硫精矿浮选—焙烧综合利用硫酸烧渣中铁试验研究

为综合利用硫酸烧渣中的铁,采用浮选—焙烧工艺对硫酸烧渣原料硫精矿进行提纯除杂试验研究,考察了磨矿细度、抑制剂、捕收剂等对试验指标的影响。结果表明：在磨矿细度-0.074 mm占70%条件下,浮选作业添加高效抑制剂抑制脉石矿物,采用丁基黄药作为捕收剂,提高了精矿品位;闭路浮选试验获得的精矿进一步焙烧,通过控制适宜的焙烧条件,获得的硫酸烧渣铁品位达到65%以上,含硫低于0.4%,可作为铁精矿直接销售。该工艺能综合回收铁,使硫酸烧渣资源得到充分利用,可为企业带来显著的经济效益。     

硫酸烧渣再回收铁试验研究

通过对硫酸烧渣试验研究,确定磁选 浮选的工艺流程回收硫酸烧渣中的铁和硫等元素,取得了较好的技术指标.     

硫酸烧渣选铁工艺研究

对硫酸烧渣的物质组成,铁的赋存状态及矿物工艺特征进行了研究,并找出了既简便又合理的烧渣选铁工艺流程,取得了较好的试验指标。     

硫酸烧渣利用工艺通过鉴定

最近,中国有色金属工业协会在广东云浮召开了硫酸烧渣综合利用研究科技成果鉴定会,中国有色金属工业协会副会长周菊秋出席了会议。 该项目由湖南有色金属研究院和广东云浮硫铁矿企业集团公司共同承担。专家组充分肯定了该工艺的独创性和先进性,认为新工艺结合硫酸烧渣的物化特性,采用烧渣-筛分-漂洗-细磨-超细     

从硫精矿烧渣中回收黄金的探讨

叙述了用硫精矿焙烧,焙砂氰化提金工艺处理的过程,对回收烧渣中金银的研究有一定的参考价值。     

硫酸烧渣提铁降硫工艺试验研究

对某化工厂黄铁矿制酸烧渣的工艺矿物学性质进行检测分析,确定采用筛分水洗—溜槽重选—浮选脱硫选矿工艺流程,从烧渣中分选获得了铁品位为65.29%,铁回收率为83.74%,硅、硫含量分别为4.89%、0.15%的高品质铁精矿产品。     

硫精矿烧渣中铁、金综合回收试验研究

含硫51.6%、金0.98 g/t的硫精矿经过焙烧,在温度850℃条件下,获得含铁66.81%、金1.31 g/t的烧渣。烧渣经水浸除硫后,进行了硝酸铅预处理、氰化浸金,金的浸出率达51.9%。     

白钨精矿提纯除杂

本文介绍了净化不合格白钨精矿除去磷、砷和硫的研究结果。白钨精矿中大部分砷(占总量的70-94.3%)是毒砂(砷黄铁矿),氧化物形态的砷有臭葱石.磷主要以磷灰石形态存在.硫     

硫酸烧渣加压浸取铁

采用自制硫化物助浸剂在密闭反应釜中还原浸出硫酸烧渣中的铁。分别进行了助浸剂用量、硫酸用量、始酸浓度、反应温度、搅拌速度、反应时间等条件试验,考察各因素对浸取效果的影响。结果表明:当酸浸渣25.0g,助浸剂过剩系数1.1,硫酸过剩系数1.4,始酸浓度3.5mol/L,反应温度95℃,搅拌速度800r/min,反应时间3h时,铁浸取率达99.4%,助浸剂有效利用率达98.9%。     

硫酸烧渣还原浸取铁

采用硫化物作助剂还原浸出硫酸烧渣中的三价铁。采用L9(43)四因素三水平正交试验,考察助剂用量、硫酸用量、温度、时间对浸出效果的影响,并确定最佳配比。结果表明:影响的显著顺序为助剂用量>时间>硫酸用量>温度。在下述最佳条件下铁浸取率可以达到87.8%:起始液固比2∶1、搅拌转速1 300r/min、助剂用量17.2g、硫酸23mL、85℃反应3h。     

Copper recovery from chalcopyrite concentrates by the BRISA process

The technical viability of the BRISA process (Biolixiviación Rápida Indirecta con Separación de Acciones: Fast Indirect Bioleaching with Actions Separation) for the copper recovery from chalcopyrite concentrates has been proved. Two copper concentrates (with a copper content of 8.9 and 9.9 wt.%) with chalcopyrite as the dominant copper mineral have been leached with ferric sulphate at 12 g/L of ferric iron and pH 1.25 in agitated reactors using silver as a catalyst. Effects of temperature, amount of catalyst and catalyst addition time have been investigated. Small amounts of catalyst (from 0.5 to 2 mg Ag/g concentrate) were required to achieve high copper extractions (>95%) from concentrates at 70 °C and 8–10 h leaching. Liquors generated in the chemical leaching were biooxidized for ferrous iron oxidation and ferric regeneration with a mixed culture of ferrooxidant bacteria. No inhibition effect inherent in the liquor composition was detected. The silver added as a catalyst remained in the solid residue, and it was never detected in solution. The recovery of silver may be achieved by leaching the leach residue in an acid-brine medium with 200 g/L of NaCl and either hydrochloric or sulphuric acid, provided that elemental sulphur has been previously removed by steam hot filtration. The effect of variables such as temperature, NaCl concentration, type of acid and acidity–pulp density relationship on the silver extraction from an elemental sulphur-free residue has been examined. It is possible to obtain total recovery of the silver added as a catalyst plus 75% of the silver originally present in concentrate B (44 mg/kg) by leaching a leach residue with a 200 g/L NaCl–0.5 M H₂SO₄ medium at 90 °C and 10 wt.% of pulp density in two stages of 2 h each. The incorporation of silver catalysis to the BRISA process allows a technology based on bioleaching capable of processing chalcopyrite concentrates with rapid kinetics.     

