Investigation of kinetics of coal based reduction of oolitic iron ore

Abstract

An oolitic iron ore was isothermally reduced by coal at 1423–1573 K, and the reduction kinetics was investigated in detail. The degree of reduction and reduction rate increased with increasing temperature and C/O molar ratio to some extent at the same reduction time. In the entire reduction process, the reduction mechanism changes with changing experimental conditions. The degree of reduction under different experimental conditions should be represented by different reduction kinetic models. The reduction rate curves are similar in shape and could be analytically divided into initial, intermediate and final stages. The apparent activation energies of the three stages are 48·26, 69·80 and 127·58 kJ mol^?1 respectively. The rate controlling mechanism in the reduction process was determined by analysing the reduction process and apparent activation energy. The rate controlling steps of these stages are combined gas diffusion and interfacial chemical reaction, surface chemical reaction and combined solid state diffusion and boundary reaction.     

Isothermal reduction kinetics of self-reducing mixtures

Isothermal reduction of haematite carbon mixtures was investigated at temperatures 750–1100°C under inert atmosphere. Mass loss curves proved the stepwise reduction of haematite to metallic iron. The non-linear feature of haematite to magnetite reduction kinetics was observed and an activation energy of 209?kJ?mol^?1 was calculated. Irrespective of carbon-bearing material type, reduction rate of magnetite was linear. Activation energy values were calculated to be 293–418?kJ?mol^?1. Significant increase in the reduction kinetics in the last step (Wustite reduction) was observed and explained by the catalytic effect of freshly formed metallic iron. During the initial stages of wustite reduction, the activation energy values were calculated to be in the range of 251–335?kJ?mol^?1 for all carbon-bearing materials.     

Kinetics of solid state silica fluxed reduction of chromite with coal

Abstract

Solid state reduction of an Australian chromite with black coal and silica addition was studied thermogravimetrically in the temperature range 1000–1400°C under an argon atmosphere. Reduction was found to occur in two stages. In the first stage, reduction is initiated by the nucleation of metallic iron, and the rate is likely to be controlled by the diffusion of cations in the outer zone of the chromite particles. The Zhuravlev–Lesokhin–Tempel'man equation was determined to fit closest to the experimental data, giving an activation energy of 194 kJ mol^-1. The second stage involves the reduction of chromium, iron, and some silicon ions through the slag. The rate of reduction is proposed to be controlled by dissolution of the chromite into the slag with an activation energy of 256 kJ mol^-1. Silica addition was found to enhance significantly the rate and degree of reduction at 1300 and 1400°C.     

Leaching chalcopyrite with sodium chlorate in hydrochloric acid solution

Abstract

Sodium chlorate was used as an oxidant for the chalcopyrite leaching in a hydrochloric acid solution. The hydrochloric acid concentration has an important effect on the dissolution of chalcopyrite. The results indicate that stirring speed has a negligible effect on copper dissolution, suggesting that the reaction is not controlled by liquid phase diffusion. Chalcopyrite leaching is also affected by temperature. X-ray diffraction studies show that sulphide is oxidised to sulphate at temperatures less than or equal to 65°C, and at 85°C, it is oxidised to elemental sulphur during the leaching process. Studies of the effect of temperature on dissolution indicate that the leaching process is highly dependent on temperature for the range of 25–65°C. This result is consistent with the values for the activation energy at 25–45°C (60·0 kJ mol^?1) and 45–65°C (57·7 kJ mol^?1). Within these temperature ranges, the leaching process is controlled by a chemical reaction. However, at temperatures of 65–85°C, the activation energy in (0–180 min) is 28·17 kJ mol^?1, which suggests that the reaction is diffusion and chemically controlled during this stage. During the last stage (180–300 min) of the process at 65–85°C, the activation energy is only 0·55 kJ mol^?1. At this point, it appears that diffusion predominates. Scanning electron microscopy and energy dispersive X-ray spectroscopy results are consistent with the abovementioned kinetic data.

