Non-ferrous metal mines beneficiation research focused on low-grade complex multi-metal mines and comprehensive recovery of refractory oxide ores. In addition to the conventional beneficiation methods, the focus of the research is also reflected in the use of new and combined agents and the properness of the joint process.
1. Lead-zinc ore dressing
For separation of a Cu Pb Copper Lead Zinc sulfide ores Yunnan exist in the production index is not over, the copper-containing lead concentrate of high intermodulation problems, Yang Wu Jia copper concentrate were mixed Pb, Cu, Pb flotation separation. The results show that, when mixed concentrate regrinding to 80% -0.074 mm, sodium sulfite, sodium silicate and a combination of CMC inhibitor instead of suppressing weight chromium potassium galena to Z-200 acetate instead of xanthate as chalcopyrite The collector can obtain good sorting technical indicators, the copper concentrate grade is 23.30%, the lead content is 3.30%, the aluminum concentrate grade is 64.66%, and the copper content is 0.50%, which effectively achieves the separation of copper and lead.
In view of the coarse grain size and low copper content of a copper-lead-zinc polymetallic sulphide ore in Qinghai, Liu Shouxin et al. used a copper-lead mixed-mixed copper-lead separation-tailing zinc-selecting process to conduct a sorting test. the study. When copper and lead are mixed, Ty-1 and zinc sulfate are used as combination inhibitors of sphalerite and iron sphalerite. Ethyl xanthate and J-21 are used as collectors to achieve effective copper, lead minerals and zinc minerals. Separation, and the viscosity of the foam is moderate, which creates favorable conditions for the separation of copper and lead in the next step. The separation of copper and lead is activated carbon, CMC and potassium dichromate, and the copper is effectively separated. The ideal beneficiation technology is obtained. index. Ren Xiangjun et al. determined the use of copper-lead preferential flotation, water glass + sodium sulfite + carboxymethyl cellulose combination inhibitor for copper-lead separation and copper-lead mixed floatation through another copper-lead-zinc polymetallic sulphide ore. The experimental scheme of flotation of zinc minerals after activation of copper sulfate by tailings was systematically studied, and the effective separation of copper and lead was successfully achieved. Finally, copper concentrate with copper grade of 21.40% and copper recovery of 67.65% was obtained. The grade is 52.92%, the lead recovery rate is 95.90%, the lead concentrate and zinc grade are 50.21%, and the zinc recovery rate is 83.74% zinc concentrate.
For a copper-lead-zinc polymetallic sulphide ore in Liaoning, Liu Yalong et al. used a copper-lead mixed-copper-lead separation-mixed floating tailings to suppress the sulphur-spraying zinc flotation process. Mixed flotation with ethyl sulphide nitrogen + aniline black drug as collector, ZnS0 4 + Na 2 S0 3 as inhibitor, and control of pulp pH = about 11.5, to achieve separation of copper-lead minerals and zinc-sulfur minerals; application of water glass The combination of sodium sulfite and carboxymethyl cellulose, replacing cyanide and potassium dichromate, successfully achieved the separation of copper and lead; through the closed circuit test, copper concentrate with a copper grade of 28.54% copper recovery of 65.62% was obtained. The lead grade is 55.69%, the lead recovery rate is 83.21%, the lead concentrate and zinc grade are 51.09%, and the zinc recovery rate is 90.87% zinc concentrate.
According to the nature of a complex refractory copper-lead-zinc polymetallic ore in Tibet, Li Guanqi conducted a separation test using the principle flow of copper-lead mixed flotation-copper-lead separation-copper-lead floating tail zinc-selecting, and copper-lead mixed flotation was adopted. Combination of Bp, butylammonium black and xanthate, using sodium sulfide, zinc sulfate and sodium carbonate as inhibitors of zinc minerals; when separating copper and lead, using activated carbon for drug removal, using CMC, Na 2 S0 3 As an inhibitor of lead minerals, the Na 2 Si0 3 environmentally friendly combination agent successfully achieved the separation of copper and lead, and the sorting technical index obtained was significantly improved compared with the current production.
Zheng Yajie et al. conducted a flotation separation test on a high- arsenic, low-copper, lead, zinc- silver ore in Inner Mongolia. The 261 0 test uses a copper-lead-zinc and other floatable-first-order flotation process, and uses FN as an inhibitor of arsenic minerals. The high concentration of arsenic in copper concentrates not only does not use the toxic inhibitor potassium dichromate, but also significantly improves the copper concentrate grade. The copper concentrate grade obtained by the test was 28.6%, the copper recovery rate was 66.41%, and the recovery rates of lead, zinc and silver were also greater than 90%.
