高结合率混合铜矿选冶新工艺试验研究
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摘要
楚雄混合铜矿含铜1.42%,氧化率69.10%,结合率26.09%,氧化钙含量10.73%,氧化镁含量2.99%,是一种高氧化率、高结合率、高钙镁含量的难处理混合铜矿,矿石中铜矿物嵌布粒度细、组成复杂,选矿和冶金回收都是一个世界性难题。自上世纪80年代以来,国内外许多大专院校及科研院所对其进行了深入的研究,但都未能有所突破。目前,采用人工拣选预分离该矿石,将氧化率相对低的部分进行浮选,氧化率相对高的部分进行酸浸,基本可以获得60%-70%的回收率。由于矿石性质复杂、氧化率变化大、钙镁含量高、人工拣选劳动强度大等特点,使得工业生产上处理能力小、酸浸耗酸高,难以满足规模化的大工业生产,致使该铜矿资源至今没有得到有效利用。
对矿石进行工艺矿物学性质研究,发现矿石中大部分铜矿物与脉石矿物致密共生,部分粒度极细,呈浸染状分布在钙硅泥质中,形成结合铜,导致这部分铜矿物的浮游活性降低或难以浮选。针对该矿石的特点,通过大量的试验研究与理论分析,提出了“硫化浮选—强化捕收—反浮选钙镁—中矿酸浸”的全新选冶联合工艺流程,优化组合,优势互补。采用硫化浮选法回收矿石中的硫化铜和易浮氧化铜矿物；使用捕收能力强的药剂或药剂组合对尾矿中的难浮铜矿物进行强化捕收得中矿；针对钙镁脉石含量高且铜结合率高的浮选二次尾矿,采用反浮选钙镁工艺得到低钙镁含量的铜中矿；混合中矿进入酸浸工艺流程中回收其中的铜矿物。该全新工艺流程顺应矿石性质,适应性强,为规模化开发利用该混合铜矿提供依据。
针对矿石中的硫化铜和易浮氧化铜矿物,进行硫化浮选工艺试验研究,考查磨矿细度以及药剂浓度等因素对铜矿浮选的影响,确定最佳硫化浮选工艺流程,即通过三次粗选,粗精矿集中再磨,再磨粗精矿三次精选,精选中矿顺序返回的闭路浮选流程,最终得到品位为21.53%,回收率为51.13%的铜精矿。通过对闭路各产品的铜物相分析及钙镁含量的测定,发现铜精矿中硫化铜矿的回收率只有86.55%,游离氧化铜矿的回收率为53.41%,结合氧化铜矿的回收率仅为6.03%。
直接对硫化浮选的尾矿进行反浮选钙镁,可能会使得部分铜矿物随钙镁脉石矿物同时上浮而损失于钙镁尾矿中,因此在进行反浮选钙镁之前,采用捕收能力强的药剂和药剂组合对其中的难浮铜矿物进行强化捕收,是十分必要的。试验研究结果表明,采用异戊基黄药与羟肟酸混合药剂对难浮铜矿物进行强化捕收,得到铜品位1.02%,回收率6.57%的中矿。对该中矿进行钙镁含量分析,其中CaO与MgO回收率均在8%左右,效果较为理想。
对于混合药剂强化捕收的二次尾矿,其中含有大量难以在浮选中回收的结合铜,为了进一步提高铜的综合回收率,对二次尾矿中的钙镁脉石矿物进行反浮选,降低其中的耗酸物质,为酸浸回收结合铜创造条件。油酸对铜矿物有着较好的捕收性能,国内外从未见发表过使用油酸对铜矿进行反浮选的相关报道。针对楚雄混合铜矿的矿石性质特点,选择油酸作为反浮选钙镁的捕收剂,取得了较好的效果.反浮选槽内产物为中矿,泡沫产品为尾矿,中矿的铜品位为0.71%,明显高于尾矿0.53%的铜品位,仅有12.28%的铜损失在尾矿中。含钙矿物得到较好的分选,反浮选阶段油酸对氧化钙的脱除率高达81.82%,另有33.99%的氧化镁脱除在尾矿中。
通过大量试验研究,确定了最终的浮选工艺全流程,即集中磨矿,三次粗选,粗精矿集中再磨,三次精选获得铜精矿,尾矿强化捕收,二次尾矿钙镁反浮选,强化捕收中矿与反浮选槽中产物合并为铜中矿,反浮选泡沫产品为尾矿。浮选闭路试验所得铜精矿的品位为21.69%,回收率为52.52%。中矿铜品位为0.85%,铜回收率为27.18%；CaO品位为5.39%, CaO回收率为22.71%；MgO品位为2.89%, MgO回收率为44.09%。尾矿铜品位为0.56%,铜损失率为20.30%。
通过浸出热力学计算与动力学分析,研究了矿石中铜矿物及含钙镁矿物硫酸浸出过程的机理,结合中矿的性质,提出“浓浆闪浸”工艺。试验考查了硫酸用量、浸出时间、液固比和搅拌强度等参数对铜浸出率的影响。试验结果表明,常温下搅拌强度为30m/min,液固比2：1,浸出时间30min,硫酸用量150Kg/t矿中矿的铜浸出率达到84.70%,每吨电解铜消耗17.31t硫酸,“浓浆闪浸”技术取得了良好的效果。
针对楚雄高结合率难处理混合铜矿,采用全新的“硫化浮选—强化捕收—反浮选钙镁—中矿酸浸”选冶联合工艺流程进行处理,获得了铜精矿品位21.69%,铜综合回收率75.54%,电解铜吨铜耗酸17.31t的技术经济指标,为规模化开发利用该铜矿资源奠定了重要的理论与试验基础。
The mixed copper ore in Chuxiong of Yunnan province, China, contains 1.42% copper, with an oxidation rate of 69.10% and a combination rate of 26.09%, the contents of calcium oxide and magnesium oxide in the ore are 10.73% and 2.99% respectively. The ore is very difficult to be treated either by beneficiation or metallurgy due to its high oxidation rate, high combination rate, and high content of acid-consuming gangue minerals. Moreover, the copper minerals are embedded in fine grain size with complex composition in the ore. Many researchers in domestic and foreign universities and research institutes have conducted a lot of research on it since 1980's, no breakthrough, however, was made. At present, the copper ore is mainly separated and processed by hand sorting, followed by flotation for the ore with low oxidation rate, and acid leaching for the ore with high oxidation rate, which results in a copper recovery of 60% to 70%. The ore is so difficult to treat resulting from the above complex natures, small processing capacity and high cost for acid leaching that insufficient utilization of the ore was made until now.
