Oxygen pressure leaching of Fe-Ni-Cu sulfide concentrates at 110°C — effect of low chloride addition

Nickeliferous sulfide concentrates containing pyrrhotite, pentlandite, and chalcopyrite were leached in aqueous media at 110°C under an oxygen overpressure. Addition of low concentrations of chloride (0.5–10 g/l) did not significantly affect the dissolution of pyrrhotite and pentlandite; however, the copper extraction from chalcopyrite was enhanced. The presence of the chloride ion appeared to maximize the oxidation to elemental sulfur, minimize sulfate formation, and reduce the amount of iron in solution. Minor improvement in the handling characteristics of the leach residue was also noticed. These advantages must be weighed against increased problems in the materials of construction and with plant effluents.     

Kinetics of Dissolution and Equilibrium in Solution of Complex Ores

The development of the kinetic expression for the dissolution of a nickeliferous sulfide by ferric sulfate and of pyrite by oxygen and ferric sulfate on the basis of its oxidation mechanism is discussed. The rate of dissolution of complex ores is determined not only by the kinetics of the heterogenous reactions (solid-liquid reactions) but also by the homogenous reactions taking place simultaneously in the leach liquor. The liquid phase in a leaching system contains a large number of species which are simultaneously reacting with one another such as in acid-base and complexation reactions at near equilibrium or at equilibrium. The heterogenous mineral solid-leaching solution reactions are limited by the kinetics of dissolution and often do not reach equilibrium. During the leaching process, the concentrations of chemical species in the liquid phase adjust rapidly to the changes in the liquid phase analytical concentrations. Chemical equilibrium is maintained in the homogenous phase although the mineral-leach solution reactions are far from equilibrium. If the reactions of the solid with the leach solution is considered to take place in very small increments, the pertubations to the analytical concentrations within the liquid phase can adjust quickly and thus remain at equilibrium as the leaching reaction proceeds. Following the changes in the liquid phase concentrations during the leaching step can be useful in optimizing the dissolution process and specifying the influent conditions to achieve optimum conditions. It may also be possible to specify the conditions under which selective leaching can be attained. The partial equilibrium model is capable of characterizing the minerals dissolution reactions and the associated changes in solution species concentrations. The modeling of the dissolution of chalcopyrite by ferric sulfate, ilmenite by hydrochloric acid, uranium dioxide-pyrite by ferric sulfate-sulfuric acid, and uraninite by ammonmium carbonate-hydrogen peroxide systems are discussed.     

Hydrometallurgical Process for Copper Recovery from Waste Printed Circuit Boards (PCBs)

Present paper focuses on the selective recovery of copper from the enriched ground printed circuit boards (PCBs) using leaching and solvent extraction. The metal-enriched ground sample obtained from the beneficiation of the sized PCBs in a laboratory scale column type air separator contained mainly 49.3% Cu, 3.83% Fe, 1.51% Ni, 5.45% Sn, 4.71% Pb, and 1.85% Zn. The leaching of the enriched sample with 3.5 mol/L nitric acid dissolved 99% copper along with other metals at 323 K temperature and 120 g/L pulp density in 1 h time. The composition of the leach liquor with wash solution was found to be 42.11 g/L Cu, 2.12 g/L Fe, 4.02 g/L Pb, 1.58 g/L Zn, and 0.4 g/L Ni. The McCabe–Thiele plot indicated the requirements of three counter-current stages for maximum extraction of copper from the leach liquor at pH 1.5 using 30, 40, and 50% (v/v) LIX 984 N at the phase ratios (A/O) of 1:3, 1:2, and 1:1.5, respectively. The counter-current simulation studies show the selective extraction of 99.7% copper from the leach liquor feed of 1.5 pH in three stages with 50% LIX 984 N at A/O phase ratio of 1:1.5. The stripping of copper from the loaded organic with sulfuric acid produced copper sulfate solution from which copper metal/powder could be recovered by electrolysis/ hydrogen reduction.     

