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By M S. Henderson, R D. Alorro, L G. Dyer

An alternative leaching system for rare earth element recovery from monazite is under development. The initial stage utilises oxalic acid to convert the rare earth phosphate into an oxalate salt through dissolution and reprecipitation and it is the efficiency of this stage that governs the overall recovery achieved. It has previously been proposed that lixiviant consumption is the primary limitation in this process. This work further investigates a range of conditions in an attempt to further elucidate the factors influencing the achievable conversion in this stage. It has been shown that the pH may be governed by multiple reactions both consuming and producing acid (thus the mild change over time). The importance of proper mixing and potential to operate at lower temperature with only moderate impact on kinetics has also been demonstrated. Staged/sequential leaching is more effective in fresh solution as opposed to restoring the lixiviant concentration with subsequent reagent addition, suggesting some impact of solution composition. However, contrary to previous findings, subjecting the leach residue to fresh solution resulted in significant decline in leach rate and overall dissolution. This may be because a greater proportion of conversion was achieved in stage 1 in this work than in previous investigation. This may indicate that eventual passivation or the presence of some more refractory REE minerals limits dissolution. Iron dissolution was relatively consistent upon exposure to fresh solution, displaying the ability of the solution to continue to leach the goethite component of the concentrate.

Aug 24, 2022

By C. Dan Kealy

Recent antipollution legislation for the control of mine tailings and waste-water has required an increasing number of tailings ponds. Many of these new waste-retention ponds are being constructed in congested locations, or the older ponds have reached such a size that--if failure occurs--they pose a threat to human life and public property. Although the typical tailings-disposal embankment represents--to the casual-, observer or the economy-minded mine official--a rather ugly but benign and stable pile of waste, they are in fact an extremely sensitive structural mass, retained behind the frailest of dikes. Failure of the waste mass requires only a minor change in water retention or weather conditions, erosion or piping of the ground strata, shifts in load or slight earth tremors. When failures occur, the results can be disastrous. In the U.S., we have been fortunate that, so far, pond failures have not caused loss of life. Thus, failures in this country have attracted little public attention. Present tailings-pond construction methods used by the mining industry require the barest money investment, and little engineering effort. Unfortunately, this superficial approach may increase short-term profits, but the practice may be conducive to the construction of needlessly weak structures. In coming years many of these designs will probaly become corporate liabilities. A typical pond is portrayed in Figure IA. This particular pond in Arizona is 1 mile long and 3/4 mile wide. Illustrated in Figure lB is a representative pond in cross section. This paper in concerned with two basic features common to all tailings disposal: (1) location of the phreatic surface within a tailings-pond embankment because of its influence on embankment-slope stability, and (2) prediction of

Jan 1, 1971

By Adrian C. Deneys, David G. C. Robertson

"Fluid flow phenomena on the surface of a molten slag have been documented using video photography in a laboratory scale D.C. arc furnace. The flow on the surface of the slag appeared to be electromagnetically influenced. This was confirmed by varying the position of the anode, which changed the direction of flow on the surface of the slag. Temperature measurements in the bath also confirmed the change in the directional nature of the fluid flow induced by the arc.During some periods, solids were fed into the arc attachment zone (AAZ) down the center of the hollow graphite electrode. Significant movement of the arc around the basal annular surface of the hollow electrode takes place. The point of attachment of an arc on a hollow electrode was more variable than an arc attached to a solid graphite electrode. Video images showed that the solids were rapidly incorporated into the bulk of the slag. Temperature measurements confirmed the existence of a superheated slag layer at the slag-gas interface, which allowed the solids to be rapidly incorporated into the melt.This work has been conducted to study fluid flow, temperatures, and solid behavior during feeding, in the arc attachment zone of a D.C. arc furnace. The work also has relevance to optimizing furnace design for pyrometallurgical slag cleaning processes."