Process mineralogy research of the ore, indicated that most of the copper ore embedded in the gangue mineral, and some copper mineral particle size are tiny in a disseminated distribution in the calcium and silicon mud, forming so-called combined copper oxide mineral, which leads to poor floatability and performance of the copper ore.
A novel combined process with beneficiation and metallurgy was proposed as "Sulfidization flotation-Intensifying collection-Reverse flotation of acid-consuming gangue minerals-Acid leaching for middlings" based on the characteristics of the ore and lots of tests and theoretical analysis, i.e. recovering copper sulfide minerals and part of easy-to-float oxidized copper minerals by sulfidization flotation; using strong collective reagents or combined reagents to intensify the collecting of difficult-to-float copper minerals in the tailings after sulfidization flotation; transferring the secondary tailings with high content of acid-consuming gangue minerals and high combination rate into reverse flotation to remove the calcium oxide and magnesium oxide; mixed middlings into the acid leaching process to recover the copper minerals. The new process conforms to the nature of the ore, is adaptable for large-scale industrial production.
The flotation tests were conducted for the flotation of copper sulfide minerals and part of easy-to-float oxidized copper minerals, and the best flotation flowsheet and conditions are determined as three-stage roughing-rough concentrate regrinding-three-stage cleaning (closed-circuit of middlings returned by turns) through the investigations of grinding fineness, various reagents dosages, stages of rougher and cleaner, as well as regrinding time of rough concentrate. The final copper concentrate is obtained with a grade of 21.53% and a recovery of 51.13%, in which the phase recoveries of copper sulfide ore, free copper oxide ore, and combined copper oxide ore are 86.55%,53.41% and 6.03% respectively.
The main purpose of reverse flotation is to remove the acid-consuming gangue minerals. It is, however, very necessary to recover the difficult-to-float copper minerals in the tailings with strong collector before carrying out reverse flotation because part of the copper minerals may lose in the tailings in process of reverse flotation. The reverse flotation was performed using blending collectors with isoamyl xanthate and hydroxamic acid, the middlings with copper grade of 1.02% and recovery of 6.55% was obtained, in which CaO and MgO recovery is about 8%, respectively.
