Kinetics of chloridization of nickel oxide with gaseous hydrogen chloride

The kinetics of chloridization of finely divided as well as granulated nickel oxide by gaseous HCl were studied in the temperature range 150 °C to 400 °C. The rate of chloridization depended upon temperature, partial pressure of HCl, and granule size. The conversion of NiO to NiCl₂ follows the logarithmic rate law or a pore-blocking model, which is attributed to the large increase in molar volume of the product phase. The blocked pores decreased the diffusion of gaseous reactant through the product layer. This observation has been supported by the low value of activation energy, the high order of reaction with respect to the partial pressure of HCl (g), and a slope value of −2 for the log-log plot of rate constant vs grain size. The morphological characteristics of both unreacted and chloridized nickel oxide have been studied with the help of a scanning electron microscope.     

Study of chloridizing volatilization roasting of cinnabar as a basis for a process to obtain mercuric and mercurous chlorides

This paper presents a study of the chloridizing volatilization reaction of concentrated cinnabar ore from Almadén, Spain in an air atmosphere in order to obtain mercuric and mercurous chlorides in a single step. The chloridizing agents used were ferric or aluminum chloride 6-hydrates. These chlorides decompose to HC1 at temperatures lower than required to break down cinnabar. According to the results obtained, aluminum chloride 6-hydrate does not seem to be suitable for the cinnabar chloridizing reaction. By the use of ferric chloride 6-hydrate, this reaction can be completed in 15 minutes, using a thermal cycle of 100 °C to 400 °C. The exhaust gases are partially neutralized by the roasting residue, reducing the SO₂ given off. The paper discusses the different behaviors of some metal chloride 6-hydrates used to study the chloridizing reaction of cinnabar ores. This reaction with hydrated metal chlorides may be used as the basis for a process to obtain mercuric chloride, mercurous chloride, or mercuric/ mercurous chloride mixtures.     

Upgrading Titanium Ore Through Selective Chlorination Using Calcium Chloride

To develop a simple and effective process for upgrading low-grade titanium ore (ilmenite, mainly FeTiO₃), a new selective chlorination process based on the use of calcium chloride (CaCl₂) as the chlorine source was investigated in this study. Titanium ore and a titanium ore/CaCl₂ mixture were placed in two separate crucibles inside a gas-tight quartz tube that was then positioned in a horizontal furnace. In the experiments, the titanium ore in the two crucibles reacted with either HCl produced from CaCl₂ or CaCl₂ itself at 1100 K (827 °C), leading to the selective removal of the iron present in the titanium ore as iron chlorides [FeCl _x (l,g) (x = 2, 3)]. Various kinds of titanium ores produced in different countries were used as feedstock, and the influence of the particle size and atmosphere on the selective chlorination was investigated. Under certain conditions, titanium dioxide (TiO₂) with purity of about 97 pct was directly obtained in a single step from titanium ore containing 51 pct TiO₂. Thus, selective chlorination is a feasible method for producing high purity titanium dioxide from low-grade titanium ore.     

Reducibility of laterite ores

The reducibility of several types of lateritic nickel bearing ores was investigated. The ores were reduced with hydrogen over a temperature range of 673 to 1273 K and reaction times from 5 to 80 minutes. The fraction of nickel, iron, and cobalt reduced to the metallic state was determined by leaching the reacted samples with a bromine-methanol solution followed by atomic absorption analysis for the individual elements. The reducibility of nickel increased with increasing iron concentration of the ore. Increased reduction temperature greatly raised the amount of nickel reduced for ores with high iron concentrations. The cobalt reducibility decreased with increasing iron concentration of the ore. Changes in reduction temperature affected cobalt reduction less than nickel reduction. The observed changes in reducibility have been attributed to the formation of phases which incorporate nickel and cobalt. The major ore components were plotted on the ternary phase diagram of the SiO₂+(Al₂O₃)-MgO-FeO system. It is shown how this plot can be used to predict the reducibility of different types of lateritic ores.     

