Search Documents

By A. Mezei

A process has been developed to recover cobalt, gold and bismuth from the NICO deposit in Canada. The process consists of a bulk rougher flotation followed by an optional separation by flotation to produce a bismuth concentrate and a cobalt concentrate, grading on average 12.8 % Bi and 1.8% Co, respectively. Cobalt is leached by pressure oxidation at 97% efficiency, producing a pregnant solution suitable for further processing. Before optimization, about 38% of the bismuth is recovered in the selective concentrate, wherefrom it is recovered by ferric chloride leaching followed by cementation, at overall efficiencies exceeding 98%. The gold is recovered from the cobalt and bismuth leach residues by cyanidation, at 95% efficiency. Previous work described the results of the leaching of the concentrates, recovery of gold (Part I) and the recovery of bismuth (Part 11). This work describes the results of the cobalt recovery testwork and the resulting conceptual flowsheet. The sulfur dioxide-oxygen system was the most efficient for the oxidationlprecipitation of iron and arsenic from cobalt bearing pregnant leach solutions, producing purified solutions assaying <1 mg/L As and <8 mg/L Fe. The efficiency of cobalt solvent extraction (from the resulting purified PLS containing about 1 g/L Co) ranged from 9 1 to 98%; the co-extractions ranged from 88 to 100% for copper, from 28 to 33% for magnesium, from 6 to 7% for calcium and from 96 to 98% for zinc. Sulphide precipitation processes (using either sulfur dioxidelelemental sulfur or sodium sulfide) produced residual copper concentrations in the cobalt strip solution ranging from 0.1 to 2 mg/L Cu. About 90% of the calcium and magnesium were separated by scrubbing the organic with cobalt sulfate, producing residual concentrations of 12 mg/L Ca and from 180 to 330 mg/L Mg in typical feed solutions directed to cobalt carbonate precipitation tests. Cobalt carbonate precipitation efficiencies ranged from 75 to 78% when calcium and magnesium co- precipitations were restricted from 8 to 17% and 0.3 to 0.4%, respectively. The limits required by most stringent specifications on commercial grade cobalt carbonate products were met or exceeded.

Jan 1, 2002

By Shigeatsu Nakazawa

Activities of Bi and In in the Bi-In liquid alloy were measured over the whole range of composition at the temperatures from 850K to 1050K. An alloy was vacuum-sealed in a quartz cell and heated at the temperature of interest. The absorption for 307nm radiation from Bi lamp was measured for Bi atom vapor in the cell. By heating a pure metal as a standard and measuring the absorbance as a function of the temperature, an analytical curve for Bi atom vapor was constructed and used for conversion of the absorbance to the vapor density. Bi activity was determined as the ratio of the Bi atom vapor density over the alloy to that over a pure metal. The same procedure was applied to the In 304nm radiation from In lamp and In activity was determined independently of Bi. Results showed that the liquid Bi-In system behaves like a regular solution.

Jan 1, 2003

By Shigeatsu Nakazawa

Activities of Bi and In in the Bi-In liquid alloy were measured over the whole range of eomposition at the temperatures from 850K to 1050K. An alloy was vaeuum-sealed in a quartz cell and heated at the temperature of interest. The absorption for 307nm radiation from Bi lamp was measured for Bi atom vapor in the cell. By heating a pure metal as a standard and measuring the absorbance as a function of the temperature, an analytical curve for Bi atom vapor was constructed and used for conversion of the absorbance to the vapor density. Bi activity was determined as the ratio of the Bi atom vapor density over the alloy to that over a pure metal. The same procedure was applied to the In 304nm radiation from In lamp and In activity was determined independently of Bi. Results showed that the liquid Bi-In system behaves like a regular solution.