On a utilisé du chlorate de sodium comme oxydant pour le lessivage de la chalcopyrite dans une solution d’acide chlorhydrique. La concentration de l’acide chlorhydrique a un effet important sur la dissolution de la chalcopyrite. Les résultats indiquent que la vitesse d’agitation a un effet négligeable sur la dissolution du cuivre, suggérant que la réaction ne serait pas contrôlée par la diffusion en phase liquide. Le lessivage de la chalcopyrite est également affecté par la température. Des études de diffraction des rayons X ont montré que le sulfure était oxydé en sulfate à des températures égales à ou à moins de 65°C, et à 85°C, il était oxydé en soufre élémentaire lors du procédé de lessivage. Les études de l’effet de la température sur la dissolution indiquent que le procédé de lessivage est hautement dépendant de la température dans la gamme de 25 à 65°C. Ce résultat est compatible avec les valeurs de l’énergie d’activation à 25–45°C (60·0 kJ mol^?1) et à 45–65°C (57·7 kJ mol^?1). Dans ces gammes de température, le procédé de lessivage est contrôlé par une réaction chimique. Cependant, à des températures de 65 à 85°C l’énergie d’activation au début (0–180 min) est de 28·17 kJ mol^?1, ce qui suggère que la réaction est contrôlée par diffusion et chimiquement lors de cette étape. Lors de la dernière étape (180–300 min) du procédé à 65–85°C, l’énergie d’activation est seulement de 0·55 kJ mol^?1. À ce point, il semble que la diffusion domine. Les résultats de microscopie électronique à balayage et de spectroscopie à dispersion d’énergie sont compatibles avec les données cinétiques mentionnées ci-haut.     

ASSESSMENT OF REDUCTION BEHAVIOR OF HEMATITE IRON ORE PELLETS IN COAL FINES FOR APPLICATION IN SPONGE IRONMAKING

Studies on isothermal reduction kinetics (with F grade coal) in fired pellets of hematite iron ores, procured from four different mines of Orissa, were carried out in the temperature range of 850–1000°C to provide information for the Indian sponge iron plants. The rate of reduction in all the fired iron ore pellets increased markedly with a rise of temperature up to 950°C, and thereafter it decreased at 1000°C. The rate was more intense in the first 30 minutes. All iron ores exhibited almost complete reduction in their pellets at temperatures of 900 and 950°C in < 2 hours' heating time duration, and the final product morphologies consisted of prominent cracks. The kinetic model equation 1 ? (1 ? α)^1/3 = kt was found to fit best to the experimental data, and the values of apparent activation energy were evaluated. Reductions of D. R. Pattnaik and M. G. Mohanty iron ore pellets were characterized by higher activation energies (183 and 150 kJ mol^?1), indicating carbon gasification reaction to be the rate-controlling step. The results established lower values of activation energy (83 and 84 kJ mol^?1) for the reduction of G. M. OMC Ltd. and Sakaruddin iron ore pellets, proposing their overall rates to be controlled by indirect reduction reactions.     

Oxidation Kinetics of Vanadium Slag Roasting in the Presence of Calcium Oxide

The isothermal and non-isothermal oxidation kinetics of a converter vanadium slag in the presence of calcium oxide was studied using thermal analysis. The isothermal experimental data for the whole oxidation process are described in terms of the equation [1? (1?α)^2/3] = kt with E_a = 20.42 kJ mol^–1 at lower temperatures of 400-500 °C, and described by [(1?α)^–1/3?1]² = kt with E_a = 227.66 kJ mol^–1 at temperature higher than 500 °C. In the nonisothermal oxidation study, heating rate greatly affects the oxidation process. Using a heating rate of 3 °C min^–1 results in overlapping oxidations of vanadium spinel and augite over temperature range of 608-959 °C, which is described by the 3/2 order reaction. Increasing the heating rate to 5 °C min^–1 or 10 °C min^–1, only oxidation of vanadium spinel takes place in temperature range of 657-914 °C and 691-954 °C respectively, both described by the third order chemical reaction. As the slag particle decreases from 250 µm to 48 µm, the kinetic equation for describing the overlapping oxidation process changes from the Anti–Zhuravlev equation with internal diffusion controlling to reaction limiting equations.     