The solubility of silver chloride in ferric chloride leaching media

The solubility of silver chloride in various FeCl₃-FeCl₂-HCl solutions has been measured over the temperature range 20–100°C; the densities of the associated saturated solutions have also been determined. The solubility increases systematically with either rising temperature or increasing concentrations of the constituent chlorides. The solubility is higher in a ferrous chloride medium than in an equivalent ferric chloride solution. The presence of CuCl₂ systematically raises the AgCl solubility, but increasing ZnCl₂ concentrations cause the solubility to decrease slightly to about 1.5 M ZnCl₂ and subsequently to increase gradually. The addition of NaCl to the iron chloride media substantially increases the AgCl solubility under all conditions. Although the solubility of AgCI is only a few mg/L in cold water, this increases to over 1 g/L in hot, moderately concentrated chloride media. Silver chloride solubilities at elevated temperatures are sufficiently high that silver solubility limitations should not be a problem in most commercial ferric chloride leaching processes.     

The dissolution of galena in ferric chloride media

The dissolution of galena (PbS) in ferric chloride-hydrochloric acid media has been investigated over the temperature range 28 to 95 °C and for alkali chloride concentrations from 0 to 4.0 M. Rapid parabolic kinetics were observed under all conditions, together with predominantly (>95 pet) elemental sulfur formation. The leaching rate decreased slightly with increasing FeCl₃ concentrations in the range 0.1 to 2.0 M, and was essentially independent of the concentration of the FeCl₂ reaction product. The rate was relatively insensitive to HCl concentrations <3.0 M, but increased systematically with increasing concentrations of alkali or alkaline earth chlorides. Most significantly, the leaching rate decreased sharply and linearly with increasing initial concentrations of PbCl₂ in the ferric chloride leaching media containing either 0.0 or 3.0 M NaCl. Although the apparent activation energy was in the range 40 to 45 kJ/mol (∼10 kcal/mol), this value was reduced to 16 kJ/mol (3.5 kcal/mol) when the influence of the solubility of lead chloride on the reaction rate was taken into consideration. The experimental results are consistent with rate control by the outward diffusion of the PbCl₂ reaction product through the solution trapped in pores in the constantly thickening elemental sulfur layer formed on the surface of the galena.     

Reaction mechanism on the smelting reduction of iron ore by solid carbon

The kinetics of the smelting reduction of iron ore by a graphite crucible and carbon-saturated molten iron was investigated between 1400 °C and 1550 °C, and its reaction phenomena were continuously observed in situ by X-ray fluoroscopy. In the smelting reduction by graphite, it was shown from the observation results that the smelting reduction reaction proceeded by the following two stages: an initial quiet reduction without foaming (stage I) and a following highly active reduction with severe foaming (stage II). At 1500 °C, by the graphite crucible, the reduction rate of iron ore was found to be 8.88×10⁻⁵ mol/cm² · s, and by the molten iron, 8.25×10⁻⁵ mol/cm²·s. The activation energies for the reduction by the graphite crucible and the molten iron were 24.1 and 22.9 kcal/mol, respectively. Based on the results of kinetic research and X-ray fluoroscopic observations, it can be concluded that these two types of smelting reduction reactions of iron ore by the graphite crucible and by the molten iron are essentially the same.     

Reducibility of laterite ores

The reducibility of several types of lateritic nickel bearing ores was investigated. The ores were reduced with hydrogen over a temperature range of 673 to 1273 K and reaction times from 5 to 80 minutes. The fraction of nickel, iron, and cobalt reduced to the metallic state was determined by leaching the reacted samples with a bromine-methanol solution followed by atomic absorption analysis for the individual elements. The reducibility of nickel increased with increasing iron concentration of the ore. Increased reduction temperature greatly raised the amount of nickel reduced for ores with high iron concentrations. The cobalt reducibility decreased with increasing iron concentration of the ore. Changes in reduction temperature affected cobalt reduction less than nickel reduction. The observed changes in reducibility have been attributed to the formation of phases which incorporate nickel and cobalt. The major ore components were plotted on the ternary phase diagram of the SiO₂+(Al₂O₃)-MgO-FeO system. It is shown how this plot can be used to predict the reducibility of different types of lateritic ores. Formerly with the Department of Metallurgy and Materials Science, University of Toronto, Toronto, ON, Canada     

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.     

La Lixiviation de concentres de sulfures de nickel par le chlorure ferrique

The effectiveness of ferric chloride leaching of nickel sulphide concentrates and of a nickel sulphide matte has been demonstrated. Ferric chloride concentrations ranged from 1 M to 5.7 M for treatment of the concentrates and 0.6 M to 3.8 M for that of the matte; experiments were carried out at 80 ± 1°C.The rate of the reaction was markedly slower after the dissolution of ～ 70% of the Ni. Concentrations of ferric chloride have a significant effect upon the quantity of elemental sulphur obtained. During leaching, a part of the iron from the leachant precipitated out. β-NiS is formed during the treatment of a nickel sulphide matte with ferric chloride.The reaction was followed by continuous measurements of pH, conductivity, redox potential and spectrophotometric intermittent measurements concerning nickel concentrations.     