Role of ferric ions in bioleaching of uranium from low tenor Indian ore

Abstract

The role of biogenic ferric ions in leaching of uranium by Acidithiobacillus ferrooxidans from a low grade ore of Turamdih mines, India, has been investigated. Using the enriched culture of bacterial isolate containing mainly A. ferrooxidans from the source mine water, biorecovery of 98% uranium at 20% (w/v) pulp density, pH 1·7 and 35°C temperature using <76 μm particles in 40 days was obtained. The effect of temperature on bioleaching of uranium showed higher recovery at 35°C. The uranium dissolution was facilitated by iron(III) available in the leach liquor because of bacterial oxidation of pyrite and chemical dissolution of magnetite present in the ore under acidic conditions. The biogenically generated Fe(III) ions enhanced uranium dissolution from the uraninite ore. The bioleaching of uranium appeared to follow a chemical control kinetic model with the reaction of lixiviant, Fe(III) and acid on the surface of the solid in the temperature range 25–35°C. Phase identification by XRD and the study of surface morphology of the ore and the residue by SEM study corroborated the above mechanism of uranium leaching.

On a étudié le rôle des ions ferriques biogènes dans la lixiviation de l’uranium par A. ferrooxidans d’un minerai pauvre des mines de Turamdih, en Inde. En utilisant la culture enrichie d’isolat bactérien contenant principalement Acidithiobacillus ferrooxidans de l’eau de source de la mine, on a obtenu une biorécupération de 98% d’uranium à 20% de densité de la pulpe (poids/vol), avec un pH de 1·7 et une température de 35°C, en utilisant des particules de <76 μm, en 40 jours. L’effet de la température sur la biolixiviation de l’uranium montrait une récupération plus élevée à 35°C. La dissolution de l’uranium était facilitée par le fer(III), disponible dans la liqueur de lixiviat grâce à l’oxydation bactérienne de la pyrite et à la dissolution chimique de la magnétite présentes dans le minerai en conditions acides. Les ions Fe(III) engendrés biogéniquement augmentaient la dissolution de l’uranium du minerai d’uraninite. La biolixiviation de l’uranium semblait suivre un modèle cinétique à contrôle chimique avec la réaction du lixiviant – Fe(III) et de l’acide à la surface du solide dans la gamme de température de 25 à 35°C. L’identification de phase par XRD et l’étude de la morphologie de la surface du minerai et du résidu par étude au SEM corroboraient le mécanisme ci-dessus de lixiviation de l’uranium.     

Cyanide and Other Lixiviant Leaching Systems for Gold with Some Practical Applications

Selection of a leaching system for gold involves consideration of ore texture and mineralogy, chemical requirements, leaching techniques, the development of flowsheets, and environmental management. Aqueous dissolution chemistry for alkaline, neutral, and acid systems is mainly considered here. All systems require an oxidant to oxidise gold and a ligand to complex with gold in solution. Adjustment of pH is usually necessary.

Alkaline lixiviant systems (pH > 10)include cyanide, ammonia-cyanide, ammonia, sulphide, nitriles, and a few other minor possibilities. Oxygen is the main oxidant. Cyanide, which is the main ligand in these systems, forms an anionic complex, “Au(CN)2”, with Au(I). Gold dissolution rates are controlled by oxygen solubility in solution.

Neutral lixiviant systems (pH 5-9)include thiosulphate, halogens, sulphurous acid, and bacteria plus natural organic acids as the ligand. Oxygen is the normal oxidant and either Au(I) or Au(III)complexes are formed.

Acid leaching systems (pH ? 3)may contain thiourea, thiocyanate, chlorine, aqua regia, or ferric chloride. Chloride is the ligand in the last three systems and the oxidants include chlorine, ferric chloride, hydrogen peroxide, and nitric acid which produce Au(III) anionic complexes, e.g. [AuClJ". Fast gold dissolution is possible but reagent consumptions are high. Thiourea is unusual in producing a cationic Au(I)complex, “Au(NH2CSNH2)2” and gold dissolution is slower.

