Vanadium recovery from stone coal through roasting and flotation

A new process for vanadium recovery from stone coal by roasting–flotation was investigated based on the mineralogy. The process comprised four key steps: decarburization, preferential grinding, desliming and flotation. In the decarburization stage, roasting at 550 °C effectively avoided the negative effect of the carbonaceous materials in raw ore and generation of free CaO from calcite decomposition during roasting. Through preferential grinding, the high acid-consuming minerals were enriched in the middle fractions, while mica was enriched in the fine and coarse fractions. Through flotation, the final concentrate can be obtained with V₂O₅ grade of 1.07% and recovery of 83.30%. Moreover, the vanadium leaching rate of the final concentrate increased 13.53% compared to that of the feed. The results reveal that the decarburization by roasting at 550 °C is feasible and has little negative impact on mica flotation, and vanadium recovery from stone coal is conducive to reducing handling quantity, acid consumption and production cost.     

云南某氧化银矿选矿试验研究

采用浮选+全泥氰化工艺对云南省某银矿石进行提银选矿试验研究。矿石中银多金属复硫盐经浮选富集得到银精粉;浮选尾矿再全泥氰化提银。浮选作业所得银精粉品位5761.54 g/t,回收率21.91%;浮选尾矿全泥氰化浸渣品位19.28 g/t,浸出率78.85%。综合回收率84.12%。银多金属复硫盐是影响浸出率的关键因素,预先将其有效浮选,有利于提高银氰化浸出率。     

河北某石英脉型金矿石选矿试验研究

河北某石英脉型金矿石金品位5.4 g/t,银品位6.4 g/t。针对该矿石性质,开展浮选试验,在最佳药剂制度条件下浮选闭路试验获得精矿金回收率为78.9%,银回收率35.6%,金品位44.0 g/t,银品位23.5 g/t。为提高选矿指标,开展重选与浮选工艺联合试验。与单一浮选工艺相比,重、浮联合工艺获得混合精矿金回收率提高6.8%,银回收率提高2.2%。     

高砷富银铋硫多金属精矿综合回收实验研究

某高砷富银铋硫矿为硫化矿混合精矿,表面受到浮选药剂污染,各种矿物之间的可浮性相近,给分离带来不利影响.采用"混合精矿加温脱药-脱药精矿铋银优先浮选-铋银尾矿砷硫活化浮选"工艺流程进行处理,采用高效银铋捕收剂SAC,全流程实验获得的银铋精矿含银4386 g/t、铋13.06％、砷0.61％,回收率为银88.52％,铋85...     

焙烧-氰化浸出锰银矿中银的扩大试验

采用添加剂焙烧-氰化浸出中试处理锰银矿。回转窑连续运转80 h，物料焙烧时间30±5 min，所得焙砂产率85.54%，焙砂中银含量237.73 g/t、锰含量24.68%，银回收率99.19%，锰回收率98.70%；焙砂经500 L反应釜直接氰化浸出，浸出液固比2:1~2.2:1，时间6~15 h，氰化钠用量700 g/t原矿，所得银浸出率86.5%，氰化尾渣满足冶金用锰矿石标准。     

ACTIVATION AND DEPRESSION OF CALCITE IN CALCIUM MINERALS FLOTATION

１ＩＮＴＲＯＤＵＣＴＩＯＮＳｃｈｅｅｌｉｔｅｏｒｅｓａｌｗａｙｓｃｏｎｔａｉｎｆｌｕｏｒｉｔｅａｎｄｃａｌｃｉｔｅ．Ｗｈｅｎｔｈｅｈｅａｄｇｒａｄｅｏｆｆｌｕｏｒｉｔｅｉｓｕｐｔｏ１５％ＣａＦ２,ｔｈｅｆｌｕｏｒｉｔｅｉｓｏｆｅｃｏｎｏｍｉｃｖａｌｕ...     

锌浸出渣中银的浮选回收综述

锌浸出渣是湿法炼锌过程中产生的固体废弃物,其含有的银具有巨大的经济价值,浮选回收银是重要的途径。锌浸出渣中浮选回收银,面临着含银矿物种类多而分散、矿物粒度细、酸性强及难免离子含量高等综合叠加影响。部分锌浸出渣可采用直接浮选法回收银,但对银赋存状态异常复杂的锌浸出渣,直接浮选往往难以实现银的高效选别回收,多采用进行一定的预处理后再浮选的间接浮选法。常用的预处理技术主要有浆洗、磨矿、外加载体、焙烧及热酸浸出等,预处理可针对性的调控矿浆环境、改变矿物的赋存状态及矿物表面性质,以提高选择性和捕收能力。各种预处理都有其独特的优缺点,合适的预处理技术结合湿法炼锌工艺或预处理技术的有机联合应用,对浮选回收锌浸出渣中的银将更为有效。     