Jan 1, 2003

By A. Martins, L. Goulart, E. Araujo, R. Batista, A. Costa e Silva

Steel cleanness is of great importance to the performance of almost all steel products. From very clean steel for bearings to long products used in civil construction, processing variables – in special those in steelmaking operations- must be properly balanced to achieve adequate cleanness for the desired performance. With the present complexity of steels as alloy systems, it is not efficient nor appropriate to develop these processes on empirical basis alone. Computational thermodynamics can greatly improve process development to achieve the desired cleanness level at reasonable costs. In this work, examples of these calculations and their applications to real steel processing are presented and discussed. The advantages, limitations and areas in which improvement in these calculations is desired are highlighted and discussed.

Jan 1, 2015

By C. B. Catanach

The development and in place leaching of a chalcocite ore body at Mountain City, Nevada is described. Block caving was used to induce permeability while ore removed during development was leached in heaps on the surface. Copper was recovered by precipitation on iron. Excess ground water which seeped into the mine was pumped into a deep disposal well. A parallel system was used to purify water by lime neutralization. Copper extraction was at an annualized rate of 11% during the first six full months of underground leaching and totalled approximately 30% during six months of heap leaching. Leach solution flow and distribution had a major effect on extraction rate. Enrichment of leach solutions with ferric sulfate did not produce conclusive results until analyzed with the aid of a leaching model.

Jan 1, 1976

By Jintao Gao, Zhancheng Guo, Chong Li

A new approach to enrich MFe and RO phase from the converter slag by super gravity was investigated. The samples obtained by the gravity coefficient G=600 and cooling rate V=5K/min from 1623K to 1373K appeared significant stratification and the MFe gathered at the bottom of sample. With the help of metallographic microscopy and X-ray diffraction, it is found that large quantity of white RO phase gathered at the lower part of the sample as the concentrate, with volume fraction appearing gradient distribution along the direction of super gravity, while the upper part was mainly composed of dicalcium silicate as the tailing. The mass fractions of the FeO, MgO and MnO in the concentrate are up to 40.46%, 6.9% and 3.91%, while that in the tailing are just 11.65%, 2.15% and 1.55%. The recovery ratios of FeO, MgO and MnO in the concentrate are 60.35%, 58.45% and 72.24%, respectively.

Jan 1, 2016

By Yong Li, Tao Jiang, Xiangxin Xue, He Yang

"In this paper, we describe the current status and research progress on the utilization of typical industrial solid wastes in China. We propose a creative method called “eco-utilizing wastes” with an aim of both integration and increment but without second-time waste and pollution, in terms of the characteristics of typical industrial solid wastes, such as titanium-bearing blast furnace slag, high-silicon iron tailing and boron-enriched slag as well as oil shale.Situation of Typical Industrial Solid Wastes in ChinaThough mineral resources in China are abundant, most of these are low grade ores. Moreover, the distribution of mineral resources is asymmetric, which causes the increase of cost and decrease of the economic benefit. Over-exploitation of mineral resources results in a large number of solid wastes. The accumulation of solid wastes not only means the waste of valuable resources, but also causes environment deterioration. There exist some problems in the current practice of the disposal of industrial solid waste, such as only emphasizing the final results and causing new pollution and waste [1]. So the research and development of a high value-added method has a high practical relevance to our nation. Based on the present status and research progress of industrial solid wastes in China, a creative new idea and method for eco-utilization of solid wastes is proposed in this paper."

Jan 1, 2012

By Helmut Antrekowitsch

Nowadays nearly the half of the world zinc production is used for galvanizing. After their life span most of the galvanized products return as scrap to the steel mill. During the steel process the zinc is collected in the flue dust It is astonishing to notice that between 40 and 50 % of the zinc produced is brought into the steel processing cycle by the coated scrap while only 3 to 5 % of the zinc production can be covered by the recycling of the zinc-bearing steel mill dusts. Responsible for this worse ratio is the fact that the established recycling technologies suffer from high energy consumption, low zinc yields, halogen problems and hardly any iron recovery. A development of new or an optimization of current recycling procedures for the dusts is unalterable because of increasing zinc amounts in these residues and the rising costs for the disposal. Different pyrometallurgical recycling technologies are tested in rotary kilns and vertical retorts at the Department of Nonferrous Metallurgy, University of Leoben, Austria. Various reducing agents from carbon monoxide to hydrogen and combinations are used. The results should make it possible to draw conclusions that allow an economical and ecological judgement of the different varieties. The goal is to support the affected industry to increase the low recycling rate and prevent a loss of high zinc values in the steel industry filter dusts.

