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By H. Mischo, R. Ellmies

Uranium has been mined in Namibia since the 1970s. The worldwide increasing need for energy in the early 2000s has led to the development and opening of a second uranium mine in 2007 and has triggered an increased interest in uranium exploration, often referred to as “the Namibian Uranium Rush.” In all, four mining licenses have been granted by the Namibian Government, which currently has two mines in operation, a third undertaking trial mining and the fourth in project finance negotiations. In addition, highly intensive prospecting activities at additional deposits are at an advanced stage. The finalization of the strategic environmental impact assessment (SEA) on uranium mining in 2010, the first of its kind and scale in the world, has enabled the Namibian government to assess this uranium rush and its tremendous legal, socioeconomic and environmental impacts and to prepare for the different future scenarios conceivable when this article was written in 2011, including both a skyrocketing and a complete breakdown of demand scenario. The Fukushima disaster and plans of the Namibian government to significantly increase royalties and company taxes in 2011 have threatened the market situation, forcing investors to reevaluate Namibian uranium mining projects. However, with the tax and royalty increase initiative since withdrawn, most projects are back on track at this point in time.

Jan 1, 2012

By D. J. Torpy

With copper at its host depressed price in several years, producers are looking at every opportunity to cut costs in order to continue operating. Some of the measures that have been applied by different companies include reduction in manpower, high grading, and in some cases partial shutdowns. The most beneficial cost cutting measure implemented by Southern Peru Copper Corporation was to steepen the slope of the final pit limit from a conservative 37 degrees to an average of 48 degrees. By doing so, substantial cash outlays were saved. Southern Peru Copper Corporation (SPCC) operates two open pit copper mines and concentrators in the southern part of Peru, about 150 kilometers from the Chilean bordar. The Toquepala Mine went into production in 1960 and the Cuajone Mine in 1976. Both mines use shovel, truck and train operations. The trains, which haul ore to the crusher or waste to the dumps, are loaded either directly by shovels or by a transfer system. The Toquepala Concentrator treats an average of 43,000 tons per day and the Cuajone Concentrator an average of 55,000 tons per day. Concentrates are shipped via a 200 kilometer railroad to the port city of Ilo where the SPCC smelter is located. The annual production in the smelter from both mines is 290,000 tons of blister copper.

Jan 1, 1987

By John E. Young

United States Steel Corporation's Winifrede Mine is operated by Lynch District in the mountainous region of Eastern Kentucky. This region is characterized by narrow mountain ridges and precipitous slows which present both an opportunity and challenge in the design and selection of coal handling facilities for the movement of coal from the section belts to the railroad cars. The narrow finger ridges precluded use of underground haulage and, thereby, presented the opportunity to use high capacity equipment on the outside bench. The precipitous mountain slopes, steepest at the Winifrede seam horizon, presented a challenge to select a solution for lowering the coal approximately 1000 feet from the seam level to the valley floor railroad load-out point. It was evident from the start of planning that these two factors predominated; namely, rail haulage and coal lowering, and solutions to these would determine the final design of the Winifrede Mine coal handling facilities.

Jan 1, 1970

By X. L. Song, G. W. Zhang, F. Y. He, X. Xiao, L. L. Huang

This paper presents the design of a model and process simulation of a vertical mill based on a population balance model. The simulator, Simulink, which is a systematic modeling tool within MATLAB, was used for the simulation. The model will simulate the real time variation trend of each size fraction in the feed material to the vertical mill, as well as, the size distribution of the batch grinding products. Breakage distribution and selection functions are calculated based on batch grinding laboratory experiments using Back-calculation method. The simulated size distributions correspond with those in the actual mill operation having a relative error of less than 0.50%. This work lays the foundation for future research on modelling and simulation of a closed circuit grinding operation consisting of a vertical mill with hydrocyclone classifier.