There is large quantity of combined copper minerals in the secondary tailings by intensifying collection, the tailings produced in the aforementioned process was treated by reverse flotation to remove the acid-consuming gangue minerals in order to supply good conditions for acid leaching. Oleic acid has proved to be a better flotation collector to recover this copper ore, and there is, so far, no report of utilization of oleic acid in the reverse flotation of copper ore. Oleic acid was employed as collector in reverse flotation to remove the acid-consuming gangue minerals based on the characteristics of the ore, and good results was achieved. The product in the flotation cell is middlings with a copper grade of 0.71%, which is higher than that of tailings grade of 0.53%, and the foam product forms the tailings with the loss of copper recovery of 12.28%. Moreover, good separation of calcium minerals was obtained in the reverse flotation stage with a removal rate of CaO 81.82%, as well as 33.99 percent of the magnesium oxide was removed in the stage.
The final flotation flowsheet and conditions were obtained after series of tests, the grinding fineness is 85%-0.074mm, and the flowsheet is three-stage rougher, roughing concentrate regrinding and three-stage cleaner, the final concentrate with copper grade of 21.69% and recovery of 52.52%. The tailings, subsequently, was put into the process of intensifying flotation, and a secondary concentrate was produced in the process. The tertiary concentrate was also produced after the secondary tailings putting into reverse flotation to remove the calcium oxide and magnesium oxide, the blending of secondary concentrate and tertiary concentrate forms the middlings product with a copper grade of 0.85%, and a recovery of 27.18%, the content of calcium oxide in the middlings is decreased by 5.39%. The copper grade in final tailings, is 0.56% and with a lost copper recovery of 20.30%.
Such a novel hydrometallurgy process as flash leaching of thick slurry was proposed based on thermodynamic calculation and kinetic analysis of copper leaching in the middlings. The leaching conditions such as dosage of sulfuric acid, leaching time, ratio of liquid to solid and agitation intensity were investigated. A good leaching performance was achieved with a leaching rate of 84.70% under such appropriate conditions as agitation intensity 30m/min, ratio of liquid to solid 2:1, leaching time 30 min, dosage of sulfuric acid 150 Kg per ton of ore, hence one ton electrolytic copper need 17.31 ton sulfuric acid.
In a word, a novel combined process with "Sulfidization flotation-Intensifying collection-Reverse flotation of acid-consuming gangue minerals-Acid leaching for middlings" was proposed for Chuxiong copper ore. Satisfied performance was achieved as follows, the copper in final concentrate is 21.69% and the total copper recovery of the whole process is 75.35%. The progress made in this dissertation will supply a good base for high-efficient utilization of the mixed copper ore.