Jan 1, 1999

By G. W. Zhang

Ultra-fine Stirred Mill is a kind of high energy efficient equipment for ultra-fine grinding, and has been the mainstream equipment in the submicron powder and slurry industry. This paper mainly describes flow-field numerical simulation of the three popular ultra-fine grinding mills: (respective-ly as) Disc Stirred Mill, Pin Stirred Mill and Spiral Stirred Mill, by means of the software Computational Fluid Dynamics (CFD). In the flow-field numerical simulation, the velocity gradient, shear rate dispersion of the flow-field, the viscosity dissipation rate were analyzed with CFD. It is discovered that Pin Stirred Mill has the highest velocity gradient in circular directions and the most homogeneous distribution of shear rate among the three. Flow-field numerical simulation provides a new way (and means to) for the parameter optimization design and project enlargement of Ultra-fine Stirred Mill. The experiments were also carried out, the results are consistent with the results of the simulation analysis. For example, in GCC grinding experiments under the same conditions, the specific energy consumption of Pin Stirred Mill with 3600L in volume is 15%~17% less, and its processing capability is 13%~15% better than Disc Stirred Mill. Now with its optimized structure and excellent performances, more than ten sets Pin Stirred Mill with 3600L in volume have been used in industrial production in Shandong, Guangdong and Hunan province.

Jan 1, 2014

By Manqiu Xu, Peter Wells, Gord Bacon

"A methodology is proposed for milling flowsheet design using flotation rate, selectivity and residence time distribution data. The flotation rate data and the difference in flotation rates between the main minerals to be separated are determined from laboratory batch tests. The residence time distribution (RTD) data is calculated using the tanks-in-series model assuming each flotation cell is a perfect reactor. The rate, selectivity and RTD data can be used to calculate the optimum flotation cell number in a bank for rougher/scavenger in a commercial plant.This methodology is demonstrated in a case study of a nickel sulfide ore. It is shown that miniplant rougher/scavenger data can be closely predicted from a batch test using the proposed methodology.INTRODUCTIONIt has been the trend in the past several decades that large flotation cells are increasingly used. This is largely driven by the economics as ore grades are getting lower and the rate of ore processing must be increased in order to keep mining operations economically viable. The development and evaluation of 250 to 500 m3 WEMCO SmartCell and Outokumpu TankCell flotation machines are well in progress. There are many advantages in using large cells: fewer cells, less floor space, lower power cost (although this is now being questioned for fine particle flotation), simpler circuit, simpler control, and lower auxiliary equipment cost (Arbiter, 1999). As flotation cells are getting larger, the importance of cell hydrodynamics, namely, air/slurry dispersion, power intensity, froth transfer, residence time distribution (RTD) and short-circuiting are being fully recognized (Jonaitis, 1999; Weber, et al., 1999; Gomez and Finch, 2002). It is not the intention of this paper to discuss flotation cell scale-up. The intention is to answer what is the optimum number and size of cells in a flotation bank for a given throughput. Figure 1 illustrates this problem facing design engineers. The same mean residence time required for a given throughput can be achieved in a single large cell, or many smaller cells in a row. A few cells of large size may be more cost-effective but they may not be as effective as more cells of smaller size in achieving the optimum metallurgy."

Jan 1, 2005

By I. Thobadi, C. Pawlik, K. Bisaka, S. Mokoena, M. Erwee

The PyEarth™ process developed at Mintek has been demonstrated successfully at pilot scale. Nine tons of ore containing 2.3 mass % of total rare earth oxides (TREO) from a dormant mineralogically complex iron-rich ore deposit, was subjected to a concentrating smelting step in a 200 kW DC open-arc pilot furnace and the rare earths from the resulting slag were extracted via a hydrometallurgical route. This ore deposit is practically not amenable to extraction by conventional methods. The testwork results proved that the PyEarth™ process was able to optimize the extraction of rare earths while minimizing waste through the removal of iron as a saleable iron alloy. Reduction of iron to the alloy in the smelting step was found flexible, efficient, and controllable, while rare earths were left virtually unreduced in the slag phase. The slag chemistry is selected to optimize both the smelting and leaching performances as well as the alloy composition. Leaching of the resulting slag was found possible and most efficient in hydrochloric acid amongst other lixiviants tested.