Jan 1, 2016

By Shengfan Zhou, XUEWEN WANG, Bianfang Chen, Mingyu Wang

Vanadium was leached from roasted stone coal by NaOH solution, and the effect of leaching temperature, leaching time, addition of NaOH, liquid to solid ratio, the grinding fineness of stone coal, and stirring speed on the vanadium leaching rate was studied. The kinetics analysis for the process is presented by the shrinking core model. The results show that the vanadium leaching ratio of roasted stone coal can be beyond 75% at the 6% NaOH addition, leaching temperature of 95°C, leaching time of 7h, liquid to solid ratio of 1.2mL/g, and grinding fineness of 0.111mm. The leaching process is controlled by interfacial reaction kinetics and the apparent activation energy was 59.36kJ/mol.

Jan 1, 2016

By Dorsey A. Lyon

"The ceramic industry of the United States has a total yearly manufactured product of nearly half a billion dollars, despite large foreign importation.After careful investigation of the industry by the Bureau of Mines, it is the belief of those in charge that the usefulness of the industry can be greatly enlarged and the importations greatly lessened through the development of processes that will lead to a greater and better development of the ceramic resources of the United States. This will result in a greater industry with more men employed than ever before, and perhaps a world's supremacy in ceramic trade matters.Investigations in behalf of the ceramic industry are being made at the Bureau of Mines experiment ceramic station at Columbus, Ohio, under the supervision of R. T. Stall, superintendent, who reports a number of instances of practical, beneficial cooperation between the industry and the station during the last fiscal year.Graphite to Crucibles:Prior to 1914, practically all graphite crucibles manufactured in the United States were made from Ceylon graphite bonded with German Klingenberg clay. It was the general belief among graphite crucible manufacturers and users that good crucibles could not be made from American bond clays and graphites.Graphite crucibles were essential for melting certain alloys and special steels urgently needed for the manufacture of war materials. The demand for such metals was so great that it created an unprecedented demand for graphite crucibles.After the European war began no Klingenberg clay could be shipped out of Germany. However, it was fortunate for the cyucibly makers that they had large stocks of Klingenberg clay on hand. When the stock of Klingenberg clay became exhausted, it was necessary for the credible makers to turn their attention to American clays. There were no data or other information available to guide the crucible makers in the selection of good bond clays for their crucibles. It was necessary for them to select their clays more or less by ""guess"", which resulted in the period of service or life of graphite crucibles being shortened 15 to as much as 60 per cent."

Feb 1, 1921

By Verônica Scarpini Candido

Este estudo tem por objetivo caracterizar a escória de alto forno proveniente de uma siderúrgica do estado do Pará visando a sua incorporação em argilas para produção de telhas. O resíduo passou pela etapa de beneficiamento e, depois, foi realizada a difração de raios-X (DRX) e fluorescência de raios-X (FRX). Além disso, realizou-se também a análise morfológica por microscopia eletrônica de varredura (MEV). Os resultados mostraram que hematita, magnetita, goethite e wustita são as fases predominantes a associadas aos óxidos de ferro. A análise química por FRX mostrou que o resíduo apresenta elevados teores de sílica (SiO2) e óxidos de ferro associados a baixos teores de alumina (Al2O3). Dessa forma, a caracterização da escória de alto forno mostrou que esse resíduo apresenta-se como um potencial constituinte mássico cerâmico podendo contribuir par ao aumento da refratariedade das peças queimadas.