For treating simple auriferous oxide-silicate-carbonate ores, and many otfier materials, cyanide remains the preferred lixiviant.

Most non-cyanide leaching systems appear to have little wide-spread practical application. Possible niche applications include the use of chlorine or aqua regia to dissolve coarse gold from gravity concentrates, oxidising acid chloride solutions for die treatment of auriferous base metal sulphide concentrates, thiosulphate for dissolving gold from gold-copper ores, and thiourea for auriferous hydrometallurgical intermediates.     

Leaching behavior of metals from a limonitic nickel laterite using a sulfation–roasting–leaching process

The leaching behavior of metals from a limonitic laterite was investigated using a sulfation–roasting–leaching process for the recovery of nickel and cobalt. The ore was mixed with water and concentrated sulfuric acid followed by roasting and finally leaching with water. Various parameters were studied including the amount of acid added, roasting temperature and time, sample particle size, addition of Na₂SO₄ and solid/liquid ratio in leaching process. More than 88% Ni, 93% Co and < 4% Fe are extracted under the determined conditions. Simultaneously, about 90% Mn and Cu, 70% Mg, 45% Al, 25% Zn, 4% Cr and Ca are extracted respectively. The pH of the leach solution is about 2. The leaching efficiency is independent of sample particle size due to decomposition of ferric sulfate formed during roasting. The roasted mass was characterized by various physico-chemical techniques such as DSC/TGA, XRD and SEM. This process provides a simple and effective way for the extraction of nickel and cobalt from laterite ore.     

The Dissolution of Sphalerite in Ferric Chloride Solutions

The kinetics of dissolution of both sintered sphalerite disks and untreated sphalerite particles in ferric chloride-hydrochloric acid solutions have been investigated. Over the temperature interval 25 to 100°C, the dissolution occurred according to a linear rate law and with an associated apparent activation energy of about 10 kcal/mole. Most of the oxidized sulfide ion reported as elemental sulfur in the leach residues. The leaching rate was independent of the disk rotation speed and this fact, together with various hydrodynamic calculations, indicated that the reaction was chemically controlled. The dissolution rate increased as the 0.36 power of the ferric chloride concentration and it also increased substantially in the presence of dissolved CuCl₂. The accumulation of the ferrous chloride reaction product severely retarded the leaching reaction, but the presence of dissolved zinc chloride only slightly impeded it. The leaching rate was relatively insensitive to low levels of HC1 (>1 M), but increased dramatically at higher acid concentrations because of direct acid attack of the ZnS.     

生物浸出低品位镍铜硫化矿

阐述了氧化亚铁硫杆菌 (TF5)和氧化硫硫杆菌 (TT)浸出金川低品位镍铜硫化矿的机理、过程动力学、工艺条件和反应工程。研究表明 ,含镍磁黄铁矿的细菌浸出以细菌氧化生成的Fe3 +的作用为主 ,浸出速率受表面反应控制 ;镍黄铁矿的细菌浸出以矿物表面吸附菌的作用为主。细菌对Mg2 +离子的耐受浓度因驯化而提高 ,极限浓度可达 15～ 2 0g/L。低品位镍铜矿的细菌浸出过程中 ,pH控制、细菌的初始接种量、矿浆浓度及TF和TT的混合比是影响镍、铜、钴等有价金属元素浸出速率和最终浸出率的主要因素。优化条件下气升式和搅拌式反应器中试验表明 ,优化条件下 ,在生物浸出低品位镍铜硫化矿 ,镍浸出率可达到 92 %～ 94 % ,铜达 4 8%～ 50 % ,钴达 88%～ 91%。     

Hydrochloric acid leaching of nickel sulfide precipitates

Nickel sulfide precipitates produced by the AMAX Acid Leach Process for oxide nickel ores were leached in hydrochloric acid. The effects of process variables such as temperature, acid concentration, stoichiometric excess of HCl, gas sparging and heat treatment of feed were investigated. The nickel leachability was found to be in the 60–80% range. Chemical and mineralogical examination of the leach residues indicated the presence of NiS₂. This higher nickel sulfide is insoluble in hydrochloric acid, and its presence hinders the leaching of NiS. Several methods are suggested to reduce the sulfur content in order to attain complete dissolution. The thermodynamics and kinetics of nickel sulfide leaching are briefly discussed.     