Jan 1, 2017

By Aline Tavares

This paper describes the investigative studies carried out at Vale´s Center of Mineral Development (CDM) for a nickel sulfide deposit aiming to develop an adequate flotation flowsheet. The CDM facilities comprises of very well equipped laboratories to develop hydrometallurgical, comminution, flotation, physical separation processes, besides chemical and mineralogical laboratories, that are required to accomplish pre-feasibility studies, new process routes developments and, eventually, brownfield projects. Close association of the flotation group with the geometallurgy and the mineralogy teams play a key role during greenfield project development. Most of the information required by process engineers responsible for the flowsheet development can be provided by the mineralogy team and its mineralogical laboratory at CDM. During the flowsheet development of this nickel sulfide deposit, two ore types were identified as the most abundant in the deposit, named biotite-chlorite hydrothermalite (HID) and magnetite hydrothermalite (MAG HID). Six samples were selected for the investigative studies based on this initial information. Nickel grade varied from 0.39% to 1.95%. The main nickel-bearing minerals were millerite and pentlandite whereas the main non-sulphide gangue minerals were quartz, plagioclase, biotite, chlorite, actinolite, and/or magnetite. Electron microprobe (EMP) identified the presence of nickel in solid solution in gangue minerals such as biotite, chlorite, and actinolite. Nickel deportment in silicates varied from 3% to almost 40%. Magnesium was distributed in biotite and chlorite, except in sample MAG HID, in which talc was responsible for almost 50% of the Mg content. Liberation studies indicated that the optimum feed grind size was around 150µm and that pyrite occurred highly associated to the nickel sulfides. The flotation tests were planned based on the information provided by the mineralogical data. Rougher flotation tests indicated that the optimum P80 of the primary grind was 212µm for the HID type and 150µm for the MAG HID. From the reagents tested, amyl xanthate proved to be the most efficient in the recovery of the sulfides. In the MAG HID sample talc was well depressed by dosing 600g/t of CMC. Nickel recoveries in the rougher stage were lower than 90% due to the nickel distributed in silicates. However, sulfur recovery was around 95%, indicating high recovery of the total sulfides. Open circuit and locked cycle tests were conducted to evaluate the effect of the regrind size in the nickel grade and recovery (P80 of 44µm and 20µm). Nickel recovery decreased from 70% to 65% when finer regrind size was tested. Although nickel recovery decreased by 5%, the finer regrind size lead to a higher Ni grade (from 11% to 23%Ni) and lower fluorine grade (from 1138ppm to 700ppm) in the concentrate. Keywords: nickel sulfide, flotation, mineralogical characterization, flowsheet development

Sep 1, 2012

By M. Bryson, J. Taguta

The Prieska deposit, mined between 1971 and 1991, is one of the world’s top thirty volcanogenic massive sulphide (VMS) base metal deposits. The Prieska Copper Mine in the Northern Cape Province was closed by Anglovaal when underground mining had extended to depths which did not allow for safe and economic extraction of the deposit at the time. The mine has a recorded historical production of over 1 Mt of zinc and 430 k tonnes of copper from 46.8 Mt of milled sulphide ore, achieving life-ofmine metal recoveries of 85% for copper and 84% for zinc into high quality concentrates (refer ORN ASX release 15 Nov 2019). Orion Minerals Ltd is currently investigating the feasibility of developing a new mine in the footprint of the historical Prieska Copper Mine, using the extensive remaining infrastructure, with the intent of extracting the unmined zinc-copper mineralisation from the Prieska deposit. This paper reports results on the flowsheet development philosophy on the hypogene ore, and challenges encountered in the process. This study initially failed to replicate historic copper and zinc recoveries employing the historic production flowsheet and reagent conditions unmodified on bench scale, as the copper and zinc were co-floating. Flowsheet modifications were therefore made to improve differential flotation and recovery, resulting in achieving comparable saleable products and recoveries from processing the hypogene ore on bench scale. Flotation variability test work on the hypogene ore focused on the use of the original flotation procedure and reagent suite using a sodium cyanide-zinc sulphate depressant system on the copper circuit, and it generated a saleable copper product of 20-26% Cu grade at 80-86% recovery. The zinc circuit generated a saleable zinc product grading at 48-55% Zn at 80-94% recoveries. Some zinc losses were due to the deportment of some zinc in sauconite. It is recommended that further test work may further improve the copper grades and recoveries. Keywords: Prieska, flotation, hypogene ore, flowsheet development, sulphide flotation, reagents

Jan 1, 2020

By Manqiu Xu, Joe Holmes, Peter Wells, Peter Langlois, Germain Labonté, Jim Barrett

