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By D. K. Bose, C. K. Gupta, K. U. Nair

A new approach to processing an indigenously available low grade molysulfide concentrate by chlorination with elemental chlorine in presence of oxygen is reported. Recovery of nickel and copper values associated with the sulfide concentrate by suitably treating the residue obtained after chlorination is also reported. In the chlorination process, a gaseous mixture of chlorine and oxygen in the volume ratio bf 2:5 has been passed at 1.66 x 10- 5 m3/s for 1800 s over a 0.01 kg batch of molysulfide concentrate at a temperature of 573°K. Most of the molybdenum has been volatilized as molybdenum oxychloride which has subsequently been extracted in water. The molybdenum value from the aqueous solution has been recovered in the form of calcium molybdate (CaO.MoO3) by treatment with calcium chloride (CaCI2). In an alternative processing scheme, the molybdenum rich solution has been treated with an oxidizing agent (K2S20f.J) and allowed to boil to recover a major portion of the molybdenum value in the form of molybdic oxide (MOOS). The spent charge obtained after molysulfide chlorination has been mixed with about 1 wt % sodium chloride and heated at 773K under an oxygen atmosphere. The reacted mass has been water-leached to yield a nickel-copper bearing solution with limited iron. The process liquor appears to be well suited for further processing by known hydrometallurgical recovery processes for nickel and copper. In the conventional process of molybdenum extraction, the molysulfide concentrate (-59% Mo) is first subjected to roasting to produce molybdic oxide. The oxide is then further processed in sequential steps involving vacuum distillation, ammonia leaching, and calcination to higher purity prior to hydrogen reduction to metal. In India there is at present some small tonnage production of molysulfide concentrate from a byproduct recovery plant of the M/s. Uranium Corp. of India Ltd. (M/s UCIL), Bihar. Efforts are being made at this laboratory to process the low grade fraction of the concentrate which contains sulfides of copper, nickel, and iron; quartz; and chlorites. Here, the conventional routes involving roasting and vacuum distillation appear unsuitable to produce an acceptable quality of molybdic oxide at high recoveries. Formation during roasting molybdates of some of the impurities present in the sulfide concentrate make subsequent vacuum distillation less efficient. In view of these problems, in earlier work reported from this laboratory the low grade concentrate has been processed hydrometallurgically to P. produce calcium molybdate. Starting from this molybdenum intermediate, processes have also been developed for the production of molybdenum and ferromolybdenum. Chlorination as an alternative to roasting for metal sulfides has been drawing considerable interest. The high reactivity of chlorine, the relatively low temperature of operation, the suitability of application to low grade resources, the convenience of recovering sulfur in non-polluting form, and the purity of chlorides which are separated by distillation all favor chlorination. In recent years, considerable work has been reported on the chlorination of common metal sulfides. In the field of molybdenum extraction from its sulfide resource, use of chlorination technology has been rather limited. Senderoff and Labrie3 were the first to chlorinate molysulfide concentrate with elemental chlorine at about 823°K to molybdenum pentachloride for use in the electrolytic preparation of molybdenum. The only other reported work, by Haver, Uchida and Wong, involved reaction between molysulfide and chlorine at 673°K to recover sulfur. The reaction products were, however, identified as molybdenum pentachloride and sulfur monochloride.

Jan 1, 1979

By C. E. Nighman, R. S. Foster

The following is a description of the methods used in rescuing men and extinguishing the underground fire at the Pennsylvania mine, Butte, Mont. , This fire, which cost the lives of 21 men, began about 9 p.m., Feb. 14, 1916, at or near a ventilating fan on the 1,200 air-shaft station. This shaft at that time was downcast, while the main hoisting shaft was upcast. Owing to the proximity of the fire to the air shaft, the air in the latter soon became heated, and as the fan on the level was not running— the power having been cut off shortly after the outbreak of the fire—the air shaft became upcast to the 300 level. At this point a blanket of cold air from the surface caused the smoke and gases to be drawn in on that level and carried down again through the workings to the 1,200. A short time after the discovery of the fire, water was turned down the air shaft through a series of pipes around the collar of the shaft, which caused the smoke to recede below the 300, so that within 2 hr. it was possible to search that level for the missing men. From the above it will be apparent why the air shaft was upcast instead of downcast, as under normal conditions, and why men were suffocated at considerable distances from the seat of the fire. The smoke and gases that came up the main hoisting shaft did not diffuse through the workings to any extent. There are doors on all levels in the connections to the main hoisting and air shafts, but in the confusion and running around many of these doors were left open, and this undoubtedly caused the smoke and gases to spread more rapidly than they would otherwise have done. It was necessary for most of the men on the various levels to pass close to the air shaft in going to the hoisting shaft, and when they encountered the smoke and gases in that vicinity a number turned back and climbed down with the direction of the air currents and were then suffocated, while all those who passed through the pall of smoke to the main shaft were saved. There were 220 men in the mine that night, and all but 25 were hoisted to the surface within 30 min. after the discovery of the fire. Of the 25, five went out through the Tramway mine, and one through the Moun-