Oct 30, 2017

By William I. Nissen, Richard Havens

Continuous flotation studies were made at a 50 - lb - per - hr feed rate to confirm the effectiveness of a batch - scale procedure for concentrating phenacite and bertrandite minerals from Mount Wheeler ores that contained , in percent , 0.47 , 0.78 , and 4.7 BeO . The procedure for selective froth flotation of the beryllium minerals from a gangue composed chiefly of calcite , fluorite , and mica comprised closely controlled preconditioning with sodium fluoride and sodium hexametaphosphate in a series of small mechanically agitated tanks and use of a mixture of oleic acid with kerosine , fuel oil , or turpentine as a collector . It was difficult to adapt the laboratory flotation procedure to continuous operation . The principal problem of achieving adequate conditioning to depress calcite and fluorite ultimately was resolved by conditioning in series . Thereafter , three ore samples of different grades treated on a continuous basis yielded concentrates containing from 12 to 25 percent Beo from these relatively complex and difficult ores . Recoveries , on a BeO basis , ranged from 75 to 88 percent . Losses , which primarily were due to incomplete liberation , probably could be lowered in a commercial operation .

Jan 1, 1964

By Carl Rampacek

Laboratory and small-scale continuous pilot-plant investigations made at the Tucson Metallurgy Research Laboratory of the Bureau of Mines have demonstrated that the copper segregation process is applicable to the treatment of oxidized and mixed oxide-sulfide copper ores regardless of the gangue constituents. The process comprises heating the ore to a temperature of 750° C. with small quantities- of sodim chloride and coke to produce metallic copper, which then is recovered by flotation. Tests were made on' 23 copper ores of different types containing 0.8 to 5 percent Cu, .The copper recoveries ranged from 73 to 96 percent in concentrates assaying 10 to 62 percent Cu, depending on the grade and type of ore, The direct operating cost for a 1,000-ton-per-day furnace and flotation plant is estimated at about- $3 per ton of ore.

Jan 1, 1960

By Tianming Sun, Gabriella Tranell, Ragnhild E. Aune, Mark William Kennedy

Rare Earth Elements (REEs) are critical for the production of many advanced materials and products. Mining and primary production of REEs are generally associated with significant environmental impacts. By recovery of REEs from secondary sources, through recycling or as a byproduct, the adverse environmental impacts may be partly avoided. In the present study apatite concentrate originating from iron ore beneficiation done by LKAB in Kiruna, Sweden, is explored as a possible source for REEs. The composition and distribution of the REEs within the concentrates are studied using ICP-MS, XRD and SEM-EDS/EPMA, as well as the melting point and heat of fusion using DTA-TGA-MS. The behavior of the material during melting has also been studied through wetting and slag saturation experiments. The overall aim of the study is to evaluate the implications for future smelting experiments aimed at concentration and recovery of the REEs.

Jan 1, 2015

By Karina Aparecida Martins Barcelos Gonçalves

Amostras dos aços inoxidáveis duplex UNS S31803 (duplex padrão) e UNS S32304 (lean duplex) foram investigadas no estado de fornecimento, após laminação a frio com redução de 64% e 62% e recozimento na temperatura de 1055ºC com 20s de encharque. Foram medidas a dureza Vickers e a quantificação das fases, austenita e ferrita, usando o método de contagem por pontos. A textura cristalográfica foi obtida via técnica de EBSD. A textura cristalográfica da fase ferrítica no estado de fornecimento e recozida apresentou as fibras a //DL e q {100} //DN mais intensas, enquanto a fibra g {111} //DN aparece com intensidade menor. Na fase austenítica as principais componentes encontradas com maior intensidade estão na fibra a //DN e na fibra b com menor intensidade.

Jan 16, 2018

By E. C. Tarpley, R. A. Drogowski, R. F. Abernethy

Three of the test methods frequently used for the determination of moisture in the analysis sample of coal were employed in analyzing a series of coals ranging in rank from anthracite to lignite . For most coals , the variation between the three methods are within the tolerances allowed between duplicate determinations by the same method . The processing variables studied included atmosphere , flow rate of atmosphere , inert atmospheres , oven temperatures , heating time , exposure of the dried coal to laboratory air , and surface area of the test sample . Unless close control is kept on these variables , erratic results will be obtained .