The extraction of copper(II) and iron(III) from chloride and sulphate solutions with LIX 64N in kerosene

A comparison has been made of the extraction of Cu(II) and Fe(III) from chloride and sulphate solutions with LIX 64N in kerosene. The effects on extraction of pH, anion concentration, and temperature were examined, and attention was paid to the ionic strength of the aqueous media, some of which contained aluminium and magnesium; extraction was carried out under ‘practical’ rather than ‘ideal’ conditions. Extraction of both Cu and Fe was enhanced from chloride solutions compared with sulphate; although separation of Cu from Fe was slightly reduced, extraction of Cu from chloride liquors appears to be applicable to commercial leach solutions.     

Fe(III)–As(V) precipitates obtained from a sulphate system and their leach stability

Fe(III)–As(V) precipitates were synthesised from the Fe(II)-As(V)-SO₄^2--H₂O system at a temperature of 90°C and a constant pH maintained at 1.5?±?0.05. The precipitates obtained were characterised by X-ray diffraction (XRD), chemical composition analysis, scanning electron microscope (SEM) and Raman spectrum, thermogravimetric analyzer (TGA) and electron probe micro-analyzer (EPMA). The precipitates were in irregular aggregation of about 1–4?μm in size. The precipitates consisted of scorodite, ferric arsenate and an amorphous ferric hydroxide sulphate formulated as Fe(OH)_x(SO₄)_y. The precipitates were stable in modified Toxicity Characteristic Leaching Procedure (TCLP) tests at pH 4.93 for 60?h. Arsenic concentrations in the leaching solutions of 0.27?mg?L^?1 and 0.59?mg?L^?1 were obtained for the precipitates prepared initial Fe(II)/As(V) molar ratios of 4.0 and 5.0, respectively. Significantly more iron than arsenic was dissolved with up to 280?mg?L^?1 of iron reporting to solution. Long-term stability tests of the precipitates were carried out by leaching them for 40 days at 25°C under various media of pH between 9.50 and 10.57. The results showed that the precipitates tested in this study were more stable than those by previous researchers owing to a preferential dissolution of the amorphous ferric hydroxide sulphate.     

Kinetics of leaching of a low-grade FeNiCuCo matte in ferric sulphate solution

A study of the dissolution of a low-grade FeNiCuCo matte in acid ferric sulphate solution is reported. This matte contained a number of different phases, and a detailed quantitative mineralogical analysis of the matte was used as a basis for mathematical modelling of the leaching process. The mathematical model assumed that the matte particles were leached by a shrinking-particle mechanism, and that the surface reaction was rate-limiting and electrochemical in nature. The experimental results verified these assumptions, and showed that the redox potential of the solution best approximated the potential of a matte particle during leaching.     

Dissolution and structural alteration of phlogopite mediated by proton attack and bacterial oxidation of ferrous iron

Microbiological weathering of a research-grade mica mineral, phlogopite, was studied using ferrous sulfate media that were inoculated with an acidophilic iron-oxidizing bacterium, Thiobacillus ferrooxidans. Weathering due to dissolution was monitored by analysis of Si, Al, Fe, K, Na, Mg, and Ca in the leach solutions and in chemical controls at pH 1.0, 1.5 and 2.0. Structural alterations of phlogopite were analyzed by X-ray diffraction. At pH 2, the oxidation of Fe(II) by T. ferrooxidans was accompanied by the formation of jarosite within 7 days of incubation at 22°C. The precipitation of jarosite was coupled with partial alteration of phlogopite to vermiculite and an interstratified (mixed-layer) phlogopite/vermiculite. Similar results were obtained with chemical controls containing 120 mM ferric sulfate. The data suggested that K incorporated into jarosite was released from interlayer positions in phlogopite; thus, jarosite constituted a sink for K. The formation of jarosite and expansible layer silicate phases was pH-dependent. At pH<1.5, jarosite was not formed and phlogopite weathering was due to chemical dissolution without detectable structural alteration.     