Jan 1, 1918

By M. E. Wadsworth, L. H. Raby, A. S. Peck

Infrared spec troscopy was used to study the adsorption of oleate collector on three varieties of hematite. Each of the minerals was found to react with either oleic acid or a solution of sodium oleate to form a mineral cation-collector anion chemisorbed layer at the solid-liquid interface. The chemisorption reaction displaced surface hydroxyls and was found to be pH dependent. The infrared band characteristic of the antisymmetrical C = 0 structural group of chemisorbed oleate appears between 1520 and 1540 cm-'. Evidence is presented which shows that molecular filming of the mineral surfaces with un-dissociated oleic acid precedes chemisorption. Infrared absorbance is correlated with the flotation behavior of the mineral. Adsorption processes are usually classified into two types which are distinguished and character- ized by the nature of the bond formed between the adsorbate and adsorbent. The process in which molecules chemically bond to solid surfaces is termed activated adsorption or chemisorption. If only weak van der Waals forces bind the molecules to the surface then the process is termed physical adsorption. Chemisorption is of primary interest in flotation systems since selectivity may be obtained if a single mineral is rendered air-avid by a specific collector mineral adsorption reaction. In previous studies infrared spectroscopy was used to define the mechanisms involved in the adsorption of oleic acid and sodium oleate on fluorite, barite, and calcite. Infrared techniques were particularly useful since adsorbed monolayers were identified in situ. Further, the infrared method permitted identification of reaction products in the solid, liquid, and vapor phases of the systems. The chemisorption reactions displaced F- from fluorite, SO4 from barite, and CO3 from calcite. This investigation was pursued in an effort to characterize and interpret the adsorption processes for synthetic hematite, compact red hematite, and specular hematite (crystalline hematite). Adsorption tests were performed using either oleic acid or

Jan 1, 1967

By F. M. Smith, R. H. Moore, J. R. Morrey

Electrodiffusion in y cerium reported by Henrie has been confirmed and a Preliminary estimate made of the relative rates of electrodiffusion of iron, cobalt, and nickel. These diffuse to the anode at rates decreasing in that order. In addition, copper and manganese exhibit slow, but detectable, diffusion to the anode and molybdenum exhibits detectable diffusion to the cathode. The electrodiffusion of carbon, .zirconium, antimony, magnesium , and silicon in y cerium could not be detected. Iron and cobalt diffuse in y cerium at rates proportional to the current density and with no apparent dependence on temperature. Decreasing polarization of iron and cobalt with increasing temperature, which cancels the expected rate increase, would account for this behavior. The electrodiffusion rate of iron in y uranium and in E plutonium has been measured. Diffusion of iron is anode-directed. Tin was found to diffuse to the cathode, in y uranium, at an appreciable rate. In all of these solvent metals, negative ions diffuse to the anode and positive ions to the cathode. The potential field effect appears to account satisfactorily for these results. FROM early experimental work summarized by Jost1 and Seith,2 the driving force for electrodiffusion was attributed to the potential field acting on ions in a metal. More recently, Heuman,3 Wever,4 and Huntington5 have shown that momentum interchange between conduction electrons and mobile entities in the metal contributes to electrodiffusion. Electron momentum interchange is anodically directed and the direction of diffusion resulting from the field force is dependent upon the charge on the diffusing entity. These two effects may either reinforce or oppose each other. Glinchuk6 has pointed out that momentum transfer in defect conductors should be cathode-directed and this appears to be the case as demonstrated by wever's4 work on iron. Barnett's7 work, on the other hand, indicates that, even in defect conductors, electrons show a negative E/m ratio when accelerated with respect to the lattice and should lead to anode-directed momentum transfer. In discussing this problem, Wever and seith8 suggest that defect electrons interact preferentially with activated ions so as to allow a net movement toward the cathode while still maintaining an electron momentum transfer in the anode direction. Williams and Huffine9 and Henriel0 have demonstrated that electrodiffusion may be useful for purification of yttrium and cerium. In yttrium, Williams and Huffine note that movement of several metallic impurities toward the anode is in keeping with observations in most other metallic systems and indicates that yttrium remains a normal electronic conductor at least to 1230°C. Close inspection of their data shows, however, that oxygen, nitrogen, and the transition elements diffused toward the anode, while nontransition elements diffused toward the cathode. This suggests that potential field effects may have been appreciable. The present work was concerned with the applicability of electrodiffusion as a technique for purification of plutonium, but, because of the obvious hazard inherent in work with this metal, experiments to develop the technique were carried out using cerium and uranium. The results of electrodiffusion measurements on these metals and on plutonium are reported here. EXPERIMENTAL The metal specimens prepared for this work were 6 in. long, 1/4 to 1/2 in. wide, and 1/16 to 3/32 in. thick. The uranium specimens were machined from a bar which analyzed 310 ppm Fe and the electrodiffusion of iron was followed by spectrographic and by chemical analysis. Cerium and plutonium specimens were cut from sheet rolled from ingots obtained from molten salt-metal equilibrations during which radioactive tracers were introduced. The electrodiffusion of the tracers was subsequently determined by counting methods. The specimens were electrolyzed between nickel electrodes containing resistance heaters used to equalize the specimen and electrode temperatures, thereby reducing thermal gradients. The temperature of the electrodes adjacent to the ends of the specimen was measured with chromel-alumel thermocouples which were connected to the heater controls. The surface temperature of the specimen at a point midway between the electrodes was measured with a sapphire rod pyrometer, the output of which controlled the dc power supply. This assembly was enclosed within an evacuable chamber containing a quartz viewing window. The temperature of the specimen over its entire length could be scanned with a portable pyrometer through