Jan 1, 1964

By C. K. Gupta, P. K. Jena

Niobium (columbium) metal in the form of a button has been produced by calciothermic reduction of niobium pentoxide using sulfur as the heat booster. In these experiments with 50 g of niobium pentoxide per batch, the influence of percentage of calcium in excess of the stoichiometric, the percentage of sulfur by weight of the niobium oxide, and the outer-wall temperature of the bomb on the quality and yield of the metal has been studied. The maximum yield on this scale was 78 pct using 50 pct excess Ca and 20 pct of S and at a bomb-wall temperature of 600°C. In experiments conducted on 500 g of niobium pentoxide, reduction with 40 pct excess Ca and 20 pct of S, and at a bomb-wall temperature of 500°C, the.yield of the metal improved to 82 to 84 pct. Some of the reduced-metal samples have been electron-beam melted and the button melts thus obtained have been observed to be extremely ductile and capable of cold reduction to 98 pct without intermediate annealing. NIOBIUM, because of its high melting point (2468oC), its ductility, its resistance to corrosion especially in liquid metals, its high-temperature mechanical properties, and its relatively low capture cross section (1.1 barns) for thermal neutrons, has potentialities as a structural material in the field of nuclear engineering. Niobium and its alloys are also gaining importance in high-temperature technology as components in aircrafts, jet engines, and missiles, electronics, chemical engineering, and surgery. In 1904 Weiss and Aichel 1 produced niobium by reduction of its pentoxide by misch-metal. Since then a large number of investigations have been carried out for the extraction of niobium from its compounds. The various routes through which the metal niobium has been extracted from its compounds can broadely be classified into the following groups: a) reduction of its halides, oxides, and fluo-salts with metals like calcium, magnesium, sodium, and aluminum;2-10 b) reduction of its halides with hydrogen and oxides with carbon or carbides;11-18 c) electrolytic winning from its halides and fluo salts;19-25 and d) thermal decomposition and disproportion ion of its halides26-30 Among the processes in commercial use at present are sodium reduction or fused-salt electrolysis of potassium niobium fluoride, carbide or carbon reduction of the oxide, and hydrogen or metallothermic reduction of the chlorides. In 1907 Von Bolton2 prepared niobium by reducing niobium pentoxide with aluminum. Dennis and Adam-sona attempted reducing the oxide with magnesium powder in argon atmosphere and the resulting metal contained as much as 5 pct O. Mondolfo8 claimed a method for the reduction of the oxides with aluminum. In the case of both magnesium and aluminum reduction, it has been stated31 that by-product oxides such as MgO or A12O3 would involve a major separation problem. Further, in the case of alumino-thermic reduction, the solubility of aluminum in niobium and intermetallic formation have to be reckoned with. In 1922 Bridge3 produced niobium by calcium reduction of niobium pentoxide. Dennis and pdamson' obtained the metal by a similar method with an oxygen content of 0.2 pct. In most of the cases of metallothermic reduction of niobium oxides mentioned above, the metal niobium is obtained in the form of powder. In an attempt to produce massive niobium metal, lock' used iodine as the thermal booster in the bomb reduction of niobium pentoxide by calcium. Massive niobium metal was obtained with a maximum yield of 75 pct containing 0.1 pct 0. Joly32 carried out experiments on calcium reduction of niobium pentoxide on a 2-kg scale using sulfur as the heat booster in a sealed bomb filled with argon. In his experiments, the reaction was initiated by passing a current through niobium spiral wire embedded in the charge. In a typical run consisting of 2000 g of niobium pentoxide, 3100 g of calcium, and 800 g of sulfur (S/Nb2O5 = 3.3), he obtained massive metal in the form of a very uneven mass with a yield of -57 pct. With lower amounts of sulfur and calcium, powder niobium mixed with small metal globules was obtained and the yield was also poor. From these experiments he concluded that this method is not a suitable one for industrial development considering the quality and yield of metal obtainable and the possible hazards with higher ratios of sulfur to niobium pentoxide in the -~charere. In the present work, a detailed investigation on the calciothermic reduction of niobium pentoxide in a stainless-steel bomb in presence of sulfur as the heat booster was undertaken on a laboratory scale.