A partial equilibrium chemical model for the dump leaching of chalcopyrite

The chemistry of the dump leaching of chalcopyrite has been studied by means of a chemically based model that makes it possible to calculate concentration changes for solution species during mineral dissolution. In a dump, chalcopyrite can dissolve in two ways: CuFeS₂ + 4Fe³⁺ → Cu²⁺ + 5Fe²⁺ + 2S° CuFeS₂ + O₂ + 4H⁺ → Cu²⁺ + Fe²⁺ + 2S° + 2H₂O CuFeS₂ dissolution is not at equilibrium in a leach dump. However, there are a large number of homogeneous reactions taking place in the leach liquor. These may be treated as rapid equilibrium reactions in this “partial equilibrium” model. The model equations are linear and consist of an equation for each aqueous phase reaction, mass and charge balances, and a kinetic equation. Dissolution by O₂ and acid should be prevented if possible. The acid consumed is partly replaced by shifts in the solution equilibria. However, the net pH increase that occurs leads to precipitation of ferric ion. The total amount of copper dissolved without precipitation is highest if high Fe(III) concentrations are used and oxygen is excluded. High H₂SO₄ concentrations are beneficial and high FeSO₄ concentrations deleterious because of their influences on the precipitation equilibria. KNONA C. LIDDELL, formerly Graduate Assistant, Ames Laboratory USDOE and Department of Chemical Engineering, Iowa State University, Ames, IA 50011     

Mathematical modeling of the zinc pressure leach process

A comprehensive mathematical model is described for the zinc pressure leaching process. Generic kinetic expressions were derived from experimental data found in the literature for the following reaction events: (1) dissolution of marmatite, (Zn,Fe)S, (2) oxidation of ferrous to ferric, and (3) precipitation of lead jarosite. Aqueous solution properties, oxygen solubility, density, enthalpy, and vapor pressure, were correlated with solution composition and temperature. Subsequently, a kineticsbased model for simultaneous sulfide dissolution and iron precipitation in a multistage, three-phase reactor was developed. The population balance method was used for sulfide mineral material balances, and apparent equilibrium was assumed for iron precipitation. A gas-phase material balance was included, which allows for prediction of oxygen utilization. The model was solve for a particular operation of the Cominco Ltd. (Trail, BC) autoclave, and prediction results were shown to be in very good comparison with actual plant performance.     

Pressure acid leaching of a Chinese laterite ore containing mainly maghemite and magnetite

The high pressure acid extraction of nickel and cobalt from a Chinese laterite containing mainly maghemite and magnetite was studied. X-ray diffraction (XRD) and scanning electron microscopy/X-ray energy dispersive spectroscopy (SEM/EDS) were employed to characterize the residues. The factors influencing the dissolution of maghemite and magnetite, nickel and cobalt extractions and iron precipitation were investigated. The results show that after 75 min at 270 °C with an acid/ore ratio of 0.55, maghemite and magnetite completely dissolved, liberating 98% Ni and 88% Co into the leach liquor. EDS analysis reveals that some nickel may be associated with the amorphous silica and/or basic ferric sulfate, resulting in a minor loss of nickel. The presence of a cobalt-containing phase in the residues, believed to be ringwoodite, is mainly responsible for the incomplete extraction of cobalt. Both maghemite and magnetite dissolved gradually with the increase in temperature from 200 to 270 °C. Maghemite dissolved more slowly than magnetite at 270 °C which also produced ferrous sulfate in the leach liquor and increased the total iron extraction. Increasing temperature and/or agitation accelerated the hydrolysis of ferric sulfate. The leaching of maghemite and magnetite corresponds to a dissolution-precipitation mechanism. In both high and low acidic environments, the precipitation of ferric sulfate proceeds through the initial formation of basic ferric sulfate and its conversion to hematite. The extent of conversion depends largely upon residual acidity and reaction time.     