Luo Jin conducted a sulfide flotation test on lead oxide ore in a complex high-oxidation lead-zinc ore. It was found that when Na 2 S was used as the vulcanizing agent for lead oxide, not only the proper amount of Na 2 S was needed, but also the initial concentration of Na 2 S was required. The grade and recovery rate of lead concentrate reached 46.02% by vulcanization flotation method. And 81.16%, effective recovery of lead oxide minerals.
In view of the fact that lead-zinc minerals in a refractory lead-zinc ore in Turpan are mutually metasomatized and wrapped, with fine grain size and difficult dissociation characteristics, Wang Fengshui has adopted a preferential process of selecting lead and zinc coarse concentrate for re-grinding and re-election. Experimental Research. Under the conditions of lead rough grinding grinding fineness -0.074 mm accounting for 80% and zinc rough grinding grinding fineness of 0.043 mm accounting for 90%, the lead grade is 40.22%, the Zn content is 6.94%, and the lead recovery rate is 82.48%. Lead concentrate and zinc grade 50.17%, zinc concentrate containing Pb 1.08% and zinc recovery rate of 86.92%.
Feng Zhongwei discovered through flotation experiments on a lead-zinc mine in Yunnan that soluble salts such as zinc sulfate and ferrous sulfate in the ore have a strong inhibitory effect on lead-zinc minerals, in order to reduce the soluble salt to lead-zinc mineral flotation. The effect is to use the alkali-free process of flotation of lead-zinc minerals under the natural pH condition of the pulp. When lead is selected, zinc sulfate and sodium sulfite are used as inhibitors of zinc minerals, and aniline black and butylammonium black are used as collectors. When zinc is used, water glass, sodium sulfite and carboxymethyl cellulose are used as adjusting agents, and PN-405 is used as collector. Finally, the lead grade is 59.57%, the lead recovery rate is 75.14%, and the lead concentrate and zinc grade are 53.93%. Recovery rate 93. 70% zinc concentrate.
In view of the high degree of oxidation, easy muddy, and difficult recovery of zinc oxide minerals in a lead-zinc ore in Shaanxi Province, Wang Hongmei and others used lead sulfide minerals and oxide minerals mixed flotation, zinc sulfide minerals and oxide minerals in turn. A systematically studied study was conducted on a separate recovery scheme. When the lead is selected, the combined collector, ethylsulfide nitrogen + butyl black liquor is used, and when the zinc oxide mineral is selected, the composite collector A-928 is collected. Finally, lead concentrates with a lead grade and recovery rate of 53.67% and 82.92%, respectively, containing 5.23% zinc, zinc grade and recovery rates of 51.08% and 40.75%, zinc sulfide concentrate containing 1.06% lead, zinc grade were obtained. The recovery rate is 22.55%, 44.28%, and 1.22% lead oxide concentrate, which realizes the sorting of lead-zinc oxide ore.
Second, copper molybdenum and nickel ore dressing
The characteristics of a low-grade copper oxide ore, and the oxidation rate higher than the rate of binding, and the like with the addition of Jianwen vulcanizing agent and the prerequisite for activation of ammonium sulfate, sodium hexametaphosphate and sodium silicate in order to constitute a combination of phosphorus inhibitors to A mixed collector was prepared by mixing xanthate 680, butylammonium black and hydroxamic acid, and a flotation test was conducted. The optimal flotation conditions and chemical system of copper oxide minerals were determined. Through the closed-circuit test, the sorting technical indexes of copper concentrate grade 17.39% and copper recovery rate 59.36% were obtained.
Ma Jiezhen et al. carried out ore properties analysis and ore dressing test on the pyrite-type copper-zinc polymetallic ore in Ashele copper mine in Xinjiang, combined with on-site production practice, using asynchronous separation of cyclone-static microbubble flotation column to enhance recovery. The technological transformation of the process has significantly improved the mineral processing technical indicators. The copper recovery rate has increased from 77.59% at the initial stage of production to 86.43%, and the zinc recovery rate has increased from 20.48% to 48.94%.
For the problem that a mixed copper- cobalt ore has a high oxidation rate, contains a large amount of carbonaceous slime, and cannot achieve the desired recovery efficiency by conventional folding method, Ou Mingming and the like remove the slime by pre-flotation to eliminate the flotation of carbonaceous slime. The effect of the process, then asynchronous flotation of copper sulphide cobalt minerals and copper oxide cobalt minerals, and the use of vulcanizing agent sodium hydrosulfide to enhance the flotation effect of copper oxide cobalt minerals. The results show that after the use of these measures, the obtained copper-cobalt concentrate has a copper grade of 21.12%, a copper yield of 88.55%, a cobalt content of 0.116%, and a diamond recovery rate of 31.39%.