Absorption of gaseous oxygen by liquid cobalt,copper, iron and nickel

An experimental investigation of the rates of oxygen solution in molten cobalt, copper, iron and nickel was carried out using pure oxygen and a constant-volume Sieverts’ method. It was found that the volume of gaseous oxygen which initially reacted with the inductively stirred metals was strongly dependent on the physical nature of the oxide film which formed during the first stage of reaction. The initial temperature of the molten iron, cobalt, and nickel was 1600°C, and for copper was 1250°C. For initial oxygen pressures above the melt of about one atmosphere both molten iron and copper, which formed liquid surface oxides, initially absorbed nearly 20 cm³ (STP) O₂/cm² of melt surface area, while molten cobalt and nickel, which formed solid oxides, absorbed about 6 cm³ (STP) O₂/cm² under the same experimental conditions. For approximately 30 s after the initial reaction between these liquid metals and gaseous oxygen, the oxygen absorption rate was proportional to the square root of the oxygen pressure above the melt, and proportional to the melt surface area, but independent of melt volume. The rate-limiting step for oxygen absorption by liquid iron, cobalt and copper can be described by dissociative adsorption of oxygen molecules at the gas/oxide interface. After 30 s of reaction, the rate of oxygen absorption became less dependent on the oxygen pressure above the melt. This indicated that the rate-controlling step was changing from a surface reaction to growth of the oxide layer by cationic diffusion in the bulk oxide. The oxidation rate of liquid nickel appears to be too complex to be described by models for dissociative adsorption of oxygen molecules at the gas/oxide interface and parabolic growth of the oxide layer. The formation of a thin layer of nickel oxide which allows oxygen to migrate through cracks or grain boundaries may be responsible for the relatively high oxygen absorption rate compared to that of liquid cobalt. R. H. RADZILOWSKI, formerly a Graduate Studient at The University of Michigan     

Equilibria between ferrous and ferric chlorides in molten chloride salts

Equilibria between ferrous and ferric chlorides in molten salts have been studied for improving magnesium electrolysis and molten salt chlorination. The apparent equilibrium constants,K, of reaction FeCl_2(melt)+0.5Cl_2(gas)=FeCl_3(melt) were obtained. Measured values ofK were in good agreement with computed ones from regression equations. The composition of the melts, the partial pressure of chlorine, and the temperature were found to have important effects onK, and the effect of dissolved iron was smaller than that of other factors. At identical other conditions, the largest values ofK were observed in system 3, which suggested that the current efficiency for electrolysis of MgCl₂ should be lower when carnallite was used as electrolyte and that catalysis of iron species in molten salt chlorination would be better when molten salt systems containing high potassium chloride were used.     

Chlorination of chalcopyrite concentrates

The chlorination behaviors of two chalcopyrite concentrates and their pure constituents in Cl₂+N₂ were investigated by thermogravimetric analysis (TGA) in nonisothermal conditions up to 1000 °C. The effect of temperature on the reaction of chlorine with both concentrates was studied between 170 °C and 300 °C under isothermal conditions. The effects of gas flow rate, chlorine content of the gas mixture, and reaction time on the reaction rate were also investigated. The reaction products were analyzed by scanning electron microscopy (SEM) and X-ray diffraction (XRD). Results showed that the kinetics of chlorination of chalcopyrite concentrates generating chlorides of Cu, Pb, Zn, Fe, and S was rapid at about 300 °C. The iron and sulfur chlorides were volatilized, leading to a residue containing valuable metal chlorides.     

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.     

Ferric chloride leaching of sulphidized chalcopyrite

Reactions between chalcopyrite and elemental sulphur at between 320°C and 400°C produce various combinations of covellite, nukundamite (Cu_5,5 FeS_6,5, and pyrite depending on the sulphur content of the products. If all the contained copper is transferred to covellite, and iron to pyrite, at least 95% of the copper can be leached by ferric chloride with negligible iron dissolution at up to 70°C. At above 70°C, iron is extracted from pyrite. When nukundamite is present, the iron and copper contents of this compound are leached simultaneously, and the consequent loss of selectivity is further enhanced at above 70°C by attack of pyrite. The increased rates of copper extraction obtained following sulphidation result from both the breakdown of chalcopyrite and a greater porosity.     