Jan 1, 1965

By John M. Lovejoy

NATURAL resources become a source of wealth as they are exploited and made available to the people in usable form. Experience has taught us that Nature does not readily give up her treasures, but the ingenuity of man in the field of techologic progress is required to find and efficiently exploit these resources. History has also demonstrated that with advancing technologic knowledge the. world has become progressively richer in its resources. Those that could not be commercially exploited are now being turned to beneficial use. Through expanding technology and the use of substitutes and synthetics, availability of the world's sources of natural wealth is today many times that of a hundred or a thousand years ago. And so it is with petroleum. Technology is also the great multiplier of our petroleum resources. Through improvements in the art of finding, drilling, and producing, our ability to meet rapidly increasing demands over the last quarter century has been maintained. On the other hand, had our

Jan 1, 1944

By B. C. Raynes

IN the early part of 1952, under the auspices of the U. S. Atomic Energy Commission, Horizons Inc. undertook an investigation dealing with the preparation of high purity thorium metal in order to develop an economically competitive production process for this metal based on a recovery of the metal by a fused salt electrolysis, since a survey of the literature of thorium chemistry indicated that electrolytic methods had not received much attention, despite the consideration that electrochemical production methods often offer attractive commercial potentialities. Thorium metal has been prepared by methods which fall into four general categories: 1) reduction of halides or double halides; 2) reduction of the oxide; 3) thermal decomposition of halides; and 4) electrolytic methods. Some of the important results obtained by various investigators since about 1900 in recorded attempts to prepare pure thorium metal are summarized in Table I. A bibliography of pertinent literature is included in this paper. Thorium is strongly electropositive and possesses a great affinity for oxygen and nitrogen. As a result the prospects for the production of high purity thorium from aqueous media or oxygen bearing compounds are remote. The most promising and profitable area of investigation for production of thorium metal lies in fused salt electrolysis. Thorium has been produced previously by this technique from halide salts in fused alkali chloride eutectic mixtures. Until recently, however, the lack of known large uses for thorium has inhibited the investigation of these methods. The basic problem in an investigation into fused salt production methods for thorium is to develop a suitable fused salt system for the economic preparation of thorium. The first portion of the work in these laboratories was a small scale investigation of the electrolysis of potassium thorium fluoride, KThF5, under a variety of conditions. High purity thorium was prepared in this manner, but the process did not appear to offer the simplest and most attractive continuous commercial operation. The work reported here relates to an all-chloride system employing thorium chloride as the cell feed. ThCL4 was produced as an anhydrous material stabilized in molten sodium chloride. This product, further diluted in NaCl, was subjected to electrolysis. The electrolytic process has been studied in cells capable of holding up to 50 lb total salt charge. Ductile, pure thorium metal has been reproducibly obtained in this operation. Considerable experimentation was carried out in the preparation of ThCL as a cell feed free from moisture, oxides, and other impurities. Thorium was cathodically deposited as a coarsely crystalline powder interspersed with residual bath salts. Aqueous procedures for separating and recovering the metal were developed. The thorium metal powder obtained has been evaluated with respect to its hardness, chemical purity, and certain other physical properties. Preparation of ThC1,—Thorium chloride, and the chlorides of some other metals (aluminum, beryllium, tantalum, zirconium, etc.) cannot be obtained from aqueous solution by direct heating and evaporation. Evaporation and heating of a solution of thorium chloride yields a hydrated salt or partially hydrated salt, and finally Tho,. Anhydrous thorium chloride is obtained by forming a hydrated ammonium chloride complex salt in acidified water solution, dehydrating this compound, and finally decomposing the double salt and subliming the ammonium chloride component. Various methods for the preparation of anhydrous thorium chloride, in accordance with this concept, were investigated as indicated in Table I1 from hydrated thorium chloride (ThCl, . 3H2O), and from thorium oxycarbonate (ThOCO,). In the method found most satisfactory for this work, thorium oxycarbonate was dissolved in an excess of concentrated HC1 to give a clear yellow solution. At least 2 mol NH,Cl were added to the solution per mol of contained thorium. Ammonium thorium chloride hydrate was crystallized from solution by cooling and saturation with HC1. The resultant complex, (NH,Cl),ThCl;xH,O, was dehydrated by evaporation at 130° to 150°C and thoroughly dried at about 250°C. The dried ammanium thorium chloride was then mixed with sodium chloride in proportions to form NaThCL or to produce a ThCL4-NaCl mixture with a given thorium content. The ammonium chloride was sublimed off at temperatures above 500°C in an inert atmosphere furnace to give an anhydrous chloride source electrolyte, essentially free of Tho, and of NKC1.