Jan 1, 2003

By Semih Perdahcioglu, Arthur E. Morris

"The curricula for the majority of Materials Science and Engineering programs is devoid of all but the most rudimentary coverage of the principles of material and energy balance calculations. As a result, graduated engineers that need to make such calculations are faced with three unpalatable options. First, attempt to learn the subject by attending workshops or short courses sponsored by corporations or industry groups. Second, outsource the needed work to a consulting engineering company. Third, purchase a software package that's capable of simulating a process. Each of these options require a considerable investment of employee time and effort. The first option might be the best if suitable computational software was readily available. We review the recent development of such software, which is entirely based in Excel. The software tools include a thermodynamic database, tools for multiple computations, and a flowsheet simulation program. These tools employ teaching by simulation.IntroductionPracticing engineers often come up with situations where they must learn a technical subject that they didn't cover in their college classes. One possibility is to obtain a couple of texts on the subject, and delve in. Or, they may sign up for an intensive short course where they can get personal instruction and guidance. But the first option requires a lot of study time, and the second option may not provide the specific instruction (or intensity of instruction) required to appropriately master the subject. Here we look at ways of using simulation techniques to learn the methodology for making material and energy balances on industrial processes.Industrial heating processes have undergone many changes over the past dozen years, as vendors have introduced new heating equipment designed to recover the greatest possible amount of heat from the fuel, and to meet stringent environmental challenges for emissions (especially NOx). Even though the recent unprecedented drop in price of natural gas may have altered the urgency for improving a plant's energy efficiency, the need is still there. Until recently, there was a dearth of readily-available software for process engineers to use for simulating a plant's flowsheet. The good news now is the existence of several Excel-based software tools that make the calculational process much easier. [ 1]. Plant engineers may also be interested in a DOEsponsored program[2] which allows plant-wide and ""what-if' analyses of the energy balance."

Jan 1, 2013

By Yafei Xiong, Yue Wang, Guotao Xu, Shuzhong Li, Huaiyuan Li

"The composition, properties and structure of taphole clay with high strength and friendly surrounding used in a new blast furnace with 3800 M3 volume was studied. The developed taphole clay had been applied in the blast furnace for four months, pouring times were 11-13 every day, flowing rate of hot iron was 7.5 t/min and the consumption of taphole clay was 0.49Kg per ton iron, which can make the new blast furnace attach the designed production quickly and decrease the cost of hot iron. The taphole clay is easy to be opened in the channel of hot iron which can be remained in stable shape under the corrosion of slag or hot iron, and it dose not produce poisonous smoke in yellow or dark color in the use process. The development of new taphole clay is useful to improve the work surrounding of the blast furnace.IntroductionIn the iron-making process, the taphole clay's function is filling up the taphole and protecting the formation as well as the growth of the clay in the blast furnace, maintaining the enough taphole depth; simultaneously it is easily to open a hole and knock out the clay. The taphole clay has a good performance of molten slag and hot iron resistance. The traditional taphole clay uses coal tar or carbolineum as a binder, and the performance of taphole clay is good, but it has the black and yellow poisonous smoke, which is harmful for work condition. Through the investigation on the binder's composition and poisonous performance of taphole clay, the environmental taphole clay is developed by some foreign companies, which is named the third generation product with benzene is less than 300ppm, and it has been used in European steel companies. This application of the taphole clay with high strength and environment-friendly using in 3800m3 blast furnace is introduced and the performance is discussed and analyzed in this paper."