Enthalpies of formation of liquid binary (copper + iron, cobalt, and nickel) alloys

The enthalpies of formation of liquid binary (Cu+Fe, Co, Ni) alloys are studied by direct reaction calorimetry in the whole range of compositions at 1873, 1823, and 1753 K, respectively. The integral molar enthalpies of mixing are found to be positive in all three systems with the maximum values approaching 10.8±0.7 kJ/mol⁻¹ at x _Fe=0.43, 7.1±0.9 kJ/mol⁻¹ at x _Co=0.55, and 3.7±0.5 kJ/mol⁻¹ at x _N1=0.53. Partial molar enthalpies at infinite dilution constitute 59.4±3.3 kJ/mol⁻¹ for iron, 44.3±4.1 kJ/mol⁻¹ for cobalt, and 14.9±2.2 kJ/mol⁻¹ for nickel in liquid copper. Similar values for copper in liquid iron, cobalt, and nickel are 36.6±3.9, 45.3±6.0, and 17.7±4.4 kJ/mol⁻¹, respectively. The results are compared with the thermodynamic data available in literature and discussed in connection to the equilibrium-phase diagrams. In particular, decreasing from Cu-Fe to Cu-Ni liquid alloys positive values of the excess thermodynamic functions of mixing are fully in accord with the growing stability of phases in these systems. The excess entropies of mixing are estimated by combining the established enthalpies with carefully selected literature data for the excess Gibbs functions. Analysis of possible contributions to the enthalpies of mixing indicates that the experimentally established regularity in ΔH values along the 3d series is likely to arise from the difference in d-band width and d-electron binding energy of the alloy constituents.     

Phase Relationship of CaO-SiO2-FeO-5 mass pct P2O5 System with Low Oxygen Partial Pressure at 1673?K (1400?°C)

Dephosphorization by using multiphase flux could considerably decrease the consumption of CaO and prevent the addition of fluorite. However, the equilibrium phase relationship within this system, which is of significant importance for understanding the formation mechanism of multiphase flux, remains unclear. Thus, it is required to provide reliable phase diagrams of the basic slag system of multiphase flux. In this research, the phase relationship of the CaO-SiO₂-FeO-5 mass pct P₂O₅ system at 1673?K (1400?°C) with $ {P}_{{{\text{O}}_{2} }} $ of 9.24?×?10^?11 atm has been studied by using the chemical equilibration method. It has been found that solid solution consists mainly of 2CaO·SiO₂-3CaO·P₂O₅, but occasionally it contains 3CaO·SiO₂. Liquidus saturated with solid solution shrinks toward the FeO corner compared with the isothermal at 1673?K (1400?°C) of the CaO-SiO₂-FeO system equilibrated with metallic iron. Thermodynamically stable CaO-FeO phase is confirmed, which could promote the condensation of 3CaO·P₂O₅ into the solid solution and increase the phosphorus partition ratio between the solid solution and molten slag. Based on the regular solution model, the effect of T.Fe and CaO content in the liquid phase on the phosphorus partition ratio between the solid solution and molten slag is discussed.     