Jan 1, 1958

By G. L. Oldright

THE development of the reverberatory furnace for smelting copper ores up to 1912 was described by E. P. Mathewson1 with details concerning the great changes in dimensions of the furnace. Hayward2 tabulated similar data, giving developments up to about 1924, in his revision of Hofman's Metallurgy of Copper. DEVELOPMENT OF PRESENT SMELTING PRACTICE Dimensions of Furnace Very great changes were made in the dimensions of the smelting hearths of the furnaces in the period from about 1800 to 1906, the length increasing from about 11 to 116 ft., and the width from 8 to 19 ft. During this time the tonnage smelted increased from about 9 to 250 tons per 24 hr., and the tons of charge smelted per ton of coal increased from 0.77 to 4.3. Dr. Mathewson3 noted no great difference in the "fuel ratio" (i. e., the amount of charge smelted per ton of coal) as the furnace at Anaconda was increased in length by stages from 85 to 116 ft. (The percentage of copper in the slag, however, was lowered from 0.42 to 0.36 per cent. The other components of the slag were not given.) Lump coal fired on grates was the fuel used during this period. The practice in operating the furnaces over such a length of time naturally varied considerably, and there were differences in composition and nature of the charge, yet it may be considered that the improvements made were due largely to increases in the size of the furnace used. Doubt-less some credit should be given to improvements in operating technique, like changes in the methods of firing the lump coal. However, the greater part of the advance was probably due to factors influencing the

Jan 1, 1928

By I. W. Wark

AN attempt has been made to compare the influences of the two most widely used depressants-alkalis and sodium cyanide-and the most widely used activator-copper sulfate-on the air-mineral contact induced by five sulfur-bearing compounds: (1) sodium di-ethyl dithio-phosphate (sodium Aerofloat); (2) potassium ethyl xanthate; (3) sodium di-ethyl dithiocarbamate; (4) potassium iso-amyl xanthate; (5) potassium di-n-amyl dithiocarbamate. Three different organic types are included in this group and for two of these types a higher and a lower homologue have been tested. The study of potassium ethyl xanthate, the preliminary results of which have been described,1 was continued until there had been obtained a comprehensive account of the influence of variations in pH value and in the concentrations of xanthate, sodium cyanide and copper sulfate on the adsorption of this xanthate by galena, sphalerite, pyrite and chalcopyrite. With concentrations of copper sulfate and collector that the work with ethyl xanthate indicated would give the most illuminating results, the other four collectors were tested in a less exhaustive way. The same general principles determine the influence of alkalis, cyanide and copper sulfate, no matter to which type the collector belongs, but there is a gradation in the quantitative measure of such influence. In the list above, the five collectors are cited in the order of their capacity to withstand the depressants, sodium Aerofloat being the least resistant, potassium di-n-amyl dithiocarbamate the most resistant. This order coincides substantially with that of the solubilities of the heavy metal salts formed by double decomposition with the collectors, those formed from sodium Aerofloat being most soluble. There is little difference between the third and fourth members of the series, and under certain conditions their positions are reversed.