Jan 1, 2012

By John H. Schloen

Copper refineries are constantly looking for methods to upgrade their plants to reduce refining costs and to improve the purity of their electrolytic cathodes at the same time that the purity of the anodes they refine is being lowered by the rehabilitation of reverberatory furnace smelters to one of the direct smelting processes. The comments given below are based on the answers to the questionnaire returned bf the 36 copper refineries that represent 66% of the western world s refining capacity. Eleven refineries employ some form of uniform weight anade casting, e.g., Outokumpu, Contilanode, Mitsubishi, Lurgi, or Hazelett. This type of anode casting presents the best possibility for improving cathode purity. The cathode press of Phelps Dadge can improve current efficiencies and cathode quality. The press is used on first-run cathodes, after 48-60 hours, on half-cells at a time without interruption of the cell current. One refinery interrupts current for sever.al hours to allow the anode slime to settle before the, cathode pressing. The cathode press is now used at eight refineries. Apart from periodic reverse current (PRC), it is not expected that current densities will be increased. Thirty refineries, using direct current, average 231 amps/M, with a low of 178 and a high of 350. Five refineries use PRC at an average of 338 amps/M, with a low of 290 and a high of 380. Current efficiencies are good at an average of 94%, with a low of 88% and a high of 99% for the 34 refineries that reported. It must be remembered that the cycle efficiency with PRC is about 93%., so that the effective current efficiency with PRC would be only .94 X .93 or 87% However, PRC should be considered seriously for refinery expansions because it can save about 25% on capital costs.

Jan 1, 1987

By C. D. Mishra

The recycling potential for any gainful utilisation of the waste products is dictated by their quality, especially the chemical characterisation. The methods commonly used for the analysis of fly ash e.g. wet chemical, AAS, ICP-AES etc are time consuming. The paper describes a rapid XRF-based method for routine analysis of fly ash which involves preparation of local standards, optimisation of sample preparation technique and use of a simple PC-based software to account for interelement effects. The matrix corrected values are found to agree well with the wet chemical data.

Jan 1, 1992

By E. V. Verhoef

The proposed EU directive on the restriction of the use of certain hazardous substances in electrical and electronic equipment promotes the substitution of lead in consumer electronics to reduce emissions from lead in the end of life phase. In this paper, the consequences of substituting lead in solders is assessed using the Eco-lndicator-99 method for the examples. The base metal production system is considered as an interconnected network modeled dynamically in Simulink 4 and Matlab 6 (Mathworks Inc 2000), where the recovery and production of metal components of the assessed product partially overlap. This approach requires global modeling of metal production processes and mass balances and resulting allocation choices. It also reveals the susceptibility of LeA scores to different choices in the modeling and data availability and uncertainty. Besides the comparison between lead-free and lead-containing solders, other side effects are revealed. Lead is carrier material for the recycling and production of many co-elements in the interconnected production and recovery system. Substitution may have adverse effects such as temporary shortages of other connected metals that are often not considered in normal LeA assessments, thus leading to erroneous results. The suggested approach provides directions for including these external effects and for further assessment of the environmental effects of substituting lead. This paper shows that in order to assess the environmental aspects of metallurgy, dynamic modeling of the total system is required based on the fundamental thermodynamics in each metallurgical reactor, the basis of this dynamic model. With this the true environmental impact of metallurgy can be explained. In the case of lead free solder, it is shown that the mining of substitutes has a very much larger impact.

