Reduction-roasting and ferric chloride leaching of copper converter slag for extracting copper,nickel and cobalt values

In a previous investigation the optimum conditions for recovering copper, nickel and cobalt from converter slag through ferric chloride leaching have been described. The study of various parameters revealed that nickel and cobalt recovery could not be improved beyond 24 to 26% respectively from converter slag, though more than 90% of the copper could be extracted. Further attempts were made to bring the metal values completely into solution through reduction-roasting followed by ferric chloride leaching of the slag. The present work comprises a study of various experimental conditions such as concentration of ferric chloride, duration of leaching, duration of reduction-roasting, temperature and nature of reducing agent, to arrive at the optimal recovery of the metal. Under identical experimental conditions a decrease in copper recovery, but an increase in nickel and cobalt recovery has been observed above a roasting temperature of 750°C. The decrease in copper recovery has been attributed to copper ferrite formation which has been confirmed both by leaching experiments with synthetic mixtures and by X-ray diffraction studies with both slag samples and synthetic mixtures. Recovery of nickel has also shown little decline when solid reductants were used above 850°C whereas cobalt recovery remains nearly the same even above 850°C. Under optimum conditions 80% copper, 95% nickel and 80% cobalt could be recovered by reducing the slag at 850°C with 10 wt % furnace oil, followed by leaching with ferric chloride.     

Recovery of cobalt,nickel and copper from converter slag through roasting with ammonium sulphate and sulphuric acid

Roasting of copper converter slag containing 4.03% copper, 1.98% nickel and 0.48% cobalt with ammonium sulphate open to atmosphere has been carried out in order to achieve sulphation of copper, nickel and cobalt followed by leaching of the metal values as soluble sulphates with water. The effect of parameters such as temperature (200–600°C), time (15–120 min), and amount of ammonium sulphate (0.5–2.5 times stoichiometric) has been studied. Under atmospheric conditions, using 2.5 times the stoichiometric requirement of ammonium sulphate, the recovery of copper, nickel and cobalt was found to be 85%, 81% and 85%, respectively. Similar studies were carried out with sulphuric acid. The influence of experimental variables such as the amount of sulphuric acid (0.25–2 times stoichiometric), roasting temperature (100–300°C) and time (15–120 min) has been studied. Under optimum conditions, i.e., at 150°C and a roasting time of 60 min with the stoichiometric amount of sulphuric acid, recoveries of copper, nickel and cobalt were 95, 90 and 99%, respectively, along with a contamination of 60–80% iron. Removal of most of the iron from the leach liquors has been effected with ammonia liquor and lime as precipitants. A two-stage roasting operation using sulphuric acid, first at 150°C and then at 650°C, has resulted in bringing the iron content down to about 3% in the sulphation product without much affecting the recovery of other metal values.     

铜熔炼渣和吹炼渣混合浮选的生产实践

东营方圆有色金属有限公司采用浮选工艺对铜熔炼渣和吹炼渣进行混合选矿,以回收炉渣中的铜等有价金属。实际生产中采用三段破碎+两段球磨的浮选工艺,通过合理控制破碎粒度、磨矿粒度以及加药量等工艺参数,有效地提高了铜等金属的回收率。     

The kinetics of the sulphuric acid leaching of nickel and magnesium from reduction roasted serpentine

