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By A. Warczok, T. A. Utigard

"Slags from smelting of copper concentrates contains from 1 to 15% of copper. Copper is present in the form of copper or copper matte inclusions of size from 2 to 1000 micro, and as dissolved cuprous oxide. Removal of the copper requires the slag reduction, primarily reduction of magnetite and co-reduction of cuprous oxide. Thus, the kinetics of the slag reduction play an important role in copper recovery. Instantaneously, at the contact of carbon and a slag the gaseous film is formed. The magnetite and cuprous oxide reduction is going through a gaseous film, separating the slag and carbon of a coke or electrodes. The thickness of a gaseous film grows to a critical value and gas bubbles, containing CO and CO2, are created and rise up in the slag. Results of measurements of kinetic of the slag reduction with graphite showed high value of activation energy, close to the value of activation energy of Boudouard reaction. Strong temperature dependence of the reaction rate points out the thermal instability of the reacting system. It was found that the slowest process in the reduction stages determines the loop of CO and CO2 transfer and magnetite diffusion to the slag/gas film boundary. Strongly endothermic Boudouard reaction seems to cause thermal instability on the reaction surface. So, heat transfer to the reaction surface is very important. Due to low thermal conductivity of a slag the convection dominates in the heat transfer. So, any slag stirring accelerates the rate of reduction and coalescence of metallic copper or copper matte inclusions."

Jan 1, 2015

By Richard Elliott

Current methods for producing low-carbon manganese alloying materials for the steel industry are energy intensive and costly. To enable wider implementation of high Mn steel alloys, a novel solid-state Mn ore reduction strategy is therefore being investigated. This solid-state process, modelled on existing Direct Reduced Iron technology, employs a methane-containing atmosphere to reduce Mn ore pellets to a metal carbide. These pellets may then be smelted, yielding a Mn-rich master alloy. It is well established that CH4 is a suitable reductant for Mn ore powders; however, the reduction behaviour of industrially relevant Mn ore pellets by CH4 is poorly understood. The present work reports the results of an investigation undertaken to address this knowledge gap. The results of this study highlight the significantly different reduction behaviour exhibited by the Mn ore pellets relative to the ore powder under similar conditions and suggest that the diffusion of CH4 into the pellet interior is the limiting factor in achieving complete reduction.

By N. Anacleto, S. Ganguly

Reduction of Groote Eylandt (Australia) and Wessels (South Africa) manganese ores using CH4-H2-Ar gas mixture was investigated in a fixed bed laboratory reactor in the temperature range 1000?1200oC. The extent and kinetics of manganese ore reduction as a function of gas composition and temperature were determined by on-line off-gas analysis using mass-spectrometer and dew point sensor. Morphology of ores and its change in the course of reduction was examined by optical and scanning electron microscopy. Phases of raw materials and reduced samples were analysed by XRD and EPMA. Manganese oxides were reduced to carbide Mn7C3. High extent and rate of reduction by methane-containing gas in comparison with carbothermal reduction were attributed to high carbon activity in the reducing gas, which was in the range 15 - 50 (relative to graphite). The reduction rate of Wessels manganese ore increased with increasing temperature. Reduction rate and extent of Groote Eylandt manganese ore achieved maximum at 1050oC. The decrease in rate and extent of reduction of Groote Eylandt ore at higher temperatures, particularly at 1150-1200oC, was due to sintering and formation of semi-liquid silicate slag. The optimum conditions for the reduction of manganese ores were established as follows: a) temperature: 1150-1200oC for Wessels and 1050-1100oC for Groote Eylandt ores, b) methane concentration: 10-15 vol%, c) hydrogen concentration: above 30 vol%, (d) size: 1-3 mm. An addition of lime (10-15 wt% CaO) to the Groote Eylandt manganese ore increased melting temperature of slag and significantly increased the rate and extent of reduction at elevated temperatures.

Jan 1, 2004

By Kai Tang, Casper van der Eijk

Natural gas has been used as reductant in the production of iron for several decades. Recently, the use of natural gas as reductant for other metals has been drawing attention because of the low price of natural gas, as well as the reduction of CO2 emissions compared to the use of coke and coal. In this work, the reduction of a mixture of Ti- and Ni-oxides by natural gas has been studied. Within the Ni-Ti system, the intermetallic compound Ni3Ti is thermodynamically the most stable one. Therefore, powders of Ni-oxide and Ti-oxides were mixed in an atomic ratio of three to one and subsequently pelletized. Reduction was performed in a TGA furnace at constant temperature with varying reducing gas atmospheres. In a natural gas containing atmosphere, the Ni-oxide was easily reduced to metallic nickel while the Ti was reduced to Ti-oxycarbides or Ti-oxynitride when nitrogen was present. Depending on the process parameters, several percent of Ti was dissolved in the nickel-based solid phase.