对矿石进行工艺矿物学性质研究,发现矿石中大部分铜矿物与脉石矿物致密共生,部分粒度极细,呈浸染状分布在钙硅泥质中,形成结合铜,导致这部分铜矿物的浮游活性降低或难以浮选。针对该矿石的特点,通过大量的试验研究与理论分析,提出了“硫化浮选—强化捕收—反浮选钙镁—中矿酸浸”的全新选冶联合工艺流程,优化组合,优势互补。采用硫化浮选法回收矿石中的硫化铜和易浮氧化铜矿物；使用捕收能力强的药剂或药剂组合对尾矿中的难浮铜矿物进行强化捕收得中矿；针对钙镁脉石含量高且铜结合率高的浮选二次尾矿,采用反浮选钙镁工艺得到低钙镁含量的铜中矿；混合中矿进入酸浸工艺流程中回收其中的铜矿物。该全新工艺流程顺应矿石性质,适应性强,为规模化开发利用该混合铜矿提供依据。
针对矿石中的硫化铜和易浮氧化铜矿物,进行硫化浮选工艺试验研究,考查磨矿细度以及药剂浓度等因素对铜矿浮选的影响,确定最佳硫化浮选工艺流程,即通过三次粗选,粗精矿集中再磨,再磨粗精矿三次精选,精选中矿顺序返回的闭路浮选流程,最终得到品位为21.53%,回收率为51.13%的铜精矿。通过对闭路各产品的铜物相分析及钙镁含量的测定,发现铜精矿中硫化铜矿的回收率只有86.55%,游离氧化铜矿的回收率为53.41%,结合氧化铜矿的回收率仅为6.03%。
直接对硫化浮选的尾矿进行反浮选钙镁,可能会使得部分铜矿物随钙镁脉石矿物同时上浮而损失于钙镁尾矿中,因此在进行反浮选钙镁之前,采用捕收能力强的药剂和药剂组合对其中的难浮铜矿物进行强化捕收,是十分必要的。试验研究结果表明,采用异戊基黄药与羟肟酸混合药剂对难浮铜矿物进行强化捕收,得到铜品位1.02%,回收率6.57%的中矿。对该中矿进行钙镁含量分析,其中CaO与MgO回收率均在8%左右,效果较为理想。
对于混合药剂强化捕收的二次尾矿,其中含有大量难以在浮选中回收的结合铜,为了进一步提高铜的综合回收率,对二次尾矿中的钙镁脉石矿物进行反浮选,降低其中的耗酸物质,为酸浸回收结合铜创造条件。油酸对铜矿物有着较好的捕收性能,国内外从未见发表过使用油酸对铜矿进行反浮选的相关报道。针对楚雄混合铜矿的矿石性质特点,选择油酸作为反浮选钙镁的捕收剂,取得了较好的效果.反浮选槽内产物为中矿,泡沫产品为尾矿,中矿的铜品位为0.71%,明显高于尾矿0.53%的铜品位,仅有12.28%的铜损失在尾矿中。含钙矿物得到较好的分选,反浮选阶段油酸对氧化钙的脱除率高达81.82%,另有33.99%的氧化镁脱除在尾矿中。
通过大量试验研究,确定了最终的浮选工艺全流程,即集中磨矿,三次粗选,粗精矿集中再磨,三次精选获得铜精矿,尾矿强化捕收,二次尾矿钙镁反浮选,强化捕收中矿与反浮选槽中产物合并为铜中矿,反浮选泡沫产品为尾矿。浮选闭路试验所得铜精矿的品位为21.69%,回收率为52.52%。中矿铜品位为0.85%,铜回收率为27.18%；CaO品位为5.39%, CaO回收率为22.71%；MgO品位为2.89%, MgO回收率为44.09%。尾矿铜品位为0.56%,铜损失率为20.30%。
通过浸出热力学计算与动力学分析,研究了矿石中铜矿物及含钙镁矿物硫酸浸出过程的机理,结合中矿的性质,提出“浓浆闪浸”工艺。试验考查了硫酸用量、浸出时间、液固比和搅拌强度等参数对铜浸出率的影响。试验结果表明,常温下搅拌强度为30m/min,液固比2：1,浸出时间30min,硫酸用量150Kg/t矿中矿的铜浸出率达到84.70%,每吨电解铜消耗17.31t硫酸,“浓浆闪浸”技术取得了良好的效果。
针对楚雄高结合率难处理混合铜矿,采用全新的“硫化浮选—强化捕收—反浮选钙镁—中矿酸浸”选冶联合工艺流程进行处理,获得了铜精矿品位21.69%,铜综合回收率75.54%,电解铜吨铜耗酸17.31t的技术经济指标,为规模化开发利用该铜矿资源奠定了重要的理论与试验基础。
The mixed copper ore in Chuxiong of Yunnan province, China, contains 1.42% copper, with an oxidation rate of 69.10% and a combination rate of 26.09%, the contents of calcium oxide and magnesium oxide in the ore are 10.73% and 2.99% respectively. The ore is very difficult to be treated either by beneficiation or metallurgy due to its high oxidation rate, high combination rate, and high content of acid-consuming gangue minerals. Moreover, the copper minerals are embedded in fine grain size with complex composition in the ore. Many researchers in domestic and foreign universities and research institutes have conducted a lot of research on it since 1980's, no breakthrough, however, was made. At present, the copper ore is mainly separated and processed by hand sorting, followed by flotation for the ore with low oxidation rate, and acid leaching for the ore with high oxidation rate, which results in a copper recovery of 60% to 70%. The ore is so difficult to treat resulting from the above complex natures, small processing capacity and high cost for acid leaching that insufficient utilization of the ore was made until now.
Process mineralogy research of the ore, indicated that most of the copper ore embedded in the gangue mineral, and some copper mineral particle size are tiny in a disseminated distribution in the calcium and silicon mud, forming so-called combined copper oxide mineral, which leads to poor floatability and performance of the copper ore.
