅等保留在浸出渣中。
对锂云母矿与固体氯化剂(CaCl2)的反应机理进行了研究。结果表明,在Ar气氛下,CaCl2和锂云母发生交互反应；在水蒸气气氛下部分氯化钙分解产生氯化氢,所以在H20蒸气气氛下锂云母与固体氯化剂发生交互反应的同时还存在与气体氯化氢反应；在O2气氛下部分CaCl2分解产生氯气,因此锂云母与CaCl2存在交互反应的同时还存在与氯气的反应。
采用硫酸盐焙烧-水浸法处理锂云母矿,可以选择性提取锂云母矿中的锂、钾、铷、铯,最佳工艺条件为：锂云矿:Na2SO4:K2SO4:CaO质量比1：0.5：0.1：0.1、焙烧温度900℃、焙烧时间30min,锂、钾、铷、铯的浸出率分别为91.61%、44.37%、29.33%和23.21%。XRD分析表明,硫酸盐焙烧可以将锂云母矿分解成NaSi3Al08、KAlSi2O6和CaAl2Si2O8。在硫酸盐焙烧过程中添加氧化钙可以有效促进锂云母矿中的氟与钙反应形成CaF2和Ca4Si2O7F2。
首次采用硫酸钠和氯化钙复合盐焙烧-水浸法处理锂云母矿。结果表明在830～930℃内锂的提取率基本不变,保持在90%以上,且焙烧物料未出现熔融现象。其最佳工艺条件为：锂云母:Na2SO4:CaCl2=1:0.5：0.3,焙烧温度850℃,焙烧时间30min,锂的浸出率达93.83%,铷、铯和钾的提取率分别为93.46%、90.08%和60.72%。
对锂云母在高温水蒸气下脱氟焙烧的条件及物理化学变化进行了系统的研究。结果表明,水蒸气气氛下于860℃反应40min,锂云母的脱氟率可达61.43%；SEM检测表明脱氟料表面变得粗糙蓬松;XRD检测表明脱氟料锂云母衍射峰几乎消失,主要物相转变成铝硅酸锂(LiAl(SiO3)2)和铝硅酸钾(KAlSi2O6)。
对硫酸钠、石灰压煮法处理脱氟锂云母工艺进行了系统的研究。结果表明,当脱氟锂云母粒度在300目以下,搅拌速度400r·min-1,反应时间3h,反应温度150℃,脱氟锂云母：硫酸钠：石灰：水=1：1：0.7：7,锂、钾、铷、铯的浸出率分别为91.98%、93.06%、70.18%、61.02%。
压煮反应过程动力学实验研究结果表明锂、钾、铷、铯浸出过程动力学符合未反应收缩核模型,属于固膜扩散控制,锂、钾、铷、铯浸出活化能分别为20.179kJ·mol-1、19.617kJ·mol-1、45274kJ·mol-1和20.179kJ·mol-1、19.617kJ·mol-1、45.274kJ·mol-1
对石灰乳压煮法处理脱氟锂云母工艺进行了系统的研究。结果表明,当脱氟锂云母球磨活化100min,石灰与脱氟锂云母质量比为1,液固比为4,压煮反应温度150℃,压煮反应时间1h,搅拌速度为400r-min-1,锂、钾、铷、铯的提取率分别为98.9%、68.67%、47.43%和43.54%。
针对氯化物、硫酸盐浸出液体系,对Li+, Na+, K+//Cl-及Li+, Na+, K+//SO42-体系相图及浸出液中制备碳酸锂方法进行了研究。结果表明氯化物体系可以通过浓缩结晶的方式将浸出液中的锂浓度提高到20～30g·L-1,而硫酸盐体系在浓缩到一定程度后会形成LiKSO4复盐而沉淀；因此Li+, Na+, K+//SO42-体系可通过冷冻到-5℃冷冻结晶,降低体系中的硫酸根浓度。去除硫酸根的溶液再次浓缩,将Li浓度提高到20-24g·L-1后与5mol/L的Na2CO3反应获得Li2CO3。
The lepidolite from Jiangxi Province (China)has a reserve of1.1million tons Li2O and accounts for30%of China's domestic proven reserves, but little of ithas been utilized. The present methods of lithium extraction from lepidolite worldwide were summarized in this paper. Comprehensive extraction and utilization of lithium, potassium, rubidium and cesium from the lepidolite of Jiangxi Province were introduced. Chlorination roasting, sulfation roasting and autoclaving method were adopted. The effect and mechanism ofmineral phase reconstruction on the extraction behavior of Li, K, Rb, Cs from lepidolite with different methods were discussed in detail and conclusions were made as follows.