Wei Dangsheng conducted a flotation test on a copper- molybdenum ore in Guangdong, and determined the process of mixing flotation-sulfur-dissolving copper-molybdenum-copper-molybdenum separation. The mixing was carried out under the conditions of grinding fineness of 75.00%-0.074mm. Flotation, after the mixed flotation coarse concentrate is further ground to 86.00%-0.043mm, the pyrite is inhibited by lime, the copper-molybdenum flotation is obtained to obtain the sulfur concentrate, and finally the copper and molybdenum separation is carried out by using Na2S copper to obtain copper respectively. Concentrate and molybdenum concentrate.
Lu Lisheng et al. conducted an optional experimental study on a low-grade refractory copper-molybdenum ore, and determined the copper-molybdenum mixed flotation-copper-molybdenum separation-selection molybdenum tailings copper-selection copper tailings return copper-molybdenum mixed flotation In the process, the collector uses isobutyl yellow drug instead of BK301C, and the process and drug addition points are appropriately adjusted to facilitate the stabilization and improvement of the molybdenum and copper selection indexes, and finally obtain an ideal technology.
Wang Ligang et al. conducted a mineral processing experiment on a copper-molybdenum ore with a high oxidation rate in Tibet. The results show that Dy-1 oil is used as the collector and the fuser as the foaming agent, and the gangue mineral and the phosphorox-inhibiting galena are treated with water glass. Good comprehensive technical indicators.
According to the characteristics of a molybdenum concentrate with low grade and high oxidation rate, Ku Jiangang et al. conducted a pressure alkali leaching test. The results show that when continuous leaching with atmospheric leaching, it can not only ensure the leaching rate of molybdenum reaches more than 95%, but also reduce the cost of the medicament. At the same time, the concentration of molybdenum in the leaching solution can be greatly improved.
Zhao Ping conducted a mixed flotation test study on a refractory molybdenum mine. The principle process of mixed flotation of molybdenum sulfide mineral and molybdenum oxide mineral is adopted. After the concentrate is concentrated, it is heated and selected under high alkalinity, and the selected concentrate is acid-leached to remove carbonate and other acid-soluble gangue minerals. Molybdenum concentrates with molybdenum grades and recovery rates of 45.65% and 70.68%, respectively, were obtained.
In view of the fine grain size and high lead content of a molybdenite ore, Xu Yinxing used water glass and phosphonox as inhibitors, fusel as foaming agent and Dy-1 oil as collector. A systematic experimental study. Since the selected tailings after regrind of the coarse concentrate are cleaned twice, the tailings are directly discarded, avoiding the formation of a vicious cycle of sulfide minerals such as galena in the flotation circuit, and finally obtaining the molybdenum grade. More than 57.00%, high-quality molybdenum containing less than 0.06% lead, indicating that the application effect of these measures is very significant.
Song Chengying et al. conducted a systematic experimental study on the leaching process of low-grade molybdenite ore. Under alkaline conditions, the molybdenum ore is not calcined, and the molybdenum disulfide is converted into sodium molybdate by oxygen oxidation. The filtrate is acidified and extracted to obtain metal molybdenum; for this leaching process, the researcher investigates The reaction time, reaction pressure, reaction temperature, sodium hydroxide concentration and stirring speed affect the molybdenum leaching rate. The optimization of the process conditions makes the leaching rate of molybdenum reach over 99%. The test results are quite satisfactory.
Lu Xinlei et al. conducted a flotation column sorting test on a selected tailings of a molybdenum ore. The semi-industrial cyclone-static microbubble flotation column is used as the sorting equipment, and the rough selection and two selected processes are used to improve the sorting index of the selected tailings re-grinding and re-election, and simplify The process of rough selection, 1 selection, and 6 selected flotations on the site obtained the concentrate product with a molybdenum grade of 38.59% and a recovery rate of 23.26%, compared with the on-site flotation machine sorting technical index. The molybdenum grade and molybdenum recovery rate increased by 1.3 percentage points and 4.72 percentage points, respectively.
Shi Weihong conducted a systematic flotation test on a nickel- poor ore. In the test, the combination of sodium carbonate, water glass and CMC was used to inhibit the easy-floating gangue minerals, control the slime trend, and reduce the adverse effects of the slime on the nickel flotation process. 2 times of rough selection, 1 sweeping, and 3 selections, the sorting index of nickel concentrate grade 3.03% and nickel recovery rate 78.67% was obtained.