Jan 1, 1934

By B. W. Prof. Frazier

(Read at the Wilkes-Barre Meeting, May, 1877.) THIS interesting application of the laws of thermodynamics to metallurgical practice has not been discussed by any writer, within my reading, except the late Prof. Gallon of Paris. In his Cours de Machines,* Prof. Callon has given a short but lucid discussion of this subject in the admirably clear style for which he was remarkable. Although the interest attached to the subject is rather theoretical than practical, I have thought that an attempt to explain the mechanical action exerted in the hot-blast oven might not be entirely devoid of utility, especially in the suggestions which it involves. The statement, that the mechanical work performed in heating the blast is, when high temperatures are produced, considerably greater than the work performed by the blowing-engine, may seem incredible to many who are practically acquainted with the subject, appearing, as it does, to conflict with the results of their experience. It is, nevertheless, literally true, and I shall attempt in this paper to prove it, and to explain why it has not rendered itself palpably manifest in practice. Before proceeding to the mathematical demonstration, it will be well to recall to mind the definitions of some terms employed by writers in mechanics and thermodynamics. The actual energy of a moving body, as defined by Rankine, is the work which it is capable of performing against a retarding resistance before being brought to rest, and is equal to the energy which must be exerted on the body to bring it from a state of rest to its actual velocity. The value of that quantity is the product of the weight of the body into the height from which it must fall to acquire its actual velocity. * Vol. I, chap. x, [§ ] 4, p. 312.

Jan 1, 1878

IT is difficult to judge how much influence the success attained during its first year, 186465, by the School of Mines at Columbia had on developments in education for the mineral industry elsewhere in the immediately subsequent years. The facts are as follows: 1. The Massachusetts Institute of Technology began giving instruction in February 1865, with a mining course included in its curriculum. 2. Early in 1865 the faculty of the Department of Science, Literature and the Arts of the University of Michigan requested permission from the Regents to establish "a course of study to be called the School of Mines . . . Also that the degree of Mining Engineering be conferred on students who should complete the course of study prescribed in the same." A resolution authorizing this was adopted by the Regents on March 28, 1865. 3. Samuel Hooper endowed a Sturgis-Hooper professorship of geology at Harvard in 1865 with the intention of making it a nucleus of a new School of Practical Mining and Geology. 4. Yale appointed Alfred P. Rockwell professor of mining in 1865. 5. Lafayette College appointed Henry S. Osborn professor of mining and metallurgy in 1865. 6. Washington College (as Washington and Lee was then known), rising out of the ashes of Civil War destruction, made a fresh start in 1865, with General Lee as its president, and offered a mining course in the first catalogue thereafter published. 7. Lehigh University, founded in 1866, had a mining curriculum from its beginning. 8. Rensselaer Polytechnic began offering a curriculum in mining and metallurgical engineering in 1867. It would be natural to suppose that these events were a reaction to the success of the school at Columbia, and possibly they

Jan 1, 1941

By C. W. Thompson

Paul D. Suloff (Goodyear Tire and Rubber Co., Inc., Akron, Ohio)—I would like first to comment on problems of the conveyor belt discussed in Mr. Thompson's excellent paper, since that is what we hope we know most about. Twists in relatively wide conveyor belt unavoidably produce a lateral maldistribution of tension, raising tension at belt edges and reducing it at the center. They also produce a lateral collapsing force on the belt at the center of the twist owing to the inherent tendency of all the longitudinal elements of the belt to try to pass through a point at the twist center. Calculation of the twist geometry by the methods shown in Mr. Thompson's paper keeps these extraordinary forces within limits which the belt designer can tolerate. No reduction in belt life due to twisting need be contemplated when this geometry is maintained. There is a minor exception that belts of extreme lateral flexibility will tend to curl laterally at the center of the twist. However, any ordinary fabric construction will perform satisfactorily in this respect. These twists are always made in regions of low tension in the conveyor so that even in the edges of the twist, belt tension does not exceed the average tension found in highly stressed regions of the conveyor. Offsetting these out-of-ordinary belt stresses is the advantage that Mr. Thompson has brought out of getting the return run up out of the dirt where it can be seen. This not only makes it easier to train, but also, in the event that it is not properly trained, frees it of the normal return run edge wear hazard. It is well known that return run edge wear is a prominent cause of belt mortality underground. An interesting aspect of this two-way conveyor is that the belt may be made what is known as a Mobius Strip. A Mobius Strip is obtained by splicing a belt after turning one end of it 180" about its longitudinal axis. In other words, one end is turned upside down before splicing. A belt spliced in this fashion turns itself upside down every time it comes around, but the twist which has been put in the splicing, of course, stays at one location on the conveyor, in this case one of the twist sections at the end. Turning the belt over every revolution might have advantages in some cases. Belts could be made with equal covers and the two sides worn equally and simultaneously. In this case there would be no problem of getting belts on upside down by mistake. However, the two-way conveyor does not have to be a Mobius Strip. It can be twisted in such fashion that the same side is up on both runs. It is simply a question of which way the final 90" twist is made before joining the ends. Another interesting aspect of the two-way conveyor is the problem of operating two-way conveyors in series. Here the sequencing of starting brings up some new problems. It will be recognized, although not always at first glance, that if the starting sequence is planned for one run of the conveyor the reverse will result on the other run. With the two runs carrying bulk material in both directions a reverse sequence on one run would be intolerable. In this situation the only solution appears to be a simultaneous starting of all conveyors in the series. However, with the coal in one direction and intermittent supplies in the other it would be entirely practical to sequence the conveyors for the coal run and accept a reverse sequence on the supply run. The two-way conveyor also lends itself to new driving possibilities. First, it is quite possible to drive at the head end of each run, which of course, means a drive at each end of the two-way conveyor. Driving in this way a given belt can be extended to substantially greater lengths than a conventional conveyor with drive at one end only. In addition to this, under certain conditions the conveyor can be extended to extreme length by driving at one end and at some intermediate point on the most heavily loaded run. As a particular case, a belt carrying coal downgrade and supplies back upgrade could be extended to extreme lengths by driving at the head of the coal run and at an intermediate point of the supply run. Mr. Thompson has been a pioneer in belt conveyor transportation underground and his accomplishment here with the first two-way conveyor of any consequence is another notable addition to the art.