Jan 1, 2014

By José R. de Tenorio

The aim of the present work was to investigate the reduction process of MnO~FeO-Si02 slags by carbon-saturated liquid iron. The effect of the initial Mn content in the bath was explored. The initial Mn content was in the range of 0 up to 40 wt. %. The MnO content in the slag was kept constant at 40 wt. %. Tests were performed in laboratory apparatus. Samples of slags and bath were taken. In the first 5 minutes, it was observed a very strong reduction of FeO. This first step promoted a MnO-Si02 enrichment of the slag and a formation of CO bubbles, which increases the slag level. The increase ofMn content decreases the MnO reduction rate.

Jan 1, 1996

By G. Castillo, J. Mortensen

"INTRODUCTION Molybdenite is typically produced as a co-product in the copper ore flotation process. At the Rio Tinto Kennecott Concentrator (RTKC) in Copperton, Utah, the molybdenite is initially floated with copper sulfides and other value minerals producing a Cu-Mo concentrate. This is then separated in the molybdenum plant by depressing copper sulfides and pyrite and floating only molybdenite. The molybdenum plant in RTKC comprises the rougher circuit and several stages of cleaning to achieve a certain grade requirement of Mo. The rougher circuit is where the first stage of separation occurs. Cu sulfides and pyrite are depressed in the rougher circuit by adding NaSH under nitrogen and in the cleaner circuit by adding NaSH under air. The depression of Cu sulfides (and pyrite) by NaSH is related to the redox activity1. It is important therefore to maintain a reducing potential in the circuit in order to achieve effective Cu-Mo separation. If the rougher circuit is under air, like RTKC’s molybdenum cleaner circuits, it is harder to maintain the target redox potential because NaSH oxidizes rapidly in the presence of air thereby increasing NaSH consumption significantly. CU-MO SEPARATION AND ROLE OF PULP POTENTIAL The collector adsorption on copper sulfide minerals is through an electrochemical mechanism2. The collector changes the surface chemistry of the Cu mineral allowing it to float. Molybdenite, or MoS2, on the other hand, floats readily in the presence of a frother as it is naturally hydrophobic. Usually, hydrocarbon oil extenders such as diesel or burner oil are added to enhance Mo recovery. In order to effectively separate Cu from Mo, the collector complex on the copper surface needs to be destabilized or desorbed using a reducing agent such as NaSH or Nokes. As long as the circuit is under highly reducing conditions, the collector remains detached from the sulfide mineral surface allowing Cu sulfide depression. Therefore, it is important to measure the oxidation-reduction potential (ORP) continually and maintain the reducing condition with addition of NaSH. For a given collector each mineral will have a characteristic potential or electrochemical window for flotation to occur3, therefore, it is necessary to reduce the potential outside of this window in order to effectively separate Cu and Fe sulfides from MoS2 minerals2. In RTKC plant, the target reducing condition is between -450 to -550 mV to effectively destabilize the collector and depress Cu and Fe minerals. Typically the dosage of NaSH used in the molybdenum plant could range from 6 to 30 gallons/min (gpm). The high dosage of NaSH could pose risk of H2S exposure. RTKC wanted to evaluate alternatives to NaSH in order to reduce its consumption and minimize risk of H2S gas generation."

Jan 1, 2016

By Patricia Martínez Nicolás

Nickel and ferronickel are some of the major components in the production of special alloys; however, the energy consumption for their production greatly increases their price. A cheaper option could be to use NiO and a reducing agent. This work presents experimental and theoretical analysis to estimate the effect of slag basicity and amount of reducing agents on the NiO reduction from the slag which interacted with molten steel at 1600 °C in an EAF. The slag system contained CaO, SiO2, CaF2 and NiO together with a reducing agent, ferrosilicon or carbon. The NiO content in the slag was 25 mass%. The amounts of the reducing agents (Si and C) were 1 to 4 times the stoichiometric mass to reduce NiO. The CaO/SiO2 ratio was 1.5 and 2.5. The kinetics of the process was very fast since after few minutes the maximum nickel yield was obtained. The slag basicity had a negligible effect on the reduction of NiO.