Nickel may be extracted with partial selectivity over magnesium from laterites containing serpentine by reduction roasting followed by sulphuric acid leaching. This paper describes the results of a kinetic study of the sulphuric acid leaching of nickel and magnesium from the reduction roasted serpentine component of a typical laterite. The serpentine used in this work analyzed 1.65% nickel, 20.2% magnesium and 6.10% iron.Initially, leaches were carried out at temperatures of 30, 50 and 70° C to determine the acid requirement for complete nickel extraction using practical leaching conditions (25% solids) under which the acidity drops to a low level. A minimum acid addition of 60 g/l was needed, which gave 80% to 83% nickel extraction from material in which 85% of the nickel was reduced using hydrogen at 700° C. Under these conditions, about 17% of the magnesium was leached at each of the temperatures studied.To facilitate an understanding of leaching kinetics, leaches were then performed using constant acidities (0.1% solids) of 60, 30 and 15 g/l acid at temperatures of 30, 50 and 70° C. Closely sized particles (?65 + 100 mesh) were leached so that magnesium dissolution rates could be tested against established “shrinking core” models.The main conclusions are that, under the experimental conditions, nickel dissolution rates were chemically controlled by either $\text{2}\text{H}{}^{\mathrm{+}}\text{+}\text{1}\text{2}\text{O}{}_2\text{+ 2e →}\text{H}{}_2\text{O}$ or 2 H⁺ + 2e → H₂ occurring at the surfaces of the 40% nickel/iron alloy platelets formed during reduction. The rate controlling process had an activation energy of 11kcal/mole.During extraction of metallic nickel, the dissolution of magnesium follows a “shrinking core” kinetic model, which assumes the reaction is unimpeded by a surface layer of silica - a reaction product. Rates of magnesium dissolution during this stage of the reaction (up to 25% dissolved) were chemically controlled with an activation energy of 12 kcal/mole.At magnesium extractions above 25%, at which point all the metallic nickel had been extracted, the rate of silicate attack was limited by diffusion through a silica coating attached to particle surfaces.The results indicate that selectivity for nickel dissolution over that of magnesium does not depend strongly on acidity and temperature at the levels investigated.     

Recovery of metal values from copper converter and smelter slags by ferric chloride leaching

A study of the recovery of copper, nickel and cobalt from copper converter and smelter slags by leaching with ferric chloride is reported. The converter slag from Ghatsila, India contained 4.03% copper, 1.99% nickel and 0.48% cobalt and the smelter slag contained 1.76% copper, 0.23% nickel and 0.19% cobalt. Various parameters including the effect of stirring, leaching time, leaching temperature, concentration of ferric chloride, solid-liquid ratio and particle size, on the extraction of copper, nickel and cobalt have been studied. 92% copper, 28% nickel and 24% cobalt could be extracted from converter slag under optimum conditions, whereas 54% copper, 71% nickel and 44% cobalt could be extracted from smelter slag.     

从含钴转炉渣中回收钴的方法及应用

为实现钴资源的综合利用, 对含钴转炉渣回收钴的方法及其研究进展与应用情况进行了综述。冶炼时, 钴主要以化学溶解的方式进入转炉渣, 根据其在冶炼过程中进入转炉渣情况及转炉渣存在的现状, 介绍了浮选、火法冶炼、湿法分离、萃取法、微生物浸出等分离回收方法, 并讨论了各种分离回收方法的优势及局限性。其中, 具有高选择性、高回收率、流程简单、操作连续、易于实现自动化等优点的溶剂萃取法, 以及具有污染小、成本低、高效等特点的微生物浸出法, 拥有较好的应用前景。     

Pressure leaching of copper minerals in perchloric acid solutions

The leaching of covellite (CuS), chalcocite (Cu₂S), bornite (Cu₅FeS₄), and chalcopyrite (CuFeS₂) was carried out in a small, shaking autoclave in perchloric acid solutions using moderate pressures of oxygen. The temperature range of investigation was 105° to 140°C. It was found that covellite, chalcocite, and bornite leach at approximately similar rates, with chalcopyrite being an order of magnitude slower. It was found that chalcocite leaching can be divided into two stages; first, the rapid transformation to covellite with an activation energy of 1.8 kcal/mole, followed by a slower oxidation stage identified as covelite dissolution with an activation energy of 11.4 kcal/mole. These two stages of leaching were also observed in bornite with chalcocite (or digenite) and covellite appearing as an intermediate step. No such transformations were observed in covellite or chalcopyrite. Two separate reactions were recognized as occurring simultaneously for all four minerals during the oxidation process; an electrochemical reaction yielding elemental sulfur and probably accounting for pits produced on the mineral surface, and a chemical reaction producing sulfate. The first reaction dominates in strongly acidic conditions, being responsible for about 85 pct of the sulfur released from the mineral, but the ratio of sulfate to elemental sulfur formed increases with decreasing acidity. Above 120°C the general oxidation process appears to be inhibited by molten sulfur coating the mineral particles; the sulfate producing reaction, however, is not noticeably affected above this temperature. For chalcopyrite, activation energies were determined separately for the oxygen consumption reaction and for the production of sulfate, with values of 11.3 and 16.0 kcal/mole respectively. This paper is based upon a thesis submitted by F. LOEWEN in partial fulfillment of the requirements of the degree of M.A. Sc. in Metallurgical Engineering at The University of British Columbia.     