Combustion Property and Kinetic Modeling of Pulverized Coal Based on Non-isothermal Thermogravimetric Analysis

Non-isothermal combustion kinetics of two kinds of low volatile pulverized coals （HL coal and RU coal） were investigated by thermogravimetrie analysis. The results show that the combustibility of HL coal was better than that of RU coal, and with increasing heating rate, ignition and burnout characteristics of pulverized coal were improved. The volume model （VM）, the random pore model （RPM）, and the new model （NEWM） in which the whole combustion process is considered to be the overlapping process of volatile combustion and coal char combustion, were used to fit with the experimental data. The comparison of these three fitted results indicated that the combustion process of coal could be simulated by the NEWM with highest precision. When calculated by the NEWM, the activation energies of volatile combustion and coal char combustion are 130.5 and 95.7 kJ · mol^-1 for HL coal, respectively, while they are 114.5 and 147.6 kJ ·mol^-1 for RU coal, respectively.     

Dislocation creep controlled superplasticity in the high carbon steel 140NiCr16-6

Superplasticity in the alloyed high carbon-steel 140NiCr16-6 with phosphorus additions and a fine grained microdupiex structure – containing cementite in volume fractions of 22 % (Fe,Cr,Ni)₃C, particle size of about 1 μm and with a medium ferrite grain size of about 2 μm – has been investigated in the temperature regime of 550 to 675°C and in the strain rate range of 10^?5 to 5 · 10^?2 s^?1. Maximum strain rate exponents of m = 0,45 at 675°C with strain rates of the order of 10^?4 s^?1 have been determined. Maximum superplastic elongations of about 700 % were detected. At higher strain rates of 10^?3 s^?1 superplastic elongations of about 570 % were achieved. At relatively low test temperatures of 550°C elongations up to 230 % were recorded. The activation analysis in the temperature regime of 550 to 650°C show an activation energy for superplastic flow of 250 ± 20 kJ/mol. This is in agreement with the activation energy for lattice self diffusion of iron in α-iron. Above 650°C the activation energy decreases to 70 kJ/mol. This is due to a stress induced decrease in the eutectoid α-γ-transformation temperature from 685°C to somewhat lower temperatures during superplastic deformation. The superplastic deformability (m > 0.3) of this steel in a wide strain rate range at relatively low temperatures above 550°C allows near net shape forming of complex parts applying low flow stresses.     

Kinetics of Alkaline Decomposition and Cyaniding of Argentian Rubidium Jarosite in NaOH Medium

The alkaline decomposition of Argentian rubidium jarosite in NaOH media is characterized by an induction period and a progressive conversion period in which the sulfate and rubidium ions pass to the solution, leaving an amorphous iron hydroxide residue. The process is chemically controlled and the order of reaction with respect to hydroxide concentration in the range of 1.75 and 20.4?mol OH^? m^?3 is 0.94, while activation energy in the range of temperatures of 298?K to 328?K (25?°C to 55?°C) is 91.3?kJ mol^?1. Cyaniding of Argentian rubidium jarosite in NaOH media presents a reaction order of 0 with respect to NaCN concentration (in the range of 5 to 41?mol m^?3) and an order of reaction of 0.62 with respect to hydroxide concentration, in the range of 1.1 and 30?mol [OH^?] m^?3. In this case, the cyaniding process can be described, as in other jarosites, as the following two-step process: (1) a step (slow) of alkaline decomposition that controls the overall process followed by (2) a fast step of silver complexation. The activation energy during cyaniding in the range of temperatures of 298?K to 333?K (25?°C to 60?°C) is 43.5?kJ mol^?1, which is characteristic of a process controlled by chemical reaction. These results are quite similar to that observed for several synthetic jarosites and that precipitated in a zinc hydrometallurgical plant (Industrial Minera México, San Luis Potosi).     