Based on the analysis of thermodynamics, the experimental conditions of chlorination roasting-water leaching were studied. The conditional experiments indicated that the optimum mass ratio of lepidolite/CaCl2/NaCl is1:0.6:0.4during the roasting process. The extraction efficiencies of Li, K, Rb and Cs are92.86%,88.49%,94.05%and93.06%, respectively after roasting for0.5h at880℃. Li, K, Rb and Cs are selectively leached by chlorination roasting and water leaching at atmospheric pressure, and impurities extraction can be suppressed.
The reaction mechanism of lepidolite and solid chloridizing agent was investigated in details. The results indicated that base-exchange reaction occurs between CaCl2and lepidolite under argon atmosphere. CaCl2is decomposed intohCl under vapor atmosphere andhCl will react with lepidolite under vapor atmosphere. Because CaCl2is decomposed into Cl2under oxygen atmosphere, lepidolite will react with CaCl2and Cl2.
Sulfation roasting followed by water leaching process was used to selectively extract Li, K, Rb and Cs from lepidolite. Various operational parameters including roasting temperature, the amount of additives, and solid/liquid ratio in the leaching process were optimized. The extraction efficiencies of Li, K, Rb and Cs could reach91.61%,44.37%,29.33%and23.21%with a mass ratio of lepidolite/Na2SO4/K2SO4/CaO as1:0.5:0.1:0.1by roasting at850℃for0.5h. XRD analysis showed that the original aluminosilicate will decompose to NaSi3AlO8, KAlSi2O6and CaAl2Si2O8during sulfation roasting. The phases of CaF2and Ca4Si2O7F2are observed due to the addition of CaO.
Salt roasting with Na2SO4+CaCl2followed by water leaching was used to extract alkali metals from lepidolite. The experiments indicated that the optinum mass ratio of lepidolite/Na2SO4/CaCl2during roasting is1:0.5:0.3. The extraction efficiencies of Li, Rb and Cs are over90%after0.5h at880℃. The recovery efficiency of Li is essentially constant at830-930℃. The flexible roasting condition is easily maintained for industrial application. After the residue was leached in water, the aqueous solution was cooled to-5℃for2h for92.1%of the sulphate and3.9%of the chloride to be crystallized as Na2SO4·10H2O and NaCl, respectively. By evaporation and precipitation with Na2CO3, Li2CO3crystal with a purity of over99.5%was produced and a solution was obtained from which Rb and Cs could be recovered.
The lepidolite was pre-roasted athigh temperature with water steam atmosphere for defluorination. The defluorination percentage is61.43%at860℃when the duration of defluorination is40min. The XRD result indicated that as the defluorination time increases to40min, the diffraction peaks of lepidolite almost disappear and the main phases are aluminum silicate (LiAl(SiO3)2) and leucite (KAlSi2O6). The structure of the originalmineral is destroyed and transformed to aluminum silicate (LiAl(SiO3)2) and leucite (KAlSi2O6) during the defluorination process.
The defluorinated lepidolite washandled by Na2SO4and Ca(OH)2in the autoclaving leaching. The results indicated that the optimum operating parameters with the extraction efficiencies of Li, K, Rb and Cs as91.98%,93.06%,70.18%and61.02%were established as follows: defluorination temperature860℃, defluorination time30min (the defluorination percentage of lepidolite is42.3%), leaching temperature150℃, leaching time60min, defluorinated lepidolite/Na2SO4/Ca(OH)2/H2O1:1:0.7:7. The results of the leaching kinetics of defluorinated lepidolite show that the dissolution rates of Li, K, Rb, Cs accord with unreacted shrinking core models for solid film control. And the activation energies for the leaching of Li, K, Rb, Cs are 20.179kJ·mol-1,19.61J·mol-1,45.274kJ·mol-1and36.338kJ·mol-1respectively.