Jan 1, 1954

By M. M. Fine

THE problem of producing a uniform, medium-fine sand for glass-furnace feed has been of interest to the glass-container industry for many years. In the present investigation of the problem, conducted by the Bureau of Mines in cooperation with the Owens-Illinois Glass Co., Alton, Ill., a satisfactory method of producing such sand was developed. The investigation involved continuous-grinding studies of glass sand with flint and porcelain pebbles, steel balls and rods, and a simultaneous study of classification. The most satisfactory procedure consisted of grinding in a pebble or ball mill followed by a double classification: hydraulic classification to separate the oversize for return to the grinding unit and mechanical classification to remove fines and impurities from the ground sand. The proper application of this flowsheet usually yielded a graded sand containing less iron than the feed even when grinding was done in a steel-lined ball mill. The amount of sand ground into the so-called "dusting" sizes was, in all cases, less than 10 pct. Experiments were made which showed that this very fine sand could be beneficiated to a usable material. Background The advantages in utilizing a uniform, medium-fine sand as a feed for glass furnaces has motivated a considerable amount of research on the ,problem of producing such sand by glass producers and others. For the glass producer, the faster melting rate of such a graded sand offers the possibilities of reduced fuel costs and increased furnace capacity. In addition, there is the likelihood of improving the quality of the melt and thereby reducing the number of seeds (air bubbles) and stones (unrnelted grains) in the finished ware. Although the melting rate of a glass batch depends on other factors in addition to the grain size of the sand, there is experimental evidence to show that grain size influences the melting rate considerably. Potts, Brookover, and Burch' found that when the raw materials (sand, soda ash, and limestone) were always matched in- size, there was a great decrease in the melting time as the grain size of the materials was decreased. They found, too, that when the grain sizes of the soda ash and Limestone were held constant at 40 to 60 mesh, the minimum melting time occurred when 40 to 60 mesh sand was used. The increase in melting time with decrease in grain size beyond 40 to 60 mesh was due, they postulated, to accelerated demixing or separation of batch during melting as the grain size of the sand decreased in relation to the grain size of the other materials. An earlier investigatora found that the highest rate of melting and purifying ordinary glass was obtained when medium-size sand (0.385 to 0.120 mm, correspondingly roughly to 40 to 120 mesh) was used. Although batches containing smaller-size sand were found to fuse more rapidly, the time required to remove air bubbles from such melts actually increased the total duration of the process. Coarsegrained sand, it was stated, increased greatly the times of both melting and purifying. Although the laboratory data are difficult to confirm on a plant scale, observations by the Owens-Illinois Glass Co. indicate a higher furnace capacity of better-quality ware when a medium-fine sand is used in the batch. During the recent war, the glass producer was inclined to overlook a considerable variation in raw materials in the interest of higher production. For instance, table I shows the wide range in the amounts of various screen fractions in sand supplied by one producer to the Terre Haute plant of Owens-Illinois. The sand, from the northern Illinois St. Peter sandstone, represented material used by that plant over a period of a few years during the war. Since the war, the glass industry has insisted on