"The flowsheet for the Voisey’s Bay mill had to meet the challenge of processing an ore containing high pentlandite and chalcopyrite values in a largely pyrrhotite/troilite matrix. In the 1994-97 period, circuit development focused on making a high-grade copper concentrate and a relatively low-grade nickel concentrate. However, in 1998 a circuit with copper concentrate, high-grade nickel concentrate and a Ni/Cu middlings concentrate was specified. Following a three year project delay, work resumed on the flowsheet in 2001, and the target for the overall grade of the nickel concentrate was raised from 14% to 20% nickel. To achieve this target triethylamine tetramine (TETA) and sulfite pyrrhotite depressants were used in an all lime circuit. It later became evident that, with appropriate selection of collector and adjustment of the circuit by the addition of a scavenger on the scavenger cleaner, it would be possible to approach the grade of concentrate required without the use of pyrrhotite depressants. The recycle of water and the contained thiosalts was also proven to be instrumental in raising the grade.Construction and commissioning of the mill were fast-tracked with start-up of the mill in September 2005 some six months ahead of the original schedule. Ramp-up of both throughput and metallurgy was faster than any other recent nickel installations with tonnage being limited mainly by concentrate dewatering due to the higher than initially planned ore grades.INTRODUCTIONNickel was discovered in the vicinity of Voisey’s Bay by Chislett and Verbiski working for Diamond Field Resources in 1993 and the Ovoid deposit was delineated in the winter of 1994/95. By the end of 1995 the Ovoid had been well explored and at that time was thought to have a reserve of at least 30 million tonnes with an average grade of 2.83% Ni and 2.0% Cu. Drilling to the east of the Ovoid had indicated the presence of significant mineralization. In June of 1995, Inco Limited bought a 25% share in the Voisey’s Bay Nickel Company (VBNC) which was a subsidiary holding company set up by Diamond Fields. Inco also bought 8% of Diamond Fields and negotiated the right to sell all the nickel produced from VBNC for twenty years as part of this agreement. In April 1996, Inco bought the 75% of VBNC shares that it did not own."

Jan 1, 2007

By T. K. Mukherjee

The importance of the processing of secondary resources such as solid industrial wastes and liquid effuents as well as byroducts for the recovery of the associated metal values hardly needs any elaboration. Such processing contributes in no small measure.towards the augmentation of the availability and the supply position of resources. As an added benefit, it leads to a reduction in environmental pollution caused by the undesirable waste materials. As some typical examples of secondary resources in the Indian scenario, mention' can be made of spent catalysts, metal/alloy scrap, ash, consumed batteries, flue dust, slag, sludge and slime, pickling solutions, spent acids; plating rinse solutions, spent etchants, waste water from synthetic fibre manufacture and from tanneries etc. These wastes constitute a source of a variety of non ferrous metals. In the processing of secondary resources like these, the techniques of hydrometallurgy such as solvent extraction play a pivotal role. This paper outlines the efforts made to use solvent extraction to recover (1) base metals such as copper, zinc, nickel and cobalt from a number of secondary resources like cake, and sludge(2) refractory metals like molybdenum and vanadium from sulphidic, byproducts and spent catalyst and (3:) nuclear metal like uranium from phosphoric acid. In the case of base metals, process flowsheet involving the use of acidic extractants such as DEHPA and PC88A have been optimised. The problem of iron removal from the leach liquor, "which influences the metal recovery process significantly, has drawn a special emphasis. hi the case of refractory metals, basic extractants .like Alamine 336 has been pressed into service to produce kilogramme. quantities of pure salts of molybdenum and vanadium, Finally, the paper gives a brief account of plant experience in recovering uranium from phosphoric acid, using 'a solvent system comosed of DEHPA and TOPO.

Jan 1, 1996

By T. K. Mukherjee, C. K. Gupta

"The importance of the processing of secondary resources such as solid industrial wastes and liquid effluents as well as byroducts for the recovery of the associated metal values hardly needs any elaboration. Such processing contributes in no small measure towards the augmentation of the availability and the supply position of resources. As an added benefit, it leads to a reduction in environmental pollution caused by the undesirable waste materials. As some typical examples of secondary resources in the Indian scenario, mention can be made of spent catalysts, metal/alloy scrap, ash, consumed batteries, flue dust, slag, sludge and slime, pickling solutions, spent acids, plating rinse solutions, spent etchants, waste water from synthetic fibre manufacture and from tanneries etc .. These wastes constitute a source of a variety of non ferrous metals. In the processing of secondary resources like these, the techniques of hydrometallurgy such as solvent extraction playa pivotal role.This paper outlines the efforts made to use solvent extraction to recover (1) base metals such as copper, zinc, nickel and cobalt from a number of secondary resources like cake, and sludge (2) refractory metals like molybdenum and vanadium from sulphidic byproducts and spent catalyst and (3) nuclear metal like uranium from phosphoric acid. In the case of base metals, process flowsheet involving the use of acidic extractants such as DEHPA and PC88A have been optimised. The problem of iron removal from the leach liquor, which influences the metal recovery process significantly, has drawn a special emphasis. In the case of refractory metals, basic extractants like Alarnine 336 has been pressed into service to produce kilogramme quantities of pure salts of molybdenum and vanadium. Finally, the paper gives a brief account of plant experience in recovering uranium from phosphoric acid, using a solvent system comosed of DEHPA and TOPO."