Jan 1, 2006

By N. A. Ziegler

THE chief difficulty in determining oxygen in steels is its tendency to form a variety of compounds. Almost every element, found as an ingredient in steels, maybe expected to be present as an oxide. Some of these oxides have widely differing properties, and consequently behave differently when attempts are made to reduce them by the vacuum fusion method. As I pointed out in a paper published in the Transactions of the A.S.S.T., Vol. 20, No. 1, many experimenters are spending considerable time and effort in attempting to deter¬mine the degree of completeness of the above reactions, and hold different opinions on the matter. In studying this subject the following table, taken from Hodgman and Lange's "Handbook of Chemistry and Physics" may be used as a guide.

Jan 1, 1933

Hydrogen and carbon monoxide reduction data were obtained on pelletized materials of the type currently being prepared commercially in Australia.Pellets, approximately 1 cm diameter, made from Deepdale, Iron Prince, and Mt. Newman iron ore were reduced by hydrogen over the temperature range 500°-900°C and the reduced pellets were examined by thermogravimetric, microscopic, electron probe, and x-ray techniques. After an initial time delay it was found that up to 95 per cent reduction the data could be described by a porous model where gaseous diffusion through the solid is rate controlling. Independent measurements of effective diffusivity supported the values obtained from the reduction data. Above 95 per cent reduction the rate decreased markedly due to the formation of retained wustite or fayalite.Comparison with reduction by carbon monoxide over the temperature range 750°-900°C showed that the rate of reduction was approximately five times slower than for reduction by hydrogen.The rate of reduction was found to be a function of the initial properties of the pellets which may account for the widely conflicting results that have been obtained with previous investigations on porous pellets.

Jan 1, 1971

By Alain Denis, Marcelo Curtes Martins, Jeferson Rocio de Oliveira

In this work we present a study of optimization of a new inhibitor for turbulence pickling line. The efficiency of inhibition was measured and the weight of loss was evaluated in four different materials to optimize the dosage of the inhibitor. In seven months we had 8.7% of reduction in pickling losses.

Oct 1, 2019

By R. G. Reddy

Solid state reduction of roast movat chromite ore with carbon was carried out as a function of C/FeO ratio and temperature in the range 800°C to 900°C by thermogravimetry analysis (TGA). A two stage mechanism, (1) the reduction of FeO present in Roast Movat with carbon monoxide; and (2) the regeneration of CO by the oxidation of carbon with CO2 was proposed. The rate controlling step is the carbon oxidation reaction. A kinetic equation, that of random nucleation, was proposed. The activation energy for the reduction was determined to be 55800 J/mole and is compared to those activation energies determined for other chromite ores.

Jan 1, 1998

By James T. Skoch

With the ongoing pressure on the secondary aluminum business to limit the amount of waste generated from processing aluminum dross, it becomes very advantageous to preprocess the dross to reduce waste. The advantage of preprocessing is that it significantly reduces the oxides and salt cake generated when melted. Various products can be produced from the oxides for the steel industry. The following paper will demonstrate that removal of oxide fines before melting will result in large volumes of material not being landfilled. We will show that the end result of this technology is to significantly reduce the amount of salt cake sent to a landfill while maximizing the recyclability of the aluminum dross.

Jan 1, 1995

By Nitin Bhavsar, Kaushik Vakil, Lakshmikanth Reddy, Sanjay Sarkar, Tathavadkar. Vilas, Bhavin Desai, Raghavendra Adiga