The effect of microwave irradiations on the leaching of zinc from bulk sulphide concentrates produced from Rampura–Agucha tailings

This work reports the effect of microwave irradiation on extraction of zinc from a bulk sphalerite (ZnS)/pyrrhotite (FeS) concentrate produced by flotation of tailings from the Rampura Agucha Lead–Zinc mines in India. This material could not be treated economically by conventional hydrometallurgy or pyrometallurgical methods due to low zinc concentration. Consequently, microwave assisted leaching was tested. Zinc leaching in sulphuric acid was rapid with > 90% dissolution after 6 min, further irradiation increased the zinc to 96% after 16 min. Iron dissolution increased up to 12 min irradiation and then decreased due to precipitation. Power consumption for > 90% Zn recovery was 0.36 kWh/kg concentrate. The leach solution contained 50–55 g/L zinc and 18–20 g/L iron which could be reduced to 0.54 g/L by jarosite precipitation. The ratio of Zn/Fe leached was 3.4 compared with the ratio of 0.6 for the concentrate showing significant selectivity for zinc over iron using this method.     

Effect of sulfides on gold dissolution in ammoniacal thiosulfate medium

To understand how various sulfide minerals affect the dissolution behavior of gold in the ammoniacal thiosulfate leaching system, an extensive study has been carried out on gold leaching in the presence of sulfides using pure gold plates. Special emphasis has been placed on gold leaching in association with sulfide dissolution, thiosulfate decomposition, and dissolved oxygen depletion in leach solutions. The results demonstrated that the leaching behavior of gold depended strongly on the solubilities of the sulfides, the thiosulfate decomposition, and the oxygen concentration in slurries. Gold dissolution was enhanced or diminished, depending on the sulfide types and the sulfide concentrations in slurries. An increase in the stirring speed accelerated the gold dissolution rates due to the improved mass transfer occurring in the gold leaching process and the increased dissolved oxygen content in leach solutions. The addition of sulfate in the sulfide slurries increased the gold leaching rates because of the depression of the sulfide dissolution in the leaching systems. Topological studies by scanning electron microscopy (SEM) demonstrated that the existence of passivating layers at the leached gold surfaces could result in the retardation of gold dissolution in the presence of sulfide minerals. An erratum to this article is available at .     

The recovery of nickel from laterites by chelate formation and reduction with hydrogen

The interaction between alkaline solutions of the sodium salts of ethylenediaminetetraacetic acid (E.D.T.A.) and nickel-bearing laterites has been studied at temperatures between 25°C and 90°C. Experiments were carried out with a serpentine ore containing 1.65% nickel, 6.10% iron, 20.2% magnesium and a limonite ore containing 1.51% nickel, 49.7% iron, 0.66% magnesium. Reduction with 400 p.s.i. or 800 p.s.i. hydrogen at 120°C or 140°C has been examined as a means of recovering nickel from the leach solution and regenerating the leachant.For both ores, an increase in pH from 8 to between 11 and 13, or a rise in temperature, gave increased nickel dissolution rates and improved selectivity for nickel dissolution over that of iron, manganese, magnesium, aluminum, chromium and calcium. After 48 hour leaches at 90°C and pH 13 with 1.5 moles E.D.T.A. per mole of nickel in the ore, nickel extractions from the serpentine and limonite were 87% and 27% respectively. Rates of nickel dissolution from both materials were markedly increased by precalcination at temperatures which gave partial decomposition of the mineral structures.Tests showed that reduction with hydrogen could be used to recover nickel from leach solutions at pH 13 but not at pH 12. Reduction from a leach solution at pH 13 with 800 p.s.i. hydrogen at 140°C for 3 hours gave 91% nickel recovery and a solution that was used effectively as recycle leachant.The substantial increase in nickel dissolution rates that resulted from precalcination suggests that leaching characteristics are strongly dependent on laterite mineralogy. With further information covering additional materials of differing mineralogy, the feasibility of any process based on E.D.T.A. leaching of raw laterite could be assessed more readily.