Jan 1, 2003

By Mikiya Tanaka, Mikio Kobayashi, Mansour A. S. A1Ghamdi, Kenji Tatsumi, Yukinori Saiki, Hirotaka Senba

"In order to recover nickel ion in a spent electroless nickel - phosphorus alloy plating bath and, at the same time, to remove impurities such as iron and zinc, the applicability of solvent extraction with LlX84I as the extractant has been investigated using actual spent solution. Although the extraction efficiency of nickel from the spent solution is considerably lower than that from the nickel sulfate-nitrate solution, nickel in the spent solution is selectively extracted with high efficiency at the pH of 6 - 7. The extracted nickel is effectively stripped by sulfuric acid at a pH of less .than· 3. On the basis of these results, a· flowsheet for recovering nickel from the spent solution is proposed Chemical equilibrium calculations of the spent solution has shown that the complex formations of nickel with anions such as lactate and phosphonate ions become predominant in the pH range between 2 and 6. This results in the depression of nickel extraction efficiency from the spent solution as compared with that from the .nickel sulfate-nitrate solution. By combining these aqueous equilibrium calculations with the· extraction model considering the extraction stoichiometry and the aggregation of the extractant in the organic phase, the extraction behavior of nickel is quantitatively predicted.IntroductionElectroless plating is indispensable for the recent advanced industries such as the electronic and mechanical industries. Electroless plating baths are repeatedly used by supplying the metal salts; however, the accumulation of the counter anions of the metal ions, the components formed by the oxidation of the reducing reagents in the plating baths, and the metal ions as impurities dissolved from the surface of the objective materials degrades the· quality of the plating. Thus, the baths are finally treated by a precipitation method to form a large amount of metal sludge, which is disposed in dumping sites. Therefore, from the standpoints of environmental protection and resource preservation, it is highly desired to establish a recycling process of the plating baths with minimum emission. Solvent extraction seems to be one of the 'suitable methods to recover metals from the spent electroless plating baths, because the concentration of the metal ion in the baths is relatively high, which is advantageous to solvent extraction as compared with the ion-exchange and adsorption methods. Some studies about the application of solvent extraction have already been carried out using simulated spe11t electroless plating baths [1-3]."

Jan 1, 2000

By Motohiro Horiguchi, Kazuo Ichii, Toshio Oishi

"As a fundamental study to prepare copper-based alloys with rare earths (RE: La, Gd, Nd, Ce and Y) without any inclusions, the solubility of rare earth in liquid copper was experimentally determined. Liquid copper alloys were equilibrated with carbon and rare earth carbide in a graphite crucible in an argon atmosphere at 1523-1873 K.Solubilities of rare earth in liquid copper were determined by approaching the equilibrium values from both the lower and the ·higher concentrations than those expected· in the preliminary experiments. They were expressed as a function of temperature. Carbon in Cu-RE alloy was also analyzed and thermodynamically discussed.IntroductionOver the past several years, the demand for highly conductive and high strength copper-based alloys, such as Cu-Cr, Cu-Zr and Cu-Cr-Zr for use in electronic materials has increased.1 -7 These alloying elements are very reactive and easily oxidized to form oxide inclusions during melting. To avoid the oxide inclusions, it may be effective to utilize the strong reducing power of carbon in conditions under which no carbides form.In this study, to provide fundamental data for producing Cu-RE( rare earth) alloys without any inclusions, the solubility of rare earth in liquid copper in equilibrium with .carbon .was determined and discussed on the basis of the experimental results and the thermodynamic data.ExperimentalSample preparation for equilibrium experimentsOxygen-free copper (99.99%), lanthanum (99.9%), gadolinium (99.9%), neodymium (99.9%), cerium (99.9%) and yttrium (99.9%), and their oxides were used as starting materials.Each rare earth carbide was prepared by mixing rare earth oxide with carbon, which is a little more than the stoichiometric composition, pressing it into a tablet form and sintering at about 2000.K for 20 ks in a vacuum furnace using a graphite crucible. Powder X-ray analysis showed that only the carbon and rare earth carbide (REC2) was present in each sample.Cu-RE mother alloys (5-20 mass % RE) were prepared beforehand by mixing metallic rare earth with oxygen-free copper, and melting in an electric arc furnace."