Jan 1, 1964

By A. G. Starliper, H. Kenworthy

A procedure was developed for the separation of lead and sulfur in industrial concentrates of galena by a combination of chlorination and fused - salt electrolysis . Galena could not be satisfactorily electrolyzed in the various media tried ; however , when it was treated with chlorine the resulting lead chloride could be effectively separated electrically into lead and chlorine . An apparatus was designed and constructed to chlorinate galena and to volatilize simultaneously most of the chloridized impurities . By electrolyzing the chlorinated concentrate , a high percentage of lead was recovered . Anodic chlorine was drawn from the cell and was used to chlorinate new feed . In many runs , the recovered metal analyzed from 99.8 to 99.9 percent lead . Lead was also distributed as a chloride in the bath , and a small portion was in the volatile fraction produced during the chlorination step . During a tinuous operation , only 0.1 percent of the total lead was discarded ; 99.9 percent was accounted for in the metal , electrolytic bath , and washed or leached chlorination condensate which was essentially lead chloride . About 95 percent of the sulfur from the galena was recovered as a byproduct of over 90 - percent grade .

Jan 1, 1964

By C. B. Kettering

In the early operations of the Inspiration Consolidated Copper Company, which extended from 1914 to 1926, the ore as mined contained principally sulphide copper minerals, manly chalcocite (Cu2S), This ore was milled in a 20,000-ton concentrator and the copper was recovered as a flotation concentrate which, was smelted at the nearby plant of the I ternational Smelting and Refining Company. In those days no attempt was made torecovercopper from any oxide minerals present. As mining operations progressed it soon became evident that the proportion of oxide minerals present would steadily increase, and. it was realized that some method of treatment would have to be developed whereby copper could' be recovered from both oxide and. sulphide minerals contained in the ore. A research program was instituted and after some years of experimental and, pilot plant work a new hydro-metallurgical process was developed whereby it became possible to extract copper from both the oxide and. sulphide minerals in a single leaching operation. The solvent adopted consisted of a mixture of sulphuric acid and ferric sulphate,' Following dissolution, the dissolved copper would be recovered by electrolysis. Also in the process of electrolyzing, spent solutions of sulphuric acid, H2SO4, and. ferric sulphate, Fe2(SOk)3, would be regeneratede and. somade available for continuous reuse. The process appeared to be a success and the decision was made to build a full-sized leaching plant for the future treatment of I spiration mixed ore. Construction was started and the leaching plant was completed in 1926, With the start of ferric-sulphate leaching the Inspiration Concentrator was shut down and, with the exception of a few short time runs has not operated since that time. The ferric sulphate leaching process served. Inspiration well for some 30 years and during that time won a name for itself as a classic example of hydro-metallurgical treatment of mixed sulphide oxide ores. Originally the plant was designed. to process some 7,500 tons per day of a mixed ore containing about 1,2% copper, of which 60% was present as oxide and. 40% as sulphide, Down through the years various improvements were made for the purpose of increasing plant capacity and as well to change the limiting 'factor of sulphide-oxide ratio, The final result was a leaching capacity of 9,600 tons per day when treating a 1% copper ore with a reversal of the sulphide-oxide ratio to 60% sulphide and 40% oxide.

Jan 1, 1958

By M. R. Yenchek

The Bureau of Mines conducted research to determine the relationship between cur-rent load and temperature rises in coal mine trailing cables. Six low-voltage, unshielded, portable power cables were continuously and intermittently loaded with direct current of various magnitudes. Temperature rises within and on the cables were measured with thermocouples, and the data were recorded with a computer. Thermal time constants were calculated, which fixed the periods of the duty-cycle tests. Relationships between average temperature at the conductor-insulation interface and current load were established. The steady-state and intermittent currents that produce a 90° C average conductor-insulation temperature were then determined. Comparisons with Insulated Cable Engineers Association (ICE A) steady-state ratings revealed that 10 to 25 pct more current than recommended ICEA ampacities is required to reach rated insulation temperature. Examination of the maximum intermittent temperatures attained showed that autoignition of coal dust and burn injuries to personnel handling the cable would not be concerns if the temperature at the conductor-insulation interface averaged 90° C.