The defluorinated lepidolite was leached in a lime-milk autoclave. Various parameters including the defluorination percentage of lepidolite, milling time, temperature, time, lime-to-defluorinated lepidolite ratio, and liquid-to-solid ratio in the leaching process were optimized. The extraction efficiencies of Li, K, Rb and Cs can reach98.9%,68.67%,47.43%and43.54%under the optimal conditions.
Aim at the chloride and sulphate leaching solution, the phase equilibria of ternary reciprocal system Li+, Na+, K+//Cl--H2O and Li+, Na+, K+//SO42--H2O were studied by isothermal method. The results indicated that lithium can be concentrated to20-30g·L-1in Li+, Na+, K+//Cl--H2O system. But the concentration of lithium was limited to8.6g·L-1due to the precipitation of LiKSO4fromhigh concentration of sulphate solutions in Li+, Na+, K+//SO42--H2O system. In order to overcome this problem, the purified solution was cooled to-5℃for2h so that92.1%of the sulphate and3.9%of the chloride ions were crystallized as Na2SO4·10H2O and NaCl, respectively. The sulphate-free solution was thermally evaporated and the concentration of lithium was increased to20-24g-L-1when5mol·L-1Na2CO3solution was added to precipitate～86%of the lithium as lithium carbonate.
在热力学理论分析的基础上,对氯化焙烧-水浸法提取锂云母矿中锂、钾、铷、铯等有价金属的工艺进行了研究。结果表明,在采用复合氯化剂质量比CaCl2:NaCl为0.6：0.4,锂云母矿与氯化剂质量比为1.0,氯化焙烧温度为880℃,氯化焙烧时间为30min时,锂、钾、铷、铯的浸出率分别达到92.86%、88.49%、94.05%、93.06%。由于氯化反应过程中锂云母矿中的铝、硅可与CaCl2、NaCl中的钙和钠反应生成相应的不溶于水的铝硅酸盐,而有价碱金属锂、钾、铷、铯与氯离子结合形成相应的可溶于水的碱金属氯化物,氯化处理后所得熟料于常压下以水作浸出剂进行选择性浸出,可有效的将锂、钾、铷、铯浸出,而铝、硅等保留在浸出渣中。
对锂云母矿与固体氯化剂(CaCl2)的反应机理进行了研究。结果表明,在Ar气氛下,CaCl2和锂云母发生交互反应；在水蒸气气氛下部分氯化钙分解产生氯化氢,所以在H20蒸气气氛下锂云母与固体氯化剂发生交互反应的同时还存在与气体氯化氢反应；在O2气氛下部分CaCl2分解产生氯气,因此锂云母与CaCl2存在交互反应的同时还存在与氯气的反应。
采用硫酸盐焙烧-水浸法处理锂云母矿,可以选择性提取锂云母矿中的锂、钾、铷、铯,最佳工艺条件为：锂云矿:Na2SO4:K2SO4:CaO质量比1：0.5：0.1：0.1、焙烧温度900℃、焙烧时间30min,锂、钾、铷、铯的浸出率分别为91.61%、44.37%、29.33%和23.21%。XRD分析表明,硫酸盐焙烧可以将锂云母矿分解成NaSi3Al08、KAlSi2O6和CaAl2Si2O8。在硫酸盐焙烧过程中添加氧化钙可以有效促进锂云母矿中的氟与钙反应形成CaF2和Ca4Si2O7F2。
首次采用硫酸钠和氯化钙复合盐焙烧-水浸法处理锂云母矿。结果表明在830～930℃内锂的提取率基本不变,保持在90%以上,且焙烧物料未出现熔融现象。其最佳工艺条件为：锂云母:Na2SO4:CaCl2=1:0.5：0.3,焙烧温度850℃,焙烧时间30min,锂的浸出率达93.83%,铷、铯和钾的提取率分别为93.46%、90.08%和60.72%。
对锂云母在高温水蒸气下脱氟焙烧的条件及物理化学变化进行了系统的研究。结果表明,水蒸气气氛下于860℃反应40min,锂云母的脱氟率可达61.43%；SEM检测表明脱氟料表面变得粗糙蓬松;XRD检测表明脱氟料锂云母衍射峰几乎消失,主要物相转变成铝硅酸锂(LiAl(SiO3)2)和铝硅酸钾(KAlSi2O6)。