Jan 1, 2000

By A. Tsubouchi, F. Noguchi, T. Takasu

"A high current density operation is desired to improve the productivity in copper electrorefining. However, a lot of problems still remain such as passivation at anode, dendritic deposition at cathode and impurity behavior during the electro refining.Effects of the current densitY on the morphology of copper deposition on the cathode, and passivation at the anode were investigated at various fluid velocity of electrolyte in the present study.Although a granular appearance of the copper deposition was obtained with increasing high current density, it was depressed by an increase of the fluid velocity of the electrolyte. A uniform deposition on the cathode was also observed when the fluid velocity of the electrolyte become higher. It was ascertained that passivation at the anode was suppressed by circulating electrolyte during electrolysis. INTRODUCTIONIt is desirable to operate at high current density in copper electrorefining in order to improve plant productivity without enlarging the size of electrodes(1). However, there are many problems in copper electrorefining at high current density with the usual commercial practice, for example, anode passivation, dendritic deposition and an increase of impurities in the cathode.In general, it is effective to agitate the electrolyte in order to suppress these problems(2). Smith et a1.(3) have performed electro refining experiments at various current densities in the range of 8 to 32 Aldm2 and at electrolyte fluid velocities in the range of 40 to 120 mlmin, and they obtained well consolidated, void free copper even at high current density in a high electrolyte velocity. However, they did not discuss the recovery ratio and the passivation behaviour at the anode. Walsh et a1.(4)(5)(6) experimented with the rotating cylinder electrode, and ascertained that electrolysis can be carried out at high current densities under the condition of pure electrolyte.With regards to anode passivation, there are several papers by previous workers(7)(8). Our group ascertained the process of producing anode slime during electrolysis(9). However, these experiments were carried out without forced circulation of electrolyte. Therefore, in the present study, to solve these problems, we electrolyzed at current densities in the range of 4 to 50 Aldm2 with circulating electrolyte in the range of 0 to 48 mlrnin. The effect of electrolyte fluid velocity on the surface morphology of deposited copper at cathode and passivation at anode were investigated when electrolysis were carried out with circulating electrolyte at high fluid velocity and at high current density."

Jan 1, 1997

By Emil H. Geus

A process has been developed to recover aluminum from dross which does not require the use of salt flux as commonly employed, This results in a salt-free residue which has potential industrial uses or alternatively can be safely land-filled. The dross is first heated by an air-fuel or oxy-fuel burner in a sealable rotary furnace. At a suitable temperature the burner is turned off and the temperature of the dross thereafter controlled by means of an atmosphere of oxidative and inert gases. When the free metal has agglomerated it is first discharged from the furnace, followed by the residues.

Jan 1, 1995

By Kazuyuki Kakegawa

"Sintered body of lead zirconate titanate (PZT) was prepared through the spark plasma sintering (SPS) technique. When the mixture of raw materials was calcined at 800°C and sintered at 900°C for 10 min. with SPS technique, the sintered body had almost theoretical density but was a mixture of lead titanate (PT) and PZT. When a calcining temperature of 1100°C was employed, sintered body had a single PZT phase but a lower density. High dielectric constant was obtained for a sample calcined at 800°C, sintered with SPS tecbriique and fired again in the air at 1100°C.IntroductionSpark plasma sintering (SPS) is a recently developed technique for sintering metals and ceramics1. The scheme of SPS is shown in Fig. 1. Sample is put in a graphite die in a vacuum chamber. It is pressed and pulse voltage is applied. The electric current heats the die. It is said that the pulse causes a spark-plasma between particles in the sample. Because of the spark- plasma, sample can be sintered at temperatures lower than that in the ordinary sintering by several hundred degrees of Celsius. The sintering period is also very short2Low sintering temperature and short sintering period in SPS technique cause low reaction rate. It is both disadvantage and advantage. In some cases, the reaction cannot be completed under the SPS conditions. If a mixture of several compositions having different properties can be sintered without a compositional change by the diffusion during the sintering, combination of properties, and therefore a design of property will be possible3. The low reaction rate of SPS may enable such combination."