提高锌浸出渣中银浮选回收率的研究

某湿法炼锌厂低酸度锌浸出渣中53.8%的银存在于难完全回收的闪锌矿上,其回收是提高浮选回收率的关键。经对比浮选和正交试验获得了浮选粗选最佳药剂制度,捕收剂为丁铵黑药(900g/t)和Z-200(50 g/t),载体活性炭(2000 g/t),起泡剂2#油(100 g/t)。一粗一精一扫开路试验表明,在非强充气和非强搅拌条件下,浮选精矿银品位为8210 g/t,较现有工艺(3000 g/t)大幅提高;银回收率为64.7%,与现有工艺(60%~64%)相当。     

吉林某金矿选矿工艺对比实验研究

对吉林某金矿进行了重选、无毒浸出和浮选3种选矿工艺对比实验,研究了磨矿细度、药剂用量、矿浆浓度等因素对选矿效率的影响.结果表明,无毒浸出工艺回收率为76.13％,重选工艺基本无优先选别作用;采用浮选工艺,在磨矿细度为-0.074 mm含量为90.52％,捕收剂异戊基黄药与丁铵黑药的药剂用量均为60 g/t,矿浆浓度为4...     

某含银高硫铜矿的综合回收试验

某含银高硫铜矿含铜0.76%、硫24.35%及银34.92 g/t,有价矿物种类多、矿石性质复杂,采用抑硫优先浮选铜-活化浮选硫的原则工艺流程进行试验,配合石灰作为硫化铁矿物抑制剂以及筛选出丁基黄药+酯-105作为硫化铜矿物的组合捕收剂,强化了银在铜精矿中的富集。在选定工艺条件下,可获得铜品位21.60%、银品位602.84 g/t的铜精矿(铜和银回收率分别为89.30%和54.39%),硫品位45.60%、银21.55 g/t的硫精矿(硫和银回收率分别为89.79%和29.59%),实现了铜、硫和银的综合回收利用。     

从氧化铜矿氨浸渣中浮选回收银

某高碱性含银氧化铜矿氨浸渣含银145.55 g/t。对氨浸渣进行了浮选回收银的试验,结果表明,采用三粗一扫两精的流程,经过闭路流程试验,可获得达到计价标准的银精矿,其银品位为1999.58 g/t、回收率80.78%。     

富银低品位锰矿和闪锌矿的联合氧压酸浸

进行了硫化锌和含银低品位锰矿联合氧压酸浸的小型试验,以考察影响浸出的各种因素,诸如浸出温度、浸出时间、氧分压和搅拌速度.实验在2L的高压釜内进行.实验结果表明,硫化锌和含银低品位锰矿能相互促进浸出,但这种耦合作用须在一定条件下才会起作用.作者对一些浸出做了探讨.为下一步扩大试验的需要,本文给出了合理的浸出条件:浸出温度, 110℃; 浸出时间,2h; 氧分压, 0.6MPa; 硫酸用量, 1.2倍理论用量.     

Recovery of iron from waste ferrous sulphate by co-precipitation and magnetic separation

Magnetite concentrate was recovered from ferrous sulphate by co-precipitation and magnetic separation. In co-precipitation process, the effects of reaction conditions on iron recovery were studied, and the optimal reaction parameters are proposed as follows: n(CaO)/n(Fe²⁺) 1.4:1, reaction temperature 80 °C, ferrous ion concentration 0.4 mol/L, and the final mole ratio of Fe³⁺ to Fe²⁺ in the reaction solution 1.9–2.1. In magnetic separation process, the effects of milling time and magnetic induction intensity on iron recovery were investigated. Wet milling played an important part in breaking the encapsulated magnetic phases. The results showed that the mixed product was wet-milled for 20 min before magnetic separation, the grade and recovery rate of iron in magnetite concentrate were increased from 51.41% and 84.15% to 62.05% and 85.35%, respectively.     

Sulphidization flotation for recovery of lead and zinc from oxide-sulfide ores

A new flowsheet was developed to recover the valuable minerals from oxide or oxide-sulfide ores of lead and zinc. The flowsheet consisted of flotation of sulfide minerals, desliming and sulphidization-flotation of oxide minerals. The corresponding reagent system and techniques to the flowsheet were investigated. Batch and continuous tests show that the dosage of sodium sulfide, temperature, and collector type are main affecting factors on the recovery of smithsonite and cerussite. For the flotation of cerussite, there is an appropriate dosage of sodium sulfide at which the recovery reaches its maximum value. The required sodium sulfide for smithsonite flotation is higher than that for cerussite and the recovery of smithsonite flotation increases with the dosage of sodium sulfide at low level and becomes insensitive at high dosage. The appropriate temperature for smithsonite and cerussite flotation is found to be 25 ～ 40℃. Amines are found to be the effective collectors for the flotation of smithsonite after sulphidization. Investigation also shows that desliming prior to sulphidization-flotation is essential to the effective recovery of smithsonite and cerussite, and the desliming process of two-stage hydrocyclon is well feasible and effective for the treatment of lead-zinc oxide ores. A further treatment on the cerussite flotation concentrate by shaking table is proposed to obtain higher lead grade.     

Behavior of silver and lead in selective chlorination leaching process of gold-antimony alloy

The behavior of silver and lead in the selective chlorination leaching process of gold-antimony alloy was analyzed in detail and appropriate recovery methods were developed.A reduction method by adding gold-antimony alloy powder was adopted to recover silver according to the thermodynamics calculation.The reducing rate of silver can exceed 99%at 80℃for 1.5 h when the dosage of gold-antimony alloy powder is 10%.The dissolution equilibrium curved surfaces of PbSO4 and PbCl2 under different conditions were dra...     

从失效有机铑催化剂中富集铑的新工艺研究

针对失效有机铑催化剂,利用还原-磨选工艺富集铑。在还原过程中,添加剂促进铁晶粒长大并富集铑,再通过磁选分离富集含铑的铁精粉。研究表明最佳工艺参数为:还原温度1200℃,还原时间6 h,添加剂配比10%,煤粉配比5%,球磨时间45 min,磁场强度1.28×105 A/m。利用X射线衍射对磁选铁精粉进行了分析,磁选后主要物相为金属铁、铑,除去了大部分脉石。最终得到铁精矿品位88.67%,回收率92.74%,铑回收率为92.08%。本工艺具有还原温度低、收率高等特点,为失效有机铑催化剂富集提供一种途径。     

Leaching of complex sulphide concentrate in acidic cupric chloride solutions

The chemical analysis of a complex sulphide concentrate by emission spectrometry and X-ray diffraction shows that it contains essentially copper, lead, zinc and iron in the form of chalcopyrite, sphalerite and galena. A small amount of pyrite is also present in the ore but does not be detected with X-ray diffraction. The cupric chloride leaching of the sulphide concentrate at various durations and solid/liquid ratios at 100 ℃ shows that the rate of dissolution of the ore is the fastest in the first several hours, and after 12 h it does not evolve significantly. If oxygen is excluded from the aqueous cupric chloride solution during the leaching experiment at 100 ℃, the pyrite in the ore will not be leached. The determination of principal dissolved metals in the leaching liquor by flame atomic absorption spectrometry, and the chemical analysis of solid residues by emission spectrometry and X-ray diffraction allow to conclude that the rate of dissolution of the minerals contained in the complex sulphide concentrate are in the order of galena 〉 sphalerite 〉 chalcopyrite.     

Extraction of manganese and zinc from their compound ore by reductive acid leaching

Comprehensive utilization of low grade manganese–zinc compound ore containing lead and silver with a method of reductive acid leaching was studied. According to the ?–pH diagram of Mn–Zn–H₂O system, Mn and Zn can be leached simultaneously in the pH range of –2 to 5.61. The results showed that both hydrogen peroxide and sucrose were effective reductants which could intensify the simultaneous leaching of Mn and Zn into leachate as well as enrich Pb and Ag in the residue. 95.88% of Mn and 99.23% of Zn were extracted when the compound ore was leached with hydrogen peroxide in sulfuric acid media, meanwhile the contents of Pb and Ag in the residue were enriched to 13.21% and 489.36 g/t, respectively. When sucrose was used as the reductant, the leaching efficiencies of Mn and Zn separately achieved 98.26% and 99.62%, and contents of Pb and Ag in the residue were as high as 13.92% and 517.87 g/t, respectively.     

溴化法浸出提取金和银

讨论了溴化法提取金银的热力学条件，用高银金精矿进行溴化法浸出实验，结果表明，一次浸出焙烧矿，金的浸出率９８％，银的浸出率９４％，效果明显。     

低品位铁锰型金银矿的硫脲浸出研究

对铁锰型金银矿采用预先浸锰工艺,使被二氧化锰包裹金银裸露出来,同时降低硫脲浸出金银时的氧化还原电位,有效减少了硫脲消耗量.结果表明,当矿样为200g,经浸锰预处理后,在pH =1.5,电位300mV,亚硫酸钠6g,浸出时间4h的最佳条件下,金、银的浸出率分别为98%、45%,硫脲消耗仅为6kg/t.