Jan 1, 1996

By S Vianna, K Caldwell, R Silva, T Olson, D Bell

Studies to enhance the FLSmidth forced air machine through changes in the design, showed important improvements through changes in the number of Rotor blades, stator designs, rotor air distribution and other designs features. The main outcomes from this study highlighted that increasing the number of rotor blades increased the selectivity while decreasing power requirements while stators with slots have better results in terms of recovery, flotation efficiency and selectivity than without slots.Also the rotor air distribution makes a difference in the design when comparing different distribution criteria whilst keeping all features of shape and number of blades etc constant. At the most efficient rotor air distribution design, copper recovery reached one of the highest values while keeping good selectivity. The possible explanation for that is the increase in the pump capacity indicated by the increase in the specific power.It was seen as significant that reduction in specific power for some designs tested compared with the then current standard design had an average of 30% less power consumption.As a result, FLSmidth was able to identify new designs that decreased power and improved recovery, as well as, identifying some designs that specifically improved fine and/or coarse particle recovery.CITATION:Silva, R, Caldwell, K, Bell, D, Olson, T and Vianna, S, 2013. FLSmidth flotation developments - Dorr Oliver machines, in Proceedings MetPlant 2013 , pp 413-430 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 15, 2013

By John E. Litz

Metal-bearing flue dusts are produced by some ferrous and non-ferrous industries. Many dusts are being or can be treated for recovery of metal values. Operating and proposed flowsheets for the most common dusts are reviewed. Advantages, disadvantages, product marketability, and cost implications are discussed for each flowsheet.

Jan 1, 1992

By John S. Lagarias

While systems to remove sulfur oxides from flue gases have been used at coal fired power plants since 1931, critical need for flue gas desulfurization has only occurred since the late 1960s following the passage of the Clean Air Act Amendments. Commercial and advanced flue gas desulfurization processes (FGD) are reviewed. Advantages and disadvantages of the processes are discussed. Regulatory requirements leading to the development of FGD systems are identified. The location and operation of coal fired plants may be determined more and more by the regulatory requirements for sludge disposal, waste, water discharge and air emissions.

Jan 1, 1978

By H. Yoon

Several processes involving injection of sorbent and humidification water downstream of the air preheater in a coal-fired boiler unit, are being developed as low capital cost retrofit S02 control options. Extensive laboratory work, pilot plant programs and larger scale programs have developed performance data and design information. Many of these programs were supported by the U. S. Department of Energy (DOE) as "proofs-of-concept" of three different processes proposed by Dravo, Bechtel and General Electric. Based on these project results, DOE formulated a comprehensive in-duct sorbent injection development program to fully commercialize the technology by 1992. Also, EPRI will conduct a 4 MWe pilot program to develop its HYPAS process. The Consol Coolside process is scheduled for utility demonstration in 1989 on a 104 MWe unit at the Ohio Edison Edgewater Station under partial sponsorship of the DOE Clean Coal Technology Program.

Jan 1, 1989

By A. Kale

A chlorination pilot plant having capacity to produce 5?10 kg/hr of titanium tetrachloride is in operation at the chlorination facility in Mintek. The bubbling fluid bed chlorination plant can be operated in batch and/or continuous process to study the chlorination kinetics of various feedstocks. The chlorination of titaniferrous ores to produce titanium tetrachloride (TiCl4) was carried out during the commissioning of the pilot plant and also as part of the titanium programme carried out at Mintek. TiCl4 is a basic raw material for producing commercial titanium dioxide pigment and titanium metal. This paper discusses the challenges faced during the commissioning and operation of the chlorination pilot plant. It also discusses the effect on the auxiliary facilities and instrumentation as a result of the harsh operating conditions of +1000°C and use of corrosive gases like chlorine and titanium tetrachloride. It discusses the effect of physico-chemical properties of the titanium tetrachloride and iron chlorides. The effect on the chlorinator refractory lining due to heating and cooling of the chlorinator is also discussed. Besides the special provisions made during the design of the chlorinator, this paper also discusses the operational experience and possible solutions to be implemented in future. Keywords Fluid bed chlorination, titanium ore chlorination, titanium tetrachloride (TiCl4) production using fluid bed technology.