Jan 1, 2000

By Mostafa O. El-Bealy

"To estimate the interdendritic cracking tendency in casting processes, a new model for interdendritic cracking criteria has been proposed. The model combines a concept of the alloying element effect on the interdendritic strain generation and therefore on the susceptibility of interdendritic cracking phenomena. A susceptibility of interdendritic cracking is modeled by using El-Bealy's interdendritic cracking susceptibility function ""BJCSF' and the opening displacement between the dendrites ""EIA "". These concepts are modeled based on the previous experimental observations and ElBealy' approach. The model is implemented numerically for a one-dimensional model problem with some similarities to the casting of steel. The variation of some key parameters that are known to influence the frequency and severity of the interdendritic cracking is then studied.I. INTRODUCTIONInterdendritic cracking ""hot tearing"" commonly encountered in both ferrous and nonferrous castings (1,2]. It is generally believed that interdendritic cracks start to nucleate and develop in the interdendritic strain region within the mushy zone [l-5]. The interdendritic strain region is then definitely coherent, but continuous films of liquid still exit (4,5]. This distinguishes interdendritic cracks from cold cracks forming during the casting process after complete solidification (6,7].Much effort has been put into the understanding of the interdendritic cracking phenomena, and several theories have been proposed. Fellini [8] stated that interdendritic crack ""hot tear"" will result if the material is subjected to too high of accumulated strain within so called vulnerable part of the solidification interval. Kinosita et. al. [9] and Campbell (10] also emphasized the role of tensile stresses in the forming of interdendritic cracks. Most interdendritic cracks criteria simply consider the size of solidification interval, stating that solidification range is associated with a lager interdendritic cracking susceptibility than a short range. Clyne and Davies (11] formulated a more refined such criteria based upon the time spent in different regimes of the solidification interval. They defined a vulnerable region in which thin continuous films of interdendritic liquid exist, and the permeability is low. When thermal strains are induced in this region, the film is not able to sustain the stresses, and an interdendritic crack will form."

Jan 1, 2001

By G. G. Richards

"The slag fuming process has operated for close to 70 years to recover zinc from lead blast furnace slag. The process is normally carried out through the batch treatment of slag by submerged injection of coal and air. Within the bath the coal acts as both a reductant, to reduce Zno to zinc vapour, and a fuel, to maintain bath temperature during the endothermic reduction process. The partition of coal between the slag, combustion and bypass is dependent on factors such as bath depth and injection pressure. The option of continuous fuming has been pursued at the D. Blagoev smelter in Bulgaria using heavy fuel oil as reductant/fuel. There has, however, been little attempt at commercial continuous fuming with coal. This paper presents an analysis of a coal-based continuous slag fuming furnace. The data on the Plovdiv fumer is used as a reference point. Critical design criteria for continuous fuming include the slag feed rate, degree of coal combustion and furnace size. Mathematical modelling analysis suggests that continuous fuming with coal can achieve respectable zinc/coal ratios.IntroductionThe slag fuming process was developed in the early part of this century to recover zinc from non-ferrous smelting slags [1-3]. The majority of installed furnaces were put in place to fume zinc from lead blast furnace slag on a batch basis. In most cases coal is used to effect reduction of zinc oxide, but one furnace uses natural gas [4] and, in a curious twist, a continuous fumer was developed based on fuel oil [5].Given the general nature of blast furnace lead smelting, it is not surprising that slag fuming has remained a batch process, in spite of the advantages that could be gained by moving to continuous fuming. Blast furnaces are usually operated as relatively independent units, and tapping has traditionally been done into ladles. The presence of a crane aisle makes direct coupling of the blast furnace to a slag fumer difficult, especially in a retrofit situation. Furthermore, steady-state blast furnace operation is not easy to achieve. The size of these furnaces and the nature of their feed make it difficult to guarantee constant operation."

Jan 1, 1996