Kinetics of Na2O evaporation from CaO–Al2O3–SiO2–MgO–TiO2–Na2O slags

The present work is carried out to study the evaporation of Na₂O from CaO–Al₂O₃–SiO₂–TiO₂–MgO–Na₂O slags with high basicity and high alumina in the temperature range of 1500–1560°C. The ratio of evaporation was determined by monitoring the Na₂O content change of the slag melt under isothermal reduction conditions. The results show that the evaporation ratio increases with increasing the temperature. Higher basicity and increasing concentrations of Na₂O, Al₂O₃ are also found to increase the evaporation ratio of Na₂O, while MgO addition only slightly enhances the evaporation ratio. With TiO₂ content increasing, the evaporation ratio first increases and then decreases. The evaporation rate of Na₂O appears to be controlled by chemical reaction at the slag/gas interface in the beginning, followed by a mixed reaction-mass transfer regime, and finally a liquid-phase mass transport step. The apparent activation energy is 134.74?kJ?mol^?1 for the chemical reaction regime and 268.53?kJ?mol^?1 for the liquid-phase mass diffusion step.     

Cupric chloride leaching of a complex copper/zinc/lead ore

A complex Cu/Zn/Pb ore from Cayeli, Turkey, was reacted with cupric chloride solutions under different conditions. Energies of activation were calculated for dissolution of copper (37 kJ mol^?1), iron (33 kJ mol^?1, zinc (26 kJ mol^?1) and lead (7.5 kJ mol^?1, values which indicate diffusion control of the reaction, probably through the sulphur layer formed round each particle. Particle size/leaching relationships corroborated microscopic assessments and indicated that chalcopyrite dissolved at a very low rate. Calculation of Fe:Cu ratios of metal leached showed considerable dissolution of pyrite from finely-ground ($\text{d}{}_{\mathrm{<mtext>1</mtext><mtext>2</mtext>}}\text{≈ 3?5 μ}\text{m}$) ore. Examination of residues using SEM X-ray fluorescence line scan techniques revealed little attack of large pyrite crystals, suggesting that fine pyrite particles in complex relationship with the sphalerite and chalcopyrite were dissolving.     

Benefits of oxygen enrichment in sponge iron rotary kiln emerging from data-driven modelling

Optimal oxygen enrichment conditions for sponge iron rotary kiln have been successfully explored on an industrial scale using a data-driven model. A multi-objective optimisation by genetic algorithm (MOGA) is employed to find the favourable conditions. The objective function for MOGA is derived from neural networks using pre-processed operational data. From industrial experimentations guided by the optimum conditions predicted by the present model, it emerged that when the coal fines injection is maintained at 1.75?tph and the oxygen enrichment is 8 Nm³?t^?1 of sponge iron, a reduction in the specific air requirement from 2609 to 2150?Nm³?t^?1 was obtained, while the end-zone bed temperature remained under control at 1132°C. These conditions resulted in a reduction of specific coal consumption by 6%, an enhancement in the sponge iron production by 6% and an increase in the rotary kiln campaign life from 50 to 100 days.     

Volatilisation behaviour of iron,silicon and magnesium during vacuum carbothermal reduction of ilmenite concentrate

The non-isothermal reduction kinetics and volatilisation behaviour of iron (Fe), silicon (Si), and magnesium (Mg) during the vacuum carbothermal reduction of ilmenite concentrate were investigated from 1350°C to 1550°C. Mg species was reduced to metallic Mg before volatilisation. Some Si elements were reduced to Si, which entered into the Fe phase to form ferrosilicon alloy, and others were reduced to SiO before volatilisation. The volatilisation of Fe came from the metallic Fe phase. For the comprehensive consideration of ?atava–?esták and Coats–Redfern methods, the rate-limiting factors of Fe, Mg, and Si were diffusion control. The integral mechanism functions of Fe, Mg, and Si were G(α)?=?[1?(1?α)^1/3]², G(α)?=?1?2α/3?(1?α)^2/3, and G(α)?=?(1?α)ln(1?α)?+?α, respectively. The apparent activation energy for volatilisation processes of Fe, Mg, and Si were 1043.02?±?12.29, 253.15?±?7.63, and 314.46?±?6.04?kJ?mol^?1, respectively.     