对硫酸钠、石灰压煮法处理脱氟锂云母工艺进行了系统的研究。结果表明,当脱氟锂云母粒度在300目以下,搅拌速度400r·min-1,反应时间3h,反应温度150℃,脱氟锂云母：硫酸钠：石灰：水=1：1：0.7：7,锂、钾、铷、铯的浸出率分别为91.98%、93.06%、70.18%、61.02%。
压煮反应过程动力学实验研究结果表明锂、钾、铷、铯浸出过程动力学符合未反应收缩核模型,属于固膜扩散控制,锂、钾、铷、铯浸出活化能分别为20.179kJ·mol-1、19.617kJ·mol-1、45274kJ·mol-1和20.179kJ·mol-1、19.617kJ·mol-1、45.274kJ·mol-1
对石灰乳压煮法处理脱氟锂云母工艺进行了系统的研究。结果表明,当脱氟锂云母球磨活化100min,石灰与脱氟锂云母质量比为1,液固比为4,压煮反应温度150℃,压煮反应时间1h,搅拌速度为400r-min-1,锂、钾、铷、铯的提取率分别为98.9%、68.67%、47.43%和43.54%。
针对氯化物、硫酸盐浸出液体系,对Li+, Na+, K+//Cl-及Li+, Na+, K+//SO42-体系相图及浸出液中制备碳酸锂方法进行了研究。结果表明氯化物体系可以通过浓缩结晶的方式将浸出液中的锂浓度提高到20～30g·L-1,而硫酸盐体系在浓缩到一定程度后会形成LiKSO4复盐而沉淀；因此Li+, Na+, K+//SO42-体系可通过冷冻到-5℃冷冻结晶,降低体系中的硫酸根浓度。去除硫酸根的溶液再次浓缩,将Li浓度提高到20-24g·L-1后与5mol/L的Na2CO3反应获得Li2CO3。
The lepidolite from Jiangxi Province (China)has a reserve of1.1million tons Li2O and accounts for30%of China's domestic proven reserves, but little of ithas been utilized. The present methods of lithium extraction from lepidolite worldwide were summarized in this paper. Comprehensive extraction and utilization of lithium, potassium, rubidium and cesium from the lepidolite of Jiangxi Province were introduced. Chlorination roasting, sulfation roasting and autoclaving method were adopted. The effect and mechanism ofmineral phase reconstruction on the extraction behavior of Li, K, Rb, Cs from lepidolite with different methods were discussed in detail and conclusions were made as follows.
Based on the analysis of thermodynamics, the experimental conditions of chlorination roasting-water leaching were studied. The conditional experiments indicated that the optimum mass ratio of lepidolite/CaCl2/NaCl is1:0.6:0.4during the roasting process. The extraction efficiencies of Li, K, Rb and Cs are92.86%,88.49%,94.05%and93.06%, respectively after roasting for0.5h at880℃. Li, K, Rb and Cs are selectively leached by chlorination roasting and water leaching at atmospheric pressure, and impurities extraction can be suppressed.
The reaction mechanism of lepidolite and solid chloridizing agent was investigated in details. The results indicated that base-exchange reaction occurs between CaCl2and lepidolite under argon atmosphere. CaCl2is decomposed intohCl under vapor atmosphere andhCl will react with lepidolite under vapor atmosphere. Because CaCl2is decomposed into Cl2under oxygen atmosphere, lepidolite will react with CaCl2and Cl2.