Recovery of cobalt, nickel, and copper from sea nodules by their extraction with alkylphosphines

The paper describes a process for the recovery of pure Co(II), Ni(II), and Cu(II) from a hydrochloric acid solution of polymetallic sea nodules. An overnight contact of the nodule powder with 4 M HCl almost quantitatively leaches these metal ions. From the leachate, Co(II) and Cu(II) were recovered by their extraction with Cyanex 923 and Ni(II) by extraction with Cyanex 301. Co(II) and Cu(II) are partitioned in the organic phase as H₂CoCl₄·2 Cyanex 923 and CuCl₂·2 Cyanex 923, whereas Ni(II) is extracted in the form of NiR₂ (HR=Cyanex 301). A solution of 0.001 M H₂SO₄ is used for the stripping of Co(II) and Cu(II) and 5% NH₄Cl in 75% NH₃ for Ni(II). Both extractants are found to be stable toward prolonged contact with HCl and show negligible loss in their extraction capacity even after recycling them for 20 cycles. The partition data have been utilized in developing conditions for the separation of Co(II), Ni(II), and Cu(II) mutually and from other metal ions, namely Ti(IV), Al(III), Fe(III), Mn(II), and Zn(II). The procedure of separation thus developed has been extended for the recovery of around 90% Co(II), Ni(II), and Cu(II) from sea nodules. The purity of the metal ions thus obtained is around 99%.     

Comparative leaching of minerals by sulphuric acid in a Chinese ferruginous nickel laterite ore

The Yuanjiang nickel laterite ore containing mainly maghemite, goethite and lizardite was leached by sulphuric acid at atmospheric pressure and the residues were characterized using X-ray diffraction and scanning electron microscopy/X-ray energy dispersive spectroscopy. The relationship was discussed between the extraction of nickel, cobalt, iron, magnesium, aluminum, and the dissolution behaviour of the laterite minerals; as well as the extent of congruency of nickel, cobalt and iron extraction. The results show that the solubility of the laterite minerals in sulphuric acid decreases in the following order: lizardite > goethite > maghemite > magnetite ≈ hematite > chromite ≈ ringwoodite. Lizardite dissolved rapidly in 0.6 mol/L sulphuric acid at 60 °C whilst goethite dissolved completely in 2.5 mol/L sulphuric acid at 80 °C. The dissolution of the primary mineral maghemite was slow, but increased with increasing acid concentration and leaching temperature. Magnetite dissolved more slowly than maghemite; and hematite was only dissolved in > 6.2 mol/L sulphuric acid at 105 °C. Chromite and ringwoodite were not dissolved. The leaching behaviour of the laterite minerals may be explained by the bond strength differences of Me–O and the substitution of metal cations in the mineral structure.     

Decomposition of iron,cobalt, nickel,and copper formates

Conclusions Investigations have shown that the decomposition of metallic formates is accompanied by a loosening of the structure of their crystals, the rate of the process depending on the nature of the metals. In the stages of decomposition the crystals acquire a porous, spongy structure and disintegrate under the action of gases being evolved and internal stresses being generated. Under these conditions the specific surfaces of specimens rapidly grow. The mean particle size of the end products of the reaction is 0.1–0.3, with a specific surface of about 60 m²/g. Exceptions are nickel and copper, whose particles have a strong tendency to sinter together, with the formation of agglomerates. X-ray studies have established that at various stages of the thermal decomposition process the intensities of reflections characterizing the starting salts change, and new reflections appear, indicating the formation of metallic phases in finely divided form. The method of thermal decomposition of metallic formates is suitable for the preparation of very fine powders of pure metals, from which various sintered materials can be produced.Deceased.Translated from Poroshkovaya Metallurgiya, No. 5(173), pp. 7–13, May, 1977.     