Gold and platinum group metals (PGM) do not dissolve in sulfate electrolyte from anode during electro-refining of primary copper. The PGM form solid ‘slimes’ which accumulates at bottom of the electrolytic cell. Typically, 8 to 10 ppm silver is present in copper cathodes, which is equivalent to almost 2% of silver loss during electro-refining. The silver can be cemented with cathode copper as soon as it is dissolved into the electrolyte from anode or during entrapment of slime particles in cathode. Over a period of time, the concentrates grade has deteriorated and change in type & amount of impurities in concentrate has resulted in higher silver loss during refining process. Therefore, an in-depth investigation was carried out in both fire- and electro-refining processes to understand the root causes of higher silver loss in copper cathodes. The detailed characterization and the plant data analysis were carried out. The results of study revealed that higher silver loss in 3rd crop cathodes as compared to 1st and 2nd crop cathodes from same anode. In anode silver was present in two forms, dissolved in copper (solid solution) and in slime phases such as Cu2-xAgxSeTe. SEM-EDX study of anode cross section confirmed higher silver and oxygen content in central region (core) of anode as compared that in both air and mold sides. This is due to non-uniform cooling of anode which was confirmed by the solidification modeling. Higher concentration of oxygen & silver at core of anode rised Ag+ concentration in electrolyte and slime particles during 3rd crop. The results of study were used to optimize the anode cooling process and also electrolyte chemistry to reduce the silver loss. The details of study carried out are discussed in this paper using the results of modeling, characterization and experimental studies. INTRODUCTION In the electrolytic refining of copper, fire refined copper anodes are used as raw material. Anodes contains copper of the range 99.5 to 99.6%, rest being the impurities. Hence, the purpose of the refining is (1) to remove the impurities so that final cathode conforms to the LME specifications and also (2) to recover the precious metal from the anode copper. When anode is dissolved, most of the impurities reports to the anode slime settled at the bottom of the cell, and rest dissolves in the electrolyte. Dissolved impurities remains in electrolyte and are removed during electrolyte recycling. However, sometime dissolved impurities and slime particles get entrapped in cathode due to disturbances in operational conditions. Metals dissolving in the electrolyte are the elements such as Ni, Co, Fe, Zn. The electrode potential of As, Sb and Bi being close to that of copper, these elements partly dissolve in the electrolyte. Noble metals like Ag, Au, Pt and Pd are generally insoluble and are thus expected to report to the anode slime. A small amount of the sliver content from the anode is dissolved in the electrolyte. Because the dissolved silver electroplates on the cathode at a smaller applied potential than copper, the cathode typically contains up to 8 ppm of silver. This is referred to as ‘silver loss’ in cathode (Cooper, 1987). Fortunately, silver is a rather benign impurity in copper due to its electrical conductivity being higher than copper. However, higher silver loss in cathode is loss of revenue for Copper smelter due to significant difference in their LME price.

Jan 1, 2019

By P. Barrios

Over a period of one year, an in-depth investigation was carried out in a conventional copper refinery to understand the behaviour of silver in copper electrolysis. The increase in silver content in the cathodes increases almost linearly with the duration of electrolysis, and the silver content of the electrolyte between the electrodes shows similar behaviour. The silver content of the cathode deposited daily depends directly on the silver concentration of the electrolyte and this, in turn, is strongly influenced by its internal temperature. The average temperature between the electrodes rises from the first to the second and to the third crop, because of the increase in the distance between the electrodes and the heat generated by the passage of the current. The mechanism of dissolution of silver from the anode, and its passage to the cathode, are discussed. Following the change to the Isa system, it was proved that the results and conclusions are analogous. Based on these results, a reduction of more than 20% in silver losses in the cathodes was achieved.

Jan 1, 1999

By E. M. Weenink

A proposed alternative to the conventional zinc extraction processes is the flash reduction of zinc calcines. As part of an overall study of this process, the kinetics of single zinc oxide pellets were investigated. Pellets, compacted from Outokumpu zinc calcine (60.9% zinc, 9.6% iron), were suspended in a flow of CO/CO, mixtures and reduced at temperatures between 1000 and 1200°C and atmospheric pressure. The pellet weight loss with time was recorded and analyzed. It was found that the reduction of calcine pellets (0.9 cm height and 1.36 cm diameter) was controlled by chemical reaction. This paper includes an analysis of the experimental results of two previous investigations of ZnO reduction.

Jan 1, 1992

By Jörg Wallner

In order to limit SO2 emissions, the Montanwerke Brixlegg AG has installed a desulphurisation plant with the primary aim of reducing the peak emissions from the copper converter that are of particular ecological concern. The author discusses the process requirements, the desulphurisation process selected and the results achieved.

Jan 1, 1996

By Vladimir Lobanov, Pritchinm Alexander, Nikifor Ageev

"Sulfuric anhydride is one of the most voluminous technological emissions. The possibility of high-temperature SO2 oxidations without catalyst in presence of calcium-bearing filling materials (CaO, Ca(OH)2 and CaCO3) with obtaining of binding material on the basis of waterless calcium sulfate is studied. Due to application package HSC-4 the thermodynamic analysis of the specified interactions is made and following conclusions are drawn: 1) Reactions are feasible in the all range of investigated temperatures (400 – 1000 °?); 2) The best results should be expected at use of ???; 3) The heat of reaction does possible autogenous behavior of process. The researches have shown: 1) Practical possibility of the process; 2) SO2 oxidation without the catalyst is feasible in presence of ???, Ca(OH)2 and CaCO3; 3) The optimum temperature range is 600-800 °?; 4) The content of waterless calcium sulfate in a product reaches 48% that makes it possible to use as the binding material; 5) At excess of ??? the degree of SO2 collection is nearer to 100%. Introduction At processing of sulfide raw materials in metallurgical furnaces the significant amount of sulfur, mainly in the form of sulfuric anhydride is escaped. Rich gases (SO2>3 %) are used for the sulfuric acid or element sulfur production. Chemism of sulfuric acid manufacturing is based on catalytic high-temperature oxidation of anhydride. Poor gases of metal works (SO2 =2,5 %) and power combustion units (SO2 =0,5 %) after clearing discharge into the atmosphere. Commonly, to recovery of weak sulfuric anhydride the chemical and physicochemical methods are applied. A few chemical ways of recycling sulfur dioxide are applied practically. In a number of methods for SO2 recovery a pulp of metals oxides (magnesium, zinc, etc.) with formation of sulfites is used. In other case sulfuric anhydride is catch by weak sulfuric acid in the presence of an oxidizer (pyrolusite ore). Methods of absorption SO2 by solid sorbents, (carbon fibers, silica gels, zeolites) are known and put into practice. Regeneration of the saturated sorbents is realized by thermal decomposition or extraction. Obtained as a result high-concentrated solutions are a commercial product."