Jan 1, 1988

By O. F. Kammerer, W. E. Miller, D. H. Gurinsky, J. Sadofsky, J. R. Weeks

Zirconium and titanium inhibit solution mass transfer of steels by liquid bismuth, mercury, and lead. It is shown that in bismuth and mercury, these adsorb on the surface of the steels and subsequently react with nitrogen and possibly carbon from the steels to form inert, adherent surface layers of ZrN, TiN, or TiN + Tic. Data are presented which describe the condition under which thase deposits form. These inhibitors decrease the solution rate of iron into bismuth, and require a higher supersaturation for precipitation of iron from bismuth. USE of the low-melting heavy metals (bismuth, lead, mercury, and their alloys) as coolants has been limited because solution mass transfer of steels occurs in these liquids; i. e., iron dissolves in the hot sections of the heat transfer circuit and deposits in the colder sections. The rate of solution of iron and the temperature coefficient of solubility are sufficiently great to cause complete or partial stoppage by the deposition in the coldest section of a closed circuit in finite time, even though the actual solubilities are extremely low. In the development of the mercury vapor turbine by the General Electric Co., Nerad and his associates1 discovered that the addition of as little as 1 ppm Ti or Zr to magnesium-deoxidized mercury reduced the mass transfer of ferrous alloys by mercury to a negligible amount. Reid2 reported that titanium was detected chemically on the surface of steels contacted with this mercury alloy in amounts varying from 2.0 to 2.6 mg per sq in., the greatest amount being found in the hottest portion of the circuit. Reid stated that the titanium forms the intermetallic compound Fe2Ti by reaction with iron on the surface of the steels. This compound was presumed to be highly insoluble in mercury. More recently, El-gert and Egan3 have reported a greater than 100-fold reduction in the rate of mass transfer of a 5 pet Cr steel by liquid bismuth upon the addition of titanium (in excess of 50 ppm) and magnesium (350 ppm) in the liquid metal, during experiments performed in thermal convection loops* over the temperature differential 700° to 615° C. Also, Shep-ard and his associates' have reported that the addition of titanium to liquid bismuth and Pb-Bi eutec-tic produced a marked decrease in the rates of solution of both iron and chromium from type 410 steel capsules under static conditions. This inhibiting effect increased with repeated reuse of the capsules. Tests performed in this laboratory under carefully controlled conditions have shown that the addition of zirconium and magnesium, or titanium and magnesium, to liquid bismuth or lead greatly reduces the rate of mass transfer of chromium alloy steels and carbon steels in thermal convection loops with a maximum temperature of 550°C.5-9 The present paper will review the data obtained to date at this laboratory on the behavior of iron and steels in contact with liquid bismuth alloys containing titanium or zirconium, and will attempt to explain the role of the above additives in reducing solution mass transfer. Reaction between the Zirconium or Titanium Dissolved In Liquid Bismuth and an Iron or Steel Surface Reaction between Zirconium Dissolved in Bismuth and the Surface of Pure Iron-—A small pure iron crucible (analyzed by the supplier to contain 0.8 ppm N was contacted with bismuth containing approximately 0.1 pet Mg and varying amounts of a radioactive zirconium tracer. The crucible was then inverted at the temperature of contact. The thin residual layer of adherent bismuth was dissolved in cold, concentrated nitric acid. The crucible surface and the solidified bismuth were then analyzed for radioactive zirconium. An analysis of the activity loss on the crucible surface and the weight loss of the crucible during the nitric acid treatment showed that the acid treatment removed the zirconium that had originally been dissolved in the adherent bismuth, but not any zirconium that may have reacted with the crucible surface. The crucible was then pickled in warm aqua regia to remove all surface activity, hydrogen-fired at 600°C, and recontacted with a new liquid alloy. The results of the experiments contacted 1 hr at 450°C show, Fig. 1, a Langmuir-type adsorption with an adsorption free energy of approximately 17 keal per g atom Zr.5 This deposit was estimated to contain 1 atom of zirconium for each 7 to 8 iron atoms on the crucible surface, assuming a surface roughness factor of the pickled crucibles to be five. Increasing the temperature to 520°C caused consi-