Jan 1, 1951

By M. M. Fine

THE problem of producing a uniform, medium-fine sand for glass-furnace feed has been of interest to the glass-container industry for many years. In the present investigation of the problem, conducted by the Bureau of Mines in cooperation with the Owens-Illinois Glass Co., Alton, Ill., a satisfactory method of producing such sand was developed. The investigation involved continuous-grinding studies of glass sand with flint and porcelain pebbles, steel balls and rods, and a simultaneous study of classification. The most satisfactory procedure consisted of grinding in a pebble or ball mill followed by a double classification: hydraulic classification to separate the oversize for return to the grinding unit and mechanical classification to remove fines and impurities from the ground sand. The proper application of this flowsheet usually yielded a graded sand containing less iron than the feed even when grinding was done in a steel-lined ball mill. The amount of sand ground into the so-called "dusting" sizes was, in all cases, less than 10 pct. Experiments were made which showed that this very fine sand could be beneficiated to a usable material. Background The advantages in utilizing a uniform, medium-fine sand as a feed for glass furnaces has motivated a considerable amount of research on the ,problem of producing such sand by glass producers and others. For the glass producer, the faster melting rate of such a graded sand offers the possibilities of reduced fuel costs and increased furnace capacity. In addition, there is the likelihood of improving the quality of the melt and thereby reducing the number of seeds (air bubbles) and stones (unrnelted grains) in the finished ware. Although the melting rate of a glass batch depends on other factors in addition to the grain size of the sand, there is experimental evidence to show that grain size influences the melting rate considerably. Potts, Brookover, and Burch' found that when the raw materials (sand, soda ash, and limestone) were always matched in- size, there was a great decrease in the melting time as the grain size of the materials was decreased. They found, too, that when the grain sizes of the soda ash and Limestone were held constant at 40 to 60 mesh, the minimum melting time occurred when 40 to 60 mesh sand was used. The increase in melting time with decrease in grain size beyond 40 to 60 mesh was due, they postulated, to accelerated demixing or separation of batch during melting as the grain size of the sand decreased in relation to the grain size of the other materials. An earlier investigatora found that the highest rate of melting and purifying ordinary glass was obtained when medium-size sand (0.385 to 0.120 mm, correspondingly roughly to 40 to 120 mesh) was used. Although batches containing smaller-size sand were found to fuse more rapidly, the time required to remove air bubbles from such melts actually increased the total duration of the process. Coarsegrained sand, it was stated, increased greatly the times of both melting and purifying. Although the laboratory data are difficult to confirm on a plant scale, observations by the Owens-Illinois Glass Co. indicate a higher furnace capacity of better-quality ware when a medium-fine sand is used in the batch. During the recent war, the glass producer was inclined to overlook a considerable variation in raw materials in the interest of higher production. For instance, table I shows the wide range in the amounts of various screen fractions in sand supplied by one producer to the Terre Haute plant of Owens-Illinois. The sand, from the northern Illinois St. Peter sandstone, represented material used by that plant over a period of a few years during the war. Since the war, the glass industry has insisted on

Jan 1, 1951

By Edward J. Way

+1KLEINFONTEIN GROUP CENTRAL ADMlNISTRATION, BENONI, TRANSVAAL, S. AFRICA. The cost and the speed of sinking a shaft are factors of so great importance in operating a mine, that the data given in Table I., herewith published, will be of interest and value to all engaged in practical mining; and for this reason the data are communicated for permanent and accessible record in our Transactions. The figures 'of cost cover the period from January 1, 1903, to May 31, 1903, during which time the total depth sunk was 858 ft. So far as I know, the cost of sinking per foot (£13,15s., 11.04d.) of a shaft of the dimensions 21 ft. by 6 ft. in hard rock, is the cheapest yet on record; and the speed of 213.5 ft. per month, which was attained during May, 1903, is also the best record for speed in sinking in rock so hard as that of this shaft. During the period from January 1 to May 31, 1903, there were 144 rounds worked, involving 4,032 holes, of a total depth of 28,963 ft. 2.4 in., which is equivalent to an average depth of 7 ft. 2.2 in. per hole bored. These data are of added interest from the fact that the work was done practically by native labor.

Jan 1, 1905

Executive Committee, Board of Directors Augustus B Kinzel, Chairman, J L Glllson, Vice-Chairman, A W Thornton, Howard C Pyle, Grover J Holt Finance Committee, Board of Directors Andrew Fletcher, Chairman, J S Vanick, Gall F Moulton Investments Gail F Moulton, '60, Chairman, Henry Krumb, '59, J Loveloy, '61, C R Dodson, ex off as Treasurer, Andrew Fletcher, ex off as Chairman, Finance Com¬mittee Endowment Andrew Fletcher, Chairman, N G Alford, H DeWitt Smith, W J Coulter, J V N Dorr, R B Gilmore, J B Haffner, O H Johnson, D H McLaughlin, J R McMillan, L F Reinartz, C E Weed, Wilfred Sykes, J R Suman Honorary Memberships H DeWitt Smith, '59, Chairman, C E Reistle, Jr, '59, Grover J Holt, '59, L F Reinartz, '59, M L Haider, '60, D H McLaughlin, '61, John M Lovejoy, '62, Augustus B Kinzel, ex off as President