Jan 1, 2008

By G. Tairova, M. Murchie, B. Esford, E. Lam

"The objective of this project is to investigate methods to lower the sulphate concentrations in the mill effluent. Currently, the concentration of sulphate in mill effluent is approximately 2 g/L. The high concentration of sulphate ion in the effluent has the potential to become a regulatory issue. The target sulphate concentration in mill effluent for this work was less than 500 mg/L. The pilot studies, using sulphate precipitation, were carried out at the mill. The results of the sulphate precipitation pilot study, using calcium hydroxide and sodium aluminate, indicated that:• The sulphate concentrations in effluent could be reduced to less than 500 mg/L.• The concentrations of most other elements of concern, such as selenium, arsenic, cobalt, nickel and vanadium, were also reduced.INTRODUCTIONHistorically, the concentration of sulphate, SO4 2-, in water has not generally been subject to regulation. However, with the increasing global consumption of water, stricter regulations on sulphate quantity may be imposed. Sulphate is a concern in process waters as it forms scale on process equipment and leads to operational problems. High sulphate concentrations may result in adverse health effects. The taste theshhold for the most sulphate salts in drinking water ranges from 250 to 500 mg/L.A literature search and review to identify methods to reduce sulphate concentration in the mill effluent was carried out. A number of methods were selected and tested in the lab. The experimental results indicated that the precipitation methods, using calcium hydroxide and aluminum compounds, barium compounds, as well as the ion exchange and the membrane methods, were effective for the removal of sulphate to below the required limits. Some other elements, such as vanadium, nickel, cobalt, selenium and molybdenum, were also removed.The process, using calcium hydroxide and aluminum compound for sulphate precipitation, similar to the Cost Effective Sulphate Removal (CESR), also known as “Walhalla” process, described in the Montana Tech report (“Demonstration Of Sulfate Removal” 1999), was selected for further consideration [1]. In the CESR process, the water is treated with lime for precipitation of gypsum, followed by the reaction with a proprietary aluminum compound for removal of a residual sulphate. The final pH adjustment is carried out using carbon dioxide. Results of a preliminary cost estimate indicated that the removal of sulphate using calcium hydroxide and aluminum compound was the preferred method of treatment, considering its low cost, availability of chemicals and minimal changes to the existing plant circuit."

Jan 1, 2012

By M. Abdel-latif, R. W. Bartlett, P. R. Taylor

A cyclone reactor system for the partial reduction and melting of taconite concentrate fines was designed and operated. A nontransferred arc plasma torch was employed as a heat source. Taconite fines, carbon monoxide and carbon dioxide were fed axially into the reactor, while the plasma gas was introduced tangentially into the cyclone. The average reactor temperature was maintained above 1673 OK (1400°C), and reduction experiments were performed under various conditions. The influence of the following parameters on the reduction of taconite was investigated: carbonmonoxide to carbon dioxide inlet feed ratio, carbon monoxide inlet partial pres- sure and average reactor temperature. The interaction of the graphite lining with carbon dioxide and taconite were also studied.

Jan 1, 1995

By P. R. Taylor

A cyclone reactor system for the partial reduction and melting of taconite concentrate fines has been engineered, designed and operated. A non-transferred arc plasma torch was employed as a heat source. Taconite fines, carbon monoxide, and carbon dioxide were fed axially into the reactor, while the plasma gas was introduced tangentially into the cyclone. The average reactor temperature was maintained at above 1673 K (1400 C), and reduction experiments were performed under various conditions. The influence of the following parameters on the reduction of taconite was investigated: carbon monoxide to carbon dioxide inlet feed ratio, carbon monoxide inlet partial pressure, and average reactor temperature. The interactions of the graphite lining with carbon dioxide and taconite were also studied.

Jan 1, 1993