Jan 1, 1959

By C. WIFLARD HAYES

(New Haven Meeting, February, 1909.) I DESIRE to make it perfectly clear at the outset that I fully realize the hazardous nature of any attempt to estimate the quantity of iron-ore or any other mineral in the ground. I should not have had the temerity to venture upon an inventory of the iron-ore supplies of the United States except for the requirements of the Conservation Commission. During the preparation of this inventory I have been impressed by the wide diversity in estimates of those best qualified to. hold an opinion, and I am therefore prepared in advance for criticism of my figures. Inasmuch as this report was prepared in large part for unprofessional readers, I considered it necessary to classify and describe the commercial iron-ores in a somewhat elementary manner, in order to explain the reasons for the large element of uncertainty in the estimates. I have therefore given a chemical and a geological classification of iron-ores, which place the commercial deposits in six classes: (1) magmatic segregations in basic igneous rocks; (2) contact-deposits formed in connection with igneous intrusions; (3) concentration-deposits; (4) replacement-deposits; (5) bedded-deposits, and (6) gossan-deposits. The iron-ores are further distinguished under two classes available and not available. This classification is in large measure arbitrary, and it, is evident that it will vary front time to time. In the available class are placed those ores which can be worked at a profit under conditions essentially as they exist at present. Since it depends wholly on the question of costs, the various factors which affect cost will determine the classification of any particular deposit. Actual production, * Abstract of Report prepared for the Conservation Commission. t Chief Geologist, United States Geological Survey. .

Apr 1, 1909

By R. Schaller, Ch. Karcher, A. Thess

"Electron beam evaporation of liquid metals is an innovative technology increasingly used in industry to produce very thin coatings of high purity. In this process the surface of a metal ingot, placed in a water-cooled copper crucible, is heated by bombarding it with a high-energy electron beam gun. The material melts, forming a free surface and eventually starting to evaporate. The rising vapor cloud condensates as a thin film on a moving substrate. The strong energy input at the free surface gives rise to vigorous thermos capillary and buoyancy driven convective motion within the melt, leading to unwelcome heat losses. This paper aims to show how these losses can be reduced by optimizing the geometry of the crucible and by using external magnetic fields. We present both experimental and numerical results.IntroductionElectron beam evaporation of liquid metals [l] is an innovative PVD-technology which is excellently suited for the industrial production of high-quality coatings. In this vacuum process a high-energy electron beam heats up the surface of a metal ingot. The ingot melts and forms a free surface. A schematic view of the arrangement is shown in Figure 1. When the temperature of the free surface exceeds the actual (pressure-dependent) saturation temperature of the ambient gas, the liquid starts to vaporize. The rising vapor cloud finally condensates on a moving substrate located at some distance above. To guarantee a superior quality of the coating the melt is typically confined in a water-cooled copper crucible. The intensive cooling is necessary to avoid chemical reactions between the melt and the crucible walls, that is, crucible materials and their reaction products are practically excluded from evaporation."