Jan 1, 1958

By J. N. Ong, J. P. Ratledge, J. H. Boyce

Since German operators opened the Tsumeb mine in the early 1900's, continuous operation has been interrupted only by enforced shutdowns during two world wars and the depression of the 1930's. Original metallurgical practice was blast furnace smelting. This was applicable only to low zinc ore, leaving large tonnages of high zinc ore to be stockpiled on the surface. A gravity concentrator using tables, buddles, and jigs operated for several years, but differential flotation of the ore did not get beyond the experimental stage. When the property was put up for sale by the South African Custodian of Enemy Property there was an estimated 434,000 tons of ore on surface dumps. The dump material, assaying 4.48 pct Cu, 15.8 pct Pb, and 10.9 pct Zn, had become highly oxidized during a stay of many years on the surface. But these dump ores had to be used in evaluating possible metallurgy because the shaft became flooded after the 1940 shutdown and underground ore was unavailable. Many companies examined Tsumeb but quickly lost enthusiasm because of the formidable metallurgical problems presented by the dump ore. However, on the basis of extensive test work, the present company decided that differential flotation would be successful and purchased the property after competitive bidding.

Jan 4, 1955

By R. R. Sayers

TWO serious hazards from coal dust confront the bituminous-coal miner- -a physical or safety hazard and a physiological or health hazard. The first threatens the miner with loss of life from coal-dint explosions, and the second with loss of health and earning capacity from inhaling large quantities of the dust. Studies by the Bureau of Mines have shown that coal dust, except some grades of anthracite dust, is capable of initiating an explosion when no explosive gas is present and that its property of extending or propagating an explosion has been responsible for nearly all the widespread explosions that have occurred in bituminous mines in the United States in the present century. For years many mining people believed that low-volatile coal dust was not explosive but not only is it explosive but if the source of ignition is intense, it is as explosive as the dust of many high-volatile coals because the war is creating increased pressure for greater coal production and because of the trend toward greater mechanization in mines, explosion and health hazards from coal dust are likely !n increase considerably.

Jan 1, 1943

By J. W. Stewart

ONE of the postwar problems posed for mineral engineering educators is the desire of some mature student veterans to finish their college education under the G. I. Bill of Rights as quickly as possible, a desire that is quite understandable. Not only have they lost several years of education, which they wish to make up as quickly as possible, but many are married and have families and want to become productive as soon as possible in order to support their families properly. The obvious solution to this problem, if it really is serious, would appear to be a continuation of the wartime Accelerated Program, as it was carried on in most colleges and universities. As a matter of fact some schools are known to have continued their wartime programs in one or another form. To ascertain just what our mineral engineering schools generally are doing to meet this postwar problem, last April a questionnaire was sent to administrative officers at colleges and universities

Jan 1, 1946

By S. L. Hoyt

ABOUT three years ago eight standard steels were prepared for the cooperative investigation of methods for oxygen analysis, sponsored by the Iron and Steel Division of the American Institute of Mining and Metallurgical Engineers and by the National Bureau of Standards. A preliminary report of the results was made at the annual meeting of the A.I.M.E. in February, 1936. ? This report showed the reliability of vacuum-fusion methods and further brought out certain advantages of the fractional vacuum-fusion method. The results obtained by the authors with the latter method will be discussed here, with some com-ments on the procedure for those who are interested in determinations of gases in steel. Some additional comments are offered on the use of the polarizing microscope for identifying the inclusions in the steel to check and amplify the findings of the fractional vacuum-fusion analysis. For details of the fractional method, see the paper by L. Reeve1.

Jan 1, 1937

In the fall of 1937, the Department of Mining at Penn State was being reorganized and the department records were moved to the Records Room of the School. An inventory of the contents of the vault disclosed six folders of records that had been forgotten for 27 years. These files, apparently preserved by former Dean W. R. Crane, plus the recollections of several former members of the Association described below, unfolded the "lost chapter" in mineral industries education.1 It was apparent at an early date in technical education that an association of mining schools with common interests and problems would have definite value. By 1902 sufficient interest had developed in this idea to cause the holding of a preliminary meeting for the purpose of organization. The first meeting was held at the Montana School of Mines on September 4, 1902. At that time, officers were elected and resolutions were adopted that there be formed an organization to be known as the National Association of State Mining Schools. The object and purpose of this Association was to advance the interests of state mining schools in every way, particularly by promoting legislation for the benefit of such schools. The second meeting was held September 10, 1903, and a resolution was passed that the American Mining

Jan 1, 1950