Extraction of nickel from ferromolybdenum leach residues

Abstract

The extraction of nickel from ferromolybdenum leach residues by sulphation roasting, water leaching and iron removal from subsequent nickel leach solutions was studied. Sulphation roasting and water leaching promote the reaction between sulphuric acid and the residue and decrease the silicon dissolution. Over 90% of Ni was leached. Ferric ions in the solution could be effectively removed as jarosite and ferric hydroxide. The recovery of nickel reached 88·3% under sulphation roasting with the sulphuric acid quality of 1472 kg t^?1 leach residue at 280°C for 4 h followed by iron removal with addition of 0·5 g NaClO₃, 6 g Na₂SO₄ and 10 g CaO/100 mL solution at 95°C for 2·5 h, while the concentration of iron in solution reduced to 0·38 from 56·6 g/L^?1.

On a étudié l’extraction du nickel à partir de résidu de lessivage de ferromolybdène par grillage sulfateur, lessivage à l’eau et enlèvement du fer des solutions obtenues de lessivage de nickel. Le grillage sulfateur et le lessivage à l’eau favorisent la réaction entre l’acide s ulfurique et le résidu et diminuent la dissolution de la silice. On a lessivé plus de 90% du Ni. On pouvait enlever efficacement de la solution les ions ferriques sous forme de jarosite et d’hydroxyde ferrique. La récupération du nickel atteignait 88·3% au moyen du grillage sulfateur, avec 1472 kg d’acide sulfurique par tonne de résidu de lessivage à 280°C pendant 4 h, suivi par l’enlèvement du fer avec l’addition de 0·5 g de NaClO₃, 6 g de Na₂SO₄ et 10 g de CaO par 100 mL de solution à 95°C pendant 2·5 h, la concentration du fer dans la solution étant réduite de 56·6 à 0·38 g L^?1.     

Kinetics of reduction of ilmenite with graphite at 1000 to 1100 °C

Pellets prepared from mixtures of synthetic ilmenite and graphite were heated in argon at 1000 to 1100 °C and the rate of reduction was followed gravimetrically. No measurable reaction was obtained at 1000 °C but the rate was high at 1100 °C. An inflection in the reduction curve at about 2 pct weight loss at 1050 and 1100 °C was due to the difficulty in nucleating iron. The rate is increased significantly by the addition of ferric chloride which promotes the nucleation of iron. The addition of rutile decreases the rate of reduction and the addition of manganese stabilizes a pseudobrookite phase.     

Surface tension measurements on pure liquid iron and nickel by an oscillating drop technique

A novel experimental technique for the measurement of surface tension of molten metals has been developed. It is based on the Rayleigh equation which relates frequency and surface tension for an oscillating drop. A systematic study has shown this equation to be valid for a liquid metal droplet levitated electromagnetically in an inert flowing gas with no prior calibration required. It is, therefore, an absolute method. The frequencies of oscillation of droplets of pure iron and nickel in a 6 pct H₂-He gas mixture were measured by high speed cinematography. Surface tensions were obtained for temperatures of 1550° to 1780°C for iron and 1475° to 1650°C for nickel. M. E. FRAZER, Formerly Graduate Student, Department of Metallurgy and Materials Science, McMaster University, Hamilton, Ontario, Canada.     

MgOHCl thermal decomposition kinetics

Experiments to determine the kinetics of the thermal decomposition of MgOHCl were performed. It was found that the decomposition of MgOHCl commenced at 649 K, and it directly converted into MgO and HCl without undergoing any intermediate step. Decomposition vs time data showed that the thermal decomposition of MgOHCl was a first-order process with respect to the amount of MgOHCl remaining, and the mass transfer of the product HCl gas away from the interface was likely the rate-limiting step. It was also found that the time required to completely decompose MgOHCl into MgO, more than 20 minutes, at the operating temperatures of electrolytic magnesium production processes, 600 °C±50 °C, was significantly longer than the time required, less than 1 minute, to digest the solid magnesium chloride containing feed material into the molten salt electrolyte in these processes. Such delay in the decomposition would mean that any MgOHCl produced during heating and digestion of the feed would not be decomposed by the heat of the electrolyte and thus the persistent MgOHCl would dissolve into the molten salt electrolyte with potentially severe negative consequences on electrolysis cell operation.