Jan 1, 1999

By J. Chipman, G. G. Hatch

One of the important functions of the iron blast furnace is the desulphur-ization of pig iron before it enters the steelmaking furnaces. However, the increasing concentrations of sulphur in the metallurgical coke, source of approximately 90 pct of the sulphur present in the blast furnace charge, and demands for higher rates of production within recent years have increased the need for greater desulphurization within the iron blast furnace. Furnace operators are beginning to look for desulphurizing agents other than blast furnace slag to accomplish the desired degree of desulphurization. A considerable amount of work has been done on desulphurization outside the furnace with soda ash, calcium carbide and various synthetic slags. Whether the desulphurization of pig iron is accomplished wholly inside the furnace or partly inside and the remainder outside, will be determined by the economics involved. Regardless of which is the case, it is believed that it is necessary to have a better understanding of the physical chemistry of desulphurization by blast furnace slags. To this end, it is the object of the present investigation to attempt what is believed to be the first equilibrium study of the distribution of sulphur between liquid pig iron and a wide range of blast furnace slag compositions. Review of Literature There is a considerable amount of information in the literature concerning the desulphurizing power of iron blast furnace slags, the solubility of various sulphides in the slags, and the effect on desulphurization of temperature, of elements dissolved in the liquid iron, and of viscosity. However, there is nothing to indicate that the equilibrium distribution of sulphur between liquid iron saturated with carbon and iron blast furnace slags has been studied experimentally. Wentrupl has made probably the most detailed study of the desulphurization of pig iron to date. He considered that there are three distinct aspects involved, namely: 1. Desulphurization within the blast furnace (by lime and manganese). 2. Subsequent desulphurization by manganese. 3. The effect of subsidiary reactions on the desulphurization by manganese. The experimental work carried out by Wentrup was devoted mainly to obtaining a better understanding of how desulphurization by manganese was accomplished in the mixer and the ladle. Particular attention was given to the part played by carbon, silicon, and phosphorus associated with manganese in the iron, and the effect of temperature on desulphurization. The experimental results indicated that desulphurization by manganese is purely a process of crystallization of manganese sulphide. The addition of silicon to iron melts containing 3.5 pct carbon and less than 0.5 pct manganese had no noticeable effect on desulphurization, but with 1-2 pct manganese the silicon additions improved the desulphurization. Additions of phosphorus also resulted in improved desulphurizati011 by manganese, but the effect was not as marked as in the case of silicon. It was also found that desulphurization by manganese was further improved by lowering the temperature. In order to explain desulphurization inside the blast furnace, Wentrup considered the system iron, sulphur, calcium, oxygen, manganese. (silicon). The distribution of sulphur between the metal and slag was represented by the following equation: (SS) _ (S)Fe + (S)Ca + (S)Mn .... [S] = [s] [1] The parentheses and the brackets represent the equilibrium concentrations in weight per cent of the slag and metal constituents, respectively. Since FeS D (FeS) _ (FeS) (S)Fe LfeS - [FeS] [S] [2] (CaO) + S e (FeO) + (S)Ca _ (FeO)(S)Ca. (S)Ca _ (CaO) Kl = (CaO)[S] [S] ~K1(FeO) [3] Mn + S D (S)Mn (S)Mn (S)Mn K' = [MnpT "1ST = *lIMnJ !4) Substitution of Eq 2, 3, and 4 into Eq 1 resulted in if = L- + *> (Sol + K^ (S) [51 Eq 5 was used to calcu1;lte -f^j and [S] at 1480°C for slags containing 30-50 pct lime, 0.1-2.5 pct iron oxide, 0-26 pct silica, 2 pct sulphur and iron analyzing 1.5 pct manganese. The value for LFaB at 1480°C was found to be equal to 4.5, based on the experimental work of Bardenheuer and Geller.2 The results of the calculations are shown in Table 1. Although the slags are hypothetical and do not represent the range of compositions found in ordinary blast furnace practice, the calculations indicate that lime is effective in controlling desulphurization only if the iron oxide and silica contents of the slag are kept low. Schenck3 did not claim K1 to be a true equilibrium constant, but an empirical value which varied with the silica content of the slag.

Jan 1, 1950