Organic acid leaching of base metals from copper granulated slag and evaluation of mechanism

A hydrometallurgical method is discussed to selectively extract base metals such as copper, cobalt, nickel and iron from the copper granulated slag (0.53% Cu) at atmospheric pressure. It involves first-stage leaching of slag with organic (citric acid) to selectively recover cobalt, nickel and iron. The residue containing high copper was subjected to second-stage leaching with inorganic (sulphuric) acid. Leaching parameters such as acid concentration, pulp density, temperature and time were optimised to extract metals from the granulated slag. A maximum recovery of 4.47% Cu, 88.3% Co, 95% Ni and 93.8% Fe were obtained in first-stage leaching with 2?N citric acid at room temperature using 10% pulp density (w/v) in 8–9?h. On subjecting the leach residue to the second-stage leaching with 2?M sulphuric acid, 66–72% Cu was recovered in 4?h. The kinetics of the metal leaching from the slag was established by the XRD and SEM–EDAX studies of the residues.     

铜闪速熔炼烟灰酸浸渣制备磁性材料

为了综合回收利用铜闪速熔炼烟灰,对铜闪速熔炼烟灰酸浸渣制备磁性材料进行了探索性研究.对熔炼烟灰酸浸渣进行了不同浓度盐酸浸出渣和除铅等工艺处理,研究了各工艺下浸出渣的磁性能.结果表明:杂质的去除有利于饱和磁化强度和剩磁的提高,但导致矫顽力下降.盐酸浸出处理的最佳性能出现在当盐酸浓度为3 mol/L时,饱和磁化强度和剩磁分别为10.965 A·m2/kg和0.964 A·m2/kg,矫顽力为3.33×103 A/m;除铅处理后浸出渣的磁化强度和剩磁分别达到11.065 A·m2/kg和0.94 A·m2/kg,矫顽力为3.228×103 A/m.      

红土镍矿微波水热法浸提镍钴

采用微波水热盐酸浸出方法对腐泥土型红土镍矿提取镍钴进行了研究,详细探讨了焙烧预处理、微波水热浸出温度和浸出时间对镍钴浸出率的影响.对于300℃焙烧预处理后的红土镍矿,微波水热温度为50℃,浸出时间为1 h时,镍的浸出率高达93.65%,钴的浸出率为87.86%.红土镍矿的微波水热浸出体系与普通水热浸出体系相比,镍和钴的浸出效果更好.研究表明,扩散过程是镍、钴浸出过程的主要限制环节.     

蛇纹石酸浸提镍过程动力学研究

研究了在蛇纹石硫酸浸出镍过程中,浸出剂初始浓度、搅拌速度、矿石粒径和浸出温度对镍浸出率的影响,并对浸出过程动力学进行了分析.结果表明:浸出剂初始浓度、浸出温度和矿石粒径对镍浸出率有显著影响,搅拌速度对镍浸出率影响不大.硫酸浸镍过程符合关系式1-(1-η)1/3=kt,反应频率因子为190.56,表观活化能为48.79 kJ/mol,浸出过程受化学反应控制.     