A novel combined process with beneficiation and metallurgy was proposed as "Sulfidization flotation-Intensifying collection-Reverse flotation of acid-consuming gangue minerals-Acid leaching for middlings" based on the characteristics of the ore and lots of tests and theoretical analysis, i.e. recovering copper sulfide minerals and part of easy-to-float oxidized copper minerals by sulfidization flotation; using strong collective reagents or combined reagents to intensify the collecting of difficult-to-float copper minerals in the tailings after sulfidization flotation; transferring the secondary tailings with high content of acid-consuming gangue minerals and high combination rate into reverse flotation to remove the calcium oxide and magnesium oxide; mixed middlings into the acid leaching process to recover the copper minerals. The new process conforms to the nature of the ore, is adaptable for large-scale industrial production.
The flotation tests were conducted for the flotation of copper sulfide minerals and part of easy-to-float oxidized copper minerals, and the best flotation flowsheet and conditions are determined as three-stage roughing-rough concentrate regrinding-three-stage cleaning (closed-circuit of middlings returned by turns) through the investigations of grinding fineness, various reagents dosages, stages of rougher and cleaner, as well as regrinding time of rough concentrate. The final copper concentrate is obtained with a grade of 21.53% and a recovery of 51.13%, in which the phase recoveries of copper sulfide ore, free copper oxide ore, and combined copper oxide ore are 86.55%,53.41% and 6.03% respectively.
The main purpose of reverse flotation is to remove the acid-consuming gangue minerals. It is, however, very necessary to recover the difficult-to-float copper minerals in the tailings with strong collector before carrying out reverse flotation because part of the copper minerals may lose in the tailings in process of reverse flotation. The reverse flotation was performed using blending collectors with isoamyl xanthate and hydroxamic acid, the middlings with copper grade of 1.02% and recovery of 6.55% was obtained, in which CaO and MgO recovery is about 8%, respectively.
There is large quantity of combined copper minerals in the secondary tailings by intensifying collection, the tailings produced in the aforementioned process was treated by reverse flotation to remove the acid-consuming gangue minerals in order to supply good conditions for acid leaching. Oleic acid has proved to be a better flotation collector to recover this copper ore, and there is, so far, no report of utilization of oleic acid in the reverse flotation of copper ore. Oleic acid was employed as collector in reverse flotation to remove the acid-consuming gangue minerals based on the characteristics of the ore, and good results was achieved. The product in the flotation cell is middlings with a copper grade of 0.71%, which is higher than that of tailings grade of 0.53%, and the foam product forms the tailings with the loss of copper recovery of 12.28%. Moreover, good separation of calcium minerals was obtained in the reverse flotation stage with a removal rate of CaO 81.82%, as well as 33.99 percent of the magnesium oxide was removed in the stage.
The final flotation flowsheet and conditions were obtained after series of tests, the grinding fineness is 85%-0.074mm, and the flowsheet is three-stage rougher, roughing concentrate regrinding and three-stage cleaner, the final concentrate with copper grade of 21.69% and recovery of 52.52%. The tailings, subsequently, was put into the process of intensifying flotation, and a secondary concentrate was produced in the process. The tertiary concentrate was also produced after the secondary tailings putting into reverse flotation to remove the calcium oxide and magnesium oxide, the blending of secondary concentrate and tertiary concentrate forms the middlings product with a copper grade of 0.85%, and a recovery of 27.18%, the content of calcium oxide in the middlings is decreased by 5.39%. The copper grade in final tailings, is 0.56% and with a lost copper recovery of 20.30%.
Such a novel hydrometallurgy process as flash leaching of thick slurry was proposed based on thermodynamic calculation and kinetic analysis of copper leaching in the middlings. The leaching conditions such as dosage of sulfuric acid, leaching time, ratio of liquid to solid and agitation intensity were investigated. A good leaching performance was achieved with a leaching rate of 84.70% under such appropriate conditions as agitation intensity 30m/min, ratio of liquid to solid 2:1, leaching time 30 min, dosage of sulfuric acid 150 Kg per ton of ore, hence one ton electrolytic copper need 17.31 ton sulfuric acid.
In a word, a novel combined process with "Sulfidization flotation-Intensifying collection-Reverse flotation of acid-consuming gangue minerals-Acid leaching for middlings" was proposed for Chuxiong copper ore. Satisfied performance was achieved as follows, the copper in final concentrate is 21.69% and the total copper recovery of the whole process is 75.35%. The progress made in this dissertation will supply a good base for high-efficient utilization of the mixed copper ore.
引文
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