Jan 1, 2011

By Casley G. E, Pritchard J

A single slurry fed fluosolids roaster was chosen to replace six multiple hearth roasters in the Copper Smelter at Mount Isa as part of the overall program to increase production from 100 000 to 155 000 tonnes blister copper per annum. The paper describes factors influencing the choice of the unit, problems encountered during the commissioning and details of the present roaster operation.

Jan 1, 1974

By K. H. Heino

Two 200 T/da Lurgi Turbulent Layer fluid-bed roasters have been in operation at Canadian Electrolytic Zinc Limited since 1966. Zinc concentrate containing 52-54% Zn, 9-11% Fe,31-33% S is treated to produce calcine containing 60-62% Zn,10-12% Fe, .l-.3% S as sulphide and .8-1.2% S as sulphate. Each roaster train comprises a waste heat boiler, two cyclones in parallel and a two-field electrostatic precipitator. Distribution of calcine collected is 10-15 percent from bed overflow, 55-65 percent from boilers, 20-25 percent from cyclones and 2-6 percent from precipitators. A Monsanto designed contact plant produces acid at a rated capacity of 365 T/da (100% H2304 basis) from roaster gases diluted to 7.0-7.5% S02. The process steps are gas cleaning and cooling by weak acid scrubbing, acid mist removal in electrostatic precipitators, gas drying in 93.5-96.5% H2304, conversion of sulphur dioxide to sulphur trioxide in a four-pass converter containing vanadium pentoxide catalyst, and absorption of the sulphur trioxide in 98.1-98.3% H2S04. Product acid is 93.2% H2S04. Conversion efficiency is 97-98 percent and overall sulphur recovery from concentrate is 87-90 percent. The adjacent fertilizer plant has a capacity of 135 T/da of P205. The products are 54 percent phosphoric acid, run-of-pile triple superphosphate, and diammonium phosphate. The plant is of Ugine-Kuhlman design and consists of an air-cooled multiple tank reactor, a tilting pan filter, and a forced circultation vacuum evaporator. Triple superphosphate is made on a continuous belt and partially cured in a rotary dryer. Diammonium phosphate is produced by ammoniating phosphoric acid in a single stage reactor. The slurry of ammonium phosphate is granulated in a blunger and dried.

Jan 1, 1970

By K. P. Wright

THREE OF the gold mines in the Red Lake Camp operates fluid bed roasters in their milling circuits. The mines are Campbell Red Lake, Cochenour Willans, and Dickenson. The roasters are used for the autogenous roasting of a flotation concentrate for subsequent gold recovery by cyanidation. Cochenour and Dickenson operate single stage reactors, whilst Campbell use a two-stage system. Ore treated by the mines is similar in that the major part of the gold is associated with arsenopyrite. Other sulphides present are pyrite, pyrrhotite, stibnite and sphalerite. Gangue minerals are mainly quartz, with carbonates and talc. Mineralization is, for the most part, extremely fine, and there is intimate association between sulphides and quartz, so much so, th at a picked specimen of apparently massive arsenopyrite will frequently contain as much as 80 per cent gangue. Gold recovery by straight cyanidation on this type of ore would not be higher than 80 per cent and

Jan 1, 1961

By Kent D. Peaslee

When supersonic oxygen jetting is used in electric furnace steelmaking, the lance jet axis is typically inclined at larger angles to the vertical than is used in basic oxygen steelmaking. The supersonic jet is important not only for refining the steel, but for sustaining the foamy slag in the electric furnace. Physical modeling was carried out using a room temperature, two-dimensional slice apparatus in which mercury represented the metal and water-glycerol the slag. High speed video photography with image analysis was used to study the effects of the angle of inclination on the slag/metal bath. Mathematical modeling using FLUENT provided a method of predicting flow patterns and the trajectories of drops injected at various points in the flowfield. A one-half scale water model of the system, in which lance heights, angles, and flowrates simulated those encountered in steelmaking operations, was used to provide a better understanding of some aspects of the oxygen blowing process in electric furnace steelmaking.

Jan 1, 1994