The leaching kinetics of cobalt and nickel from aluminum-coprecipitated products

The leaching behavior of cobalt and nickel from aluminum-coprecipitated products in ammoniacal solutions has been investigated. Cobalt and nickel were precipitated with aluminum in a solution containing 0.5 M of total ammonia, a mixture of ammonium hydroxide and ammonium nitrate at pH 10. The leaching was carried out at 2 M total ammonia concentration at pH 10. The effect of aging the coprecipitated products on the leaching behavior of cobalt and nickel was also investigated. The equation of continuity with a source term for a spherical coordinate was solved numerically using a forward finite differentiation technique to describe the leaching of cobalt. The nickel leaching kinetics was best described by chemical reaction at the solid/liquid interface of the grains of the precipitated product. The Arrhenius activation energy for cobalt and nickel was found to be very much influenced by the aging time of coprecipitated products and varying over a wide range, 10.5 to 26.8 kJ/mol for cobalt and 15.0 to 38.9 kJ/mole for nickel.     

The extraction of nickel from ammoniacal media and its separation from copper,cobalt and zinc using hydroxyoxime extractants I. SME-529

Nickel extraction from ammoniacal media, containing, initially, 0.03 mol dm^?3 of nickel, by 10% v/v Shell metal extractant 529 in MSB 210 or Shellsol A, has been studied as a function of pH, counter anions (nitrate or sulphate), ammonium salt concentration, temperature (10 to 50°C), and the presence of copper, cobalt and zinc. The extraction of copper as a function of pH was studied in the presence of nickel. The pH for maximum nickel extraction decreased from 8.2 to 7.4 as the ammonium sulphate concentration increased from 0.1 to 3.4 mol dm^?3. Nickel extraction was lower from sulphate than from nitrate media and with Shellsol A, and at higher temperature. The enthalpy and entropy of extraction were negative. The presence of copper, cobalt and zinc decreased nickel extraction. Separation factors between copper and nickel are reported as a function of pH and indicate that initial removal of copper is desirable.From a consideration of the equilibrium, the extraction of nickel (with MSB 210 diluent) at 25°C could be expressed by the equation:

$\text{log}\text{D}{}_{\mathrm{<mtext>Ni</mtext>}}\text{= ?9.25 +}\text{log}\text{α}{}_0\text{+ 2}\text{log}\text{(C}{}_{\mathrm{HR}}\text{?2[}\text{Ni}\text{]}{}_{\mathrm{<mtext>org</mtext>}}\text{) + 2}\text{pH}$

where D_Ni represents nickel distribution coefficient, C_HR the total extractant concentration, [Ni]_org the organic nickel concentration (both in mol dm^?3) and α₀, the fraction of uncomplexed nickel in the aqueous phase calculated from the overall stability constants of Ni(NH₃)_i²⁺ complexes and the formation constant of NiSO₄ ion pairs.The limiting slope of ?1 for long D_Ni versus log [NH₃]_free plots indicated an average co-ordination number of 3 for nickel by ammonia in the aqueous phase.The uptake of SO₄^2?, NO₃^? and water by the organic phase was negligible. The uptake of ammonia was dependent upon pH and the ammonia content of the aqueous phase, and was negligible in the absence of metal.     

Synergic extraction of nickel and cobalt by LIX63-dinonylnaphthalene sulfonic acid mixtures

The extraction of nickel and cobalt by mixtures of 5,8-diethyl-7-hydroxy-6-dodecanone oxime (H₂Ox, active extractant in LIX63, a product of Henkel Corp.) and dinonylnaphthalene sulfonic acid (HDNNS or HD) has been studied by using the methods of slope analysis and continuous variation. On the basis of distribution and spectrophotometric data, the observed synergic extraction of nickel and cobalt by H₂Ox-HDNNS mixtures is attributed to the formation of a mixed ligand complex with the stoichiometry: Me/H₂Ox/HDNNS = 1/3/2. In the absence of H₂Ox, log D versus log [HDNNS] plots give straight lines with slopes of unity for both nickel and cobalt extraction, thus indicating that the active extractant is the HDNNS micelle. In the presence of H₂Ox the log D versus log [HDNNS] plots go through a maximum and a minimum. It is proposed that the active extractant under these conditions is a mixed HDNNS-H₂Ox micelle of the form (HD)_p·3H₂Ox, and that the shape of the metal distribution curve, as a function of HDNNS concentration, follows the organic-phase concentration of this mixed micelle.