Aiming at the difficulty of enrichment of a complex refractory silicon-nickel ore in southern China by mineral processing method, Che Xiaokui et al. conducted a leaching test using atmospheric pressure acid leaching. In the grinding fineness of 0.074 mm accounted for 78.60%, liquid-solid ratio of 6:1, sulfuric acid concentration of 2.60 mol / L, stirring strength of 170 r / min, leaching temperature of 60 ° C, leaching for 6h, leaching nickel in the noble liquid The leaching rate is about 86%, the leaching residue contains about 0.12% of nickel, and after three times of extraction of the leaching solution, the Ni 2+ concentration can reach the requirement of sinking nickel.
3. Separation of other non-ferrous metal ores
Li Zhiwei et al. conducted a wet extraction of vanadium pentoxide from a vanadium ore in Henan Province, using a strong acid leaching-solution extraction-sulfuric acid stripping-ammonia vanadium-calcination vanadium process, using 1% oxidizing agent sodium chlorate, grinding The fineness of the mineral is 65%-0.074mm, the leaching temperature is 90°C, the liquid-solid ratio is 1:1, the sulfuric acid is 30%, and the leaching time is 10h. The leaching rate of vanadium reaches 92.50%; the leaching solution is P-204. P- 507, TBP solution was extracted and sulfonated kerosene, sulfuric acid solution was back extracted, via oxidation, ammonia precipitation, pyrolysis, comprehensive recovery to a purity of 98.56%, vanadium is greater than 85%.
Gao Yude et al. conducted a mineral processing test on a scheelite in Hunan. Adopting the preferred float sulphur-white tungsten at room temperature rough selection-tungsten concentrate concentrate heating process and sodium carbonate-water glass-F9 combination pharmacy system, the tungsten content is 0.39%, and the distribution rate of tungsten in scheelite is 85%. The raw ore of the right and left has obtained the mineral processing technical index of 67.35% of tungsten concentrate grade and recovery rate of 80.09%.
Zhang Aiping conducted a flotation test on a high-sulfur scheelite ore. The results show that under the condition of grinding fineness 70%~75% -0.074mm, pre-flotation desulfurization and flotation of white tungsten at normal temperature can obtain the ideal index of 62.87% of scheelite concentrate and recovery rate of 84.33%.
Liu Meihua conducted a tailing test study on the characteristics of a low-grade tin ore using three different re-election methods: spiral chute, jigging and shaker. The results show that the shaker tail throwing is an effective method for pre-selection of the mine. The grade of tin coarse concentrate is increased from 0.37% to 3%, the recovery rate is 72.37%, the tailings yield of thrown is 60%, and the tin is in the tailings. The loss in the case was only 15.89%, which provided favorable conditions for subsequent tin recycling operations.
Sun Yang et al. conducted a beneficiation test on a shale ore in Shangnan, Shaanxi Province. It was found that the use of dextrin can effectively inhibit pyrite in the ore, using a mixture of ethyl sulphide, butyl sulphate and butylammonium. The collecting agent can make the pyrite and the strontium minerals be well separated. By re-grinding the coarse concentrate, the strontium minerals in the coarse concentrate can be separated as much as possible, thereby improving the sorting technical index.
Cai Zhenlei et al. conducted a rare earth ore dressing test on the tailings obtained by magnetic roasting and weak magnetic separation of the strong magnetic separation coarse concentrate in Baotou Steel Concentrator. The results showed that the mixed flotation concentrate was obtained by pre-decarburization and mixed flotation, and after 1 rough selection, 3 selections and 1 sweeping, the final REO grade was 64.4 1%, and the recovery rate was 18.13%. Rare earth concentrate products.
Yu Xiulan et al. studied the reaction principle and process of adding carbon and chlorination to extract rare earth from the tailings of Baotou Steel Concentrator after desulfurization with A1cl 3 or MgO, and investigated the time of carbon thermal chlorination and the extraction rate of rare earth from defluorination agent. influences. Results Yuan Ming, carbon chlorination at 700 ° C for 2 h. When AlCl 3 is used as the defluorination agent, the rare earth extraction rate can reach 77%; when Mg0 is used as the defluorination agent, the rare earth extraction rate can reach 84%.
In order to effectively reduce the impact of slime on rutile flotation, Gao Likun and others conducted a reverse flotation test on a refractory rutile mine. The results show that the inhibition of rutile by aluminum sulfate and the reverse flotation of sodium oleate can discard a certain amount of foam products, wherein the rutile grade is 0.39%, the -0.010 mm fraction removal rate is 74.79%, which is the rutile positive. Flotation has created favorable conditions; deliming 20 pairs of rutile for positive flotation, after 1 rough selection can obtain a grade of 20.30%, a recovery rate of 83.88% rutile coarse concentrate.
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