Dissolution Kinetics of Iron from Low-Grade Mn Ore and Optimization Using Response Surface Methodology

In this study an attempt has been made to increase Mn/Fe ratio in dump Manganese ore fines so that it can be used for the production of ferromanganese. For this purpose non-coking coal was used as reductant and dilute hydrochloric acid as leaching medium for the roasted ore. The effects of acid strength, leaching time, leaching temperature, stirring speed, ore particle size and pulp density have been studied. The dissolution of iron follows the kinetic model 1 ? 2x/3 ? (1 ? x)^2/3 = k_dt. Thus product layer diffusion is the controlling mechanism and the activation energy has been determined to be 26.23 kJ/mol at 40–95 °C. Another set of experiments have been conducted according to 2³ full factorial design, and regression equation for iron dissolution has been developed.     

Enthalpies of formation of borides of iron,cobalt, and nickel by solution calorimetry in liquid copper

The enthalpies of formation at 1385 ±2 K of the following crystalline borides have been determined by high temperature solution calorimetry using liquid copper as the calorimetric solvent. Fe₂B-67.87 ±8.05 kJ mol⁻¹, Co₂B -58.1 ±7.0 kJ mol⁻¹, Ni₂B -67.66 ±4.12 kJ ml⁻¹, FeB-64.63 ±4.34 kJ mol⁻¹, CoB -69.52 ±6.0 kJ mol⁻¹, and NiB -40.2 ±3.77 kJ mol⁻¹. The enthalpy of fusion of NiB has been determined to be 28.25 ±1.54 kJ mol⁻¹ at its melting point of 1315 K. New data are reported also for the enthalpies of solution of iron, cobalt, and nickel in copper, and for the enthalpies of interaction between these metals and boron in dilute solutions in liquid copper.     

Kinetic investigations of two processes for zinc recovery from zinc plant residue

Abstract

This paper presents a kinetic study of hydrometallurgical and pyrometallurgical processes employed for the recovery of zinc from a zinc plant residue (ZPR). The hydrometallurgical process investigated involved zinc leaching from the ZPR with sulphuric acid solution, while the pyrometallurgical process was based on zinc oxide reduction from the ZPR using activated carbon as the reducting agent. At the optimum leaching conditions, the zinc recovery was 89·5% at 95°C after 60?min. The data obtained for the leaching kinetics indicated that the dissolution of ZPR is a diffusion controlled process and the activation energy is equal to 21±2?kJ?mol^?1. The maximum zinc recovery, obtained during pyrometallurgical treatment was 72% at 1300°C after 60?min. Kinetic investigations revealed that the zinc oxide reduction is carried out in the transient kinetic area, where both diffusion and the chemical reaction control the overall process rate, with an activation energy of 38±5?kJ?mol^?1.

Cet article présente une étude cinétique d'un procédé hydrométallurgique et d'un procédé pyrométallurgique utilisés pour la récupération du zinc du résidu d'une usine de zinc (ZPR). Le procédé hydrométallurgique investigué impliquait la lixiviation du zinc du résidu de l'usine de zinc avec une solution d'acide sulfurique, alors que le procédé pyrométallurgique était basé sur la réduction de l'oxyde de zinc du résidu de l'usine de zinc en utilisant du charbon activé comme agent de réduction. Aux conditions optimales de lixiviation, la récupération du zinc était de 89·5% à 95°C après 60?min. Les données obtenues pour la cinétique de lixiviation ont indiqué que la dissolution du résidu de l'usine de zinc était un procédé contrôlé par la diffusion et l'énergie d'activation est égale à 21±2?kJ?mol^?1. La récupération maximale de zinc, obtenue lors du traitement pyrométallurgique, était de 72% à 1300°C après 60?min. Les investigations de la cinétique ont révélé que la réduction de l'oxyde de zinc était effectuée dans la zone de cinétique transitoire, où tant la diffusion que la réaction chimique contrôlent la vitesse globale du procédé, avec une activation d'énergie de 38±5?kJ?mol^?1.