Sulfation roasting followed by water leaching process was used to selectively extract Li, K, Rb and Cs from lepidolite. Various operational parameters including roasting temperature, the amount of additives, and solid/liquid ratio in the leaching process were optimized. The extraction efficiencies of Li, K, Rb and Cs could reach91.61%,44.37%,29.33%and23.21%with a mass ratio of lepidolite/Na2SO4/K2SO4/CaO as1:0.5:0.1:0.1by roasting at850℃for0.5h. XRD analysis showed that the original aluminosilicate will decompose to NaSi3AlO8, KAlSi2O6and CaAl2Si2O8during sulfation roasting. The phases of CaF2and Ca4Si2O7F2are observed due to the addition of CaO.
Salt roasting with Na2SO4+CaCl2followed by water leaching was used to extract alkali metals from lepidolite. The experiments indicated that the optinum mass ratio of lepidolite/Na2SO4/CaCl2during roasting is1:0.5:0.3. The extraction efficiencies of Li, Rb and Cs are over90%after0.5h at880℃. The recovery efficiency of Li is essentially constant at830-930℃. The flexible roasting condition is easily maintained for industrial application. After the residue was leached in water, the aqueous solution was cooled to-5℃for2h for92.1%of the sulphate and3.9%of the chloride to be crystallized as Na2SO4·10H2O and NaCl, respectively. By evaporation and precipitation with Na2CO3, Li2CO3crystal with a purity of over99.5%was produced and a solution was obtained from which Rb and Cs could be recovered.
The lepidolite was pre-roasted athigh temperature with water steam atmosphere for defluorination. The defluorination percentage is61.43%at860℃when the duration of defluorination is40min. The XRD result indicated that as the defluorination time increases to40min, the diffraction peaks of lepidolite almost disappear and the main phases are aluminum silicate (LiAl(SiO3)2) and leucite (KAlSi2O6). The structure of the originalmineral is destroyed and transformed to aluminum silicate (LiAl(SiO3)2) and leucite (KAlSi2O6) during the defluorination process.
The defluorinated lepidolite washandled by Na2SO4and Ca(OH)2in the autoclaving leaching. The results indicated that the optimum operating parameters with the extraction efficiencies of Li, K, Rb and Cs as91.98%,93.06%,70.18%and61.02%were established as follows: defluorination temperature860℃, defluorination time30min (the defluorination percentage of lepidolite is42.3%), leaching temperature150℃, leaching time60min, defluorinated lepidolite/Na2SO4/Ca(OH)2/H2O1:1:0.7:7. The results of the leaching kinetics of defluorinated lepidolite show that the dissolution rates of Li, K, Rb, Cs accord with unreacted shrinking core models for solid film control. And the activation energies for the leaching of Li, K, Rb, Cs are 20.179kJ·mol-1,19.61J·mol-1,45.274kJ·mol-1and36.338kJ·mol-1respectively.
The defluorinated lepidolite was leached in a lime-milk autoclave. Various parameters including the defluorination percentage of lepidolite, milling time, temperature, time, lime-to-defluorinated lepidolite ratio, and liquid-to-solid ratio in the leaching process were optimized. The extraction efficiencies of Li, K, Rb and Cs can reach98.9%,68.67%,47.43%and43.54%under the optimal conditions.
Aim at the chloride and sulphate leaching solution, the phase equilibria of ternary reciprocal system Li+, Na+, K+//Cl--H2O and Li+, Na+, K+//SO42--H2O were studied by isothermal method. The results indicated that lithium can be concentrated to20-30g·L-1in Li+, Na+, K+//Cl--H2O system. But the concentration of lithium was limited to8.6g·L-1due to the precipitation of LiKSO4fromhigh concentration of sulphate solutions in Li+, Na+, K+//SO42--H2O system. In order to overcome this problem, the purified solution was cooled to-5℃for2h so that92.1%of the sulphate and3.9%of the chloride ions were crystallized as Na2SO4·10H2O and NaCl, respectively. The sulphate-free solution was thermally evaporated and the concentration of lithium was increased to20-24g-L-1when5mol·L-1Na2CO3solution was added to precipitate～86%of the lithium as lithium carbonate.
引文
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