Selective recovery of copper from copper-nickel sulfide concentrates by applying segregation technology

Copper can be selectively segregated from roasted copper-nickel sulfide concentrates. Conditions are reported for desulfurizing-roasting followed by segregation treatment to selectively reduce and segregate copper without nickel contamination. A process is described which is applicable for treating low grade sulfide concentrates such as are produced in Botswana (4.6 pct Cu, 3.1 pct Ni) and higher grade concentrates that can potentially be produced from sulfide deposits in Minnesota (13 pct Cu, 2.5 pct Ni). Steps include roasting to remove at least 97 pct of the sulfur, segregation treatment of calcine and flotation to recover a copper product, a middlings fraction which is recycled to the desulfurizing roast and a tailings fraction which can be subsequently treated for nickel recovery. LAMAR D. COFFIN, formerly Senior Research Metallurgist at AMAX Base Metals R&D, Inc., has retired.     

浮选金精矿微生物氧化渣提金工艺试验研究

针对某含砷、含硫微细包裹类型难处理金矿石提金工艺进行了研究。该矿石常规氰化金浸出率仅为12.54%。通过对其浮选金精矿进行生物氧化预处理,并对生物氧化渣提金工艺的氰化法、氯化法、硫脲法进行了浸金对比试验。其结果表明,氯化法浸金效果最好,金浸出率可达94.73%,指标优于其他两种工艺。     

Recovery of iron oxide concentrate from high-sulfur and low-grade pyrite cinder using an innovative beneficiating process

A high quality iron oxide concentrate, suitable as a feed for blast and electric reduction furnaces was recovered from high-sulfur and low-grade pyrite cinder. Pyrite cinder was treated with a hot strong alkali solution which dissolved most of the silica and alumina, leaving a solid residue and forming a precipitate which was an acid-soluble salt of aluminosilicate hydrate. The residue and precipitate were then treated with a diluted aqua regia to dissolve the precipitate and other impurities. Impurities such as the soluble silicates, aluminates and heavy metals can be extracted and produced high value-added products with several simple methods in later research. After analyzing the chemical composition and crystalline phase of raw material, the effects of different parameters on recovery efficiency of iron were carried out. The optimum reaction parameters were proposed as the following: melting at 200 °C for 1.0 h with 3.0 M NaOH in a liquid-solid ratio of 1.0:1.2 at atmospheric pressure, followed by leaching with aqua regia solution (3.0 vf.%) in a ratio of liquid-to-solid of 1.2:1.0 at room temperature for 1.5 h. In optimum parameters mentioned above, the raw residue only containing 38% Fe could be purified to about 65% and the recovery yield of Fe was 86%. Furthermore, the sulfur content of product could be reduced to 0.090% from 8.0%. To reveal the mechanism of this process, the phase transformation and microstructure of the process were studied by XRD and SEM with EDS.     

某硫精矿焙烧—氰化回收金试验研究

某硫精矿中含铁42.3%、有效硫47.34%、金0.72 g/t,对其进行了焙烧—氰化浸金试验研究。其结果表明:在750℃下焙烧2 h,获得含铁61.42%、金品位1.04 g/t的焙砂;在硝酸铅添加量为300 g/t条件下,对该焙砂进行氰化浸金时,浸渣金品位可降低至0.33 g/t,金浸出率可达68.27%。     

硫酸渣脱硫制备高品质铁精矿研究进展

利用硫酸渣脱硫制备高品质铁精矿具有良好的的工业利用价值,不仅可以解决烧渣的综合利用问题,而且可以解决其对环境影响的问题.本文系统介绍了硫酸渣脱硫制备高品质铁精矿的脱硫技术方法、工艺流程及最新研究进展.硫酸渣脱硫方法主要有化学法、联合法和生物法.化学法主要包括酸浸、碱浸,联合法可分为碱浸-酸浸、浮选-磁选、重选-浮选、磁化焙烧-磁选等联合工艺方法.比较了这些方法的工艺路线及存在的优缺点,提出了生物法具有良好的工业应用前景,展望了该方法未来的研究方向为：高效脱硫菌种的选育,生物脱硫液的循环使用,硫酸渣生物脱硫协同回收有价金属,生物脱硫过程基础理论及工程化技术研究等.     

超细磨-树脂矿浆法从黄铁矿烧渣中回收金的研究

研究了超细磨－树脂矿浆法从黄铁矿烧渣中提金新工艺，以氧化铝球为研磨介质，考察了磨矿粒度，液固比，磨浸时间，助磨剂，助浸剂等因素对浸出效果的影响，以AM－2b树脂为吸附剂，弱酸性硫脲为解吸剂从氰化矿浆中回收金，分别用静态法，动态法考察了时间，温度，酸度，流速等因素对AM－2b树脂吸附金和解吸金的影响，研究结果表明，在适宜的条件下，金的回收率可达85％。     

Removal of impurities from crude nickel sulfide by selective dry chlorination and leaching

Finely ground impure nickel sulfide is treated with dry gaseous chlorine at about 600°F to selectively chlorinate copper, cobalt and iron. These chlorinated impurities are leached from the nickel sulfide by agitating in water at about 70°F and bubbling chlorine gas into the slurry to maintain